8-K - SSRM - Equibles

UNITED STATES

SECURITIES AND EXCHANGE COMMISSION

WASHINGTON, D.C. 20549

FORM 8-K

CURRENT REPORT

PURSUANT TO SECTION 13 OR 15(d)

OF THE SECURITIES EXCHANGE ACT OF 1934

Date of Report (Date of earliest event reported):

September 29, 2022

SSR Mining Inc.

(Exact name of Registrant as Specified in Its Charter)

British Columbia

(State or Other Jurisdiction of Incorporation)

001-35455

(Commission File Number)

98-0211014

(I.R.S. Employer Identification No.)

6900 E. Layton Ave., Suite 1300, Denver, Colorado USA 80237

(Address of principal executive offices) (zip code)

(303) 292-1299

(Registrant’s telephone number, including area code)

Not Applicable

(Former Name or Former Address, if Changed Since Last Report)

Check the appropriate box below if the Form 8-K filing is intended to simultaneously satisfy the filing obligation of the registrant under any of the following provisions:

¨	Written communications pursuant to Rule 425 under the Securities Act (17 CFR 230.425)
¨	Soliciting material pursuant to Rule 14a-12 under the Exchange Act (17 CFR 240.14a-12)
¨	Pre-commencement communications pursuant to Rule 14d-2(b) under the Exchange Act (17 CFR 240.14d-2(b))
¨	Pre-commencement communications pursuant to Rule 13e-4(c) under the Exchange Act (17 CFR 240.13e-4(c))

Securities registered pursuant to Section 12(b) of the Act:

Title of each class	Trading Symbol	Name of each exchange on which registered
Common Shares without par value	SSRM	The Nasdaq Stock Market LLC

Indicate by check mark whether the registrant is an emerging growth company as defined in Rule 405 of the Securities Act of 1933 (§ 230.405 of this chapter) or Rule 12b-2 of the Securities Exchange Act of 1934 (§ 240.12b-2 of this chapter).

Emerging growth company ¨

If an emerging growth company, indicate by check mark if the registrant has elected not to use the extended transition period for complying with any new or revised financial accounting standards provided pursuant to Section 13(a) of the Exchange Act.

ITEM 8.01. Other Events.

SSR Mining Inc. (the “Company”) previously filed the Çöpler District Master Plan 2021 Technical Report Summary, the Marigold 2021 Technical Report Summary, the Seabee 2021 Technical Report Summary and the Puna 2021 Technical Report Summary, each with an effective date of December 31, 2021 and issued on February 23, 2022 (each, a “Technical Report Summary” and collectively, the “Technical Report Summaries”), as Exhibit 96.1, Exhibit 96.2, Exhibit 96.3 and Exhibit 96.4, respectively, to our Annual Report on Form 10-K for the year ended December 31, 2021, as amended. The Company has amended the Technical Report Summaries to reflect certain revisions in compliance with Subpart 1300 of Regulation S-K, which revisions consist of adding confirmatory statements and other modifications that the Company does not consider material. The Çöpler District Master Plan 2021 Technical Report Summary has also been amended as described in the Company’s Current Report on Form 8-K filed earlier on September 29, 2022 (the “Previous Form 8-K”) to separate the initial assessment into a standalone technical report summary filed with such Previous Form 8-K. Each such amended Technical Report Summary has been reissued as of September 29, 2022 and is presented with an effective date of December 31, 2021 (each such amended Technical Report Summary, an “Amended Technical Report Summary” and collectively, the “Amended Technical Report Summaries”) and the information in the Amended Technical Report Summaries has not been updated to reflect events, information or developments occurring after the effective date. The Amended Technical Report Summaries are being filed as Exhibit 96.1, Exhibit 96.2, Exhibit 96.3 and Exhibit 96.4 to this Current Report on Form 8-K and the corresponding consents of the qualified persons with respect to the filing of the Amended Technical Report Summaries are being filed as Exhibit 23.1, Exhibit 23.2, and Exhibit 23.3, and each is incorporated herein by reference.

ITEM 9.01. FINANCIAL STATEMENTS AND EXHIBITS

(d) Exhibits

Exhibit Number	Description of Exhibit
23.1	Consent of Gregory Gibson.
23.2	Consent of Bernard Peters.
23.3	Consent of Sharron Sylvester.
96.1	Çöpler District Master Plan 2021 Technical Report Summary.
96.2	Marigold 2021 Technical Report Summary.
96.3	Seabee 2021 Technical Report Summary.
96.4	Puna 2021 Technical Report Summary.
104	Cover Page Interactive Data File (embedded within the Inline XBRL document)

SIGNATURE

Pursuant to the requirements of the Securities and Exchange Act of 1934, as amended, the Registrant has duly caused this report to be signed on its behalf by the undersigned hereunto duly authorized.

SSR Mining Inc.
By:	/s/ Michael J. Sparks
Name:	Michael. J. Sparks
Title:	Executive Vice President and Chief Legal & Administrative Officer

Dated: September 29, 2022

Document

Exhibit 23.1

CONSENT OF EXPERT

I hereby consent to the use of and reference to my name, Gregory Gibson, SME Registered Member, and the information set forth in the Puna 2021 Technical Report Summary, as reissued as of September 29, 2022, which such information is effective as of December 31, 2021, that I reviewed and approved as described or incorporated by reference in (i) SSR Mining Inc.’s Current Report on Form 8-K filed on September 29, 2022, (ii) SSR Mining Inc.’s Annual Report on Form 10-K for the year ended December 31, 2021, as amended on Form 10-K/A, and (iii) SSR Mining Inc.’s Registration Statements on Form S-8 (File No. 333-219848, 333-185498, 333-196116, 333-198092, 333-248813, 333-259280, and 333-265661), filed with the United States Securities and Exchange Commission.

Dated this 29th day of September, 2022.

Yours very sincerely,

/s/ “Gregory Gibson”

Gregory Gibson, SME Registered Member

Document

Exhibit 23.2

CONSENT OF EXPERT

I hereby consent to the use of and reference to my name, Bernard Peters, BEng (Mining), FAusIMM, and the information set forth in each of the Çöpler District Master Plan 2021 Technical Report Summary, the Marigold 2021 Technical Report Summary, the Seabee 2021 Technical Report Summary and the Puna 2021 Technical Report Summary, each as reissued as of September 29, 2022, which such information is effective as of December 31, 2021, that I reviewed and approved as described or incorporated by reference in (i) SSR Mining Inc.’s Current Report on Form 8-K filed on September 29, 2022, (ii) SSR Mining Inc.’s Annual Report on Form 10-K for the year ended December 31, 2021, as amended on Form 10-K/A, and (iii) SSR Mining Inc.’s Registration Statements on Form S-8 (File No. 333-219848, 333-185498, 333-196116, 333-198092, 333-248813, 333-259280, and 333-265661), filed with the United States Securities and Exchange Commission.

Dated this 29th day of September, 2022.

Yours very sincerely,

/s/ “B F Peters”

Bernard Peters, BEng (Mining), FAusIMM

Technical Director – Mining

OreWin Pty Ltd.

Document

Exhibit 23.3

CONSENT OF EXPERT

I hereby consent to the use of and reference to my name, Sharron Sylvester, B.Sc. (Geol), RPGeo AIG, and the information set forth in each of the Çöpler District Master Plan 2021 Technical Report Summary, the Marigold 2021 Technical Report Summary, the Seabee 2021 Technical Report Summary and the Puna 2021 Technical Report Summary, each as reissued as of September 29, 2022, which such information is effective as of December 31, 2021, that I reviewed and approved as described or incorporated by reference in (i) SSR Mining Inc.’s Current Report on Form 8-K filed on September 29, 2022, (ii) SSR Mining Inc.’s Annual Report on Form 10-K for the year ended December 31, 2021, as amended on Form 10-K/A, and (iii) SSR Mining Inc.’s Registration Statements on Form S-8 (File No. 333-219848, 333-185498, 333-196116, 333-198092, 333-248813, 333-259280, and 333-265661), filed with the United States Securities and Exchange Commission.

Dated this 29th day of September, 2022.

Yours very sincerely,

/s/ “S T Sylvester”

Sharron Sylvester, B.Sc. (Geol), RPGeo AIG

Technical Director – Geology

OreWin Pty Ltd

Document

Exhibit 96.1

Explanatory Note

SSR Mining Inc. (the “Company”) previously filed the Çöpler District Master Plan 2021 Technical Report Summary (the “CDMP21TRS”), with an effective date of December 31, 2021 and issued on February 23, 2022, as Exhibit 96.1 to its Annual Report on Form 10-K for the year ended December 31, 2021, as amended. The CDMP21TRS has been amended to reflect certain revisions in compliance with Subpart 1300 of Regulation S-K, which revisions consist of adding confirmatory statements and other modifications that SSR does not consider material. The CDMP21TRS has also been amended to separate the initial assessment originally contained therein into a standalone Çöpler District Mineral Resource 2021 Technical Report Summary. The amended CDMP21TRS has been reissued as of September 29, 2022 and is presented with an effective date of December 31, 2021. The information in this amended CDMP21TRS has not been updated to reflect events, information or developments occurring after the effective date.

Mineral resources and mineral reserves presented in the CDMP21TRS reflect the Company’s actual ownership in the various deposits. Cash flow analyses and supporting information referenced in the CDMP21TRS are presented on a 100% ownership basis.

This page does not constitute a part of the amended CDMP21TRS.

Title Page

Project Name:	Çöpler District Master Plan 2021
Title:	Çöpler District Master Plan 2021<br>Technical Report Summary
Location:	Erzincan Province, Turkey
Effective Date of Technical Report Summary:	31 December 2021
Effective Date of Mineral Resources:	31 December 2021
Effective Date of Mineral Reserves:	31 December 2021

Qualified Persons (QPs):

•Bernard Peters, BEng (Mining), FAusIMM (201743), employed by OreWin Pty Ltd as Technical Director - Mining, was responsible for the overall preparation of the CDMP21TRS and, the Mineral Reserve estimates, Sections 1 to 5; Section 10; and Sections 12 to 25.

•Sharron Sylvester, BSc (Geol), RPGeo AIG (10125), employed by OreWin Pty Ltd as Technical Director - Geology, was responsible for the preparation of the Mineral Resources, Sections 1 to 3; Sections 6 to 9; Section 11; and Sections 22 to 25.

OreWin Pty Ltd ACN 165 722 574

140 South Terrace Adelaide 5000

P +61 8 8210 5600 E orewin@orewin.com W orewin.com

Signature Page

Project Name:	Çöpler District Master Plan 2021
Title:	Çöpler District Master Plan 2021<br>Technical Report Summary
Location:	Erzincan Province, Turkey
Date of Signing:	29 September 2022
Effective Date of Technical Report Summary:	31 December 2021

/s/ Bernard Peters

Bernard Peters, Technical Director – Mining, OreWin Pty Ltd, BEng (Mining), FAusIMM (201743)

/s/ Sharron Sylvester

Sharron Sylvester, Technical Director – Geology, OreWin Pty Ltd, BSc (Geol), RPGeo AIG (10125)

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TABLE OF CONTENTS

1EXECUTIVE SUMMARY	1
1.1Introduction	1
1.2Mineral and Surface Rights	6
1.3Accessibility, Climate, Local Resources, Infrastructure and Physiography	6
1.4History	6
1.5Geological Setting and Mineralisation	6
1.6Exploration	8
1.7Drilling	8
1.8Sampling Method, Approach and Analyses	9
1.9Data Verification	9
1.10Metallurgical Testwork	10
1.10.1Oxide Testwork	10
1.10.2Sulfide Testwork	10
1.11Mineral Resources	11
1.11.1Resource Modelling	11
1.11.2Reasonable Prospects for Eventual Economic Extraction	14
1.11.32021MR Initial Assessment	14
1.11.4Mineral Resources Estimates	17
1.12Mineral Reserves Estimates	20
1.13Mining Method	23
1.14Recovery Methods	24
1.14.1Sulfide Plant	24
1.14.2Oxide Ore Heap Leach Processing	26
1.14.3Project Infrastructure	28
1.15Market Studies	28

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1.16Environmental and Permitting	28
1.17Capital and Operating Costs	29
1.17.1Capital Costs	29
1.17.2Operating Costs	30
1.18CDMP21TRS Reserve Case	31
1.19Interpretation and Conclusions	34
1.20Recommendations	35
2INTRODUCTION	36
2.1Terms of Reference	36
2.2Qualified Persons	37
2.3Qualified Persons Property Inspection	37
2.4Units and Currency	38
2.5Effective Dates	38
3PROPERTY DESCRIPTION	39
3.1Location	39
3.2Ownership	42
3.3Mineral Tenure	43
3.4Surface Rights	45
3.5Taxation	45
3.6Royalties	46
3.7Environmental Liabilities	47
3.8Permits	47
3.9Other Significant Factors and Risks	48
4ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	49
4.1Accessibility	49

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4.2Local Resources and Infrastructure	49
4.3Climate	50
4.4Hydrogeology	50
4.4.1Existing Data Evaluation, Field Investigation, and Hydrogeology Conceptual Model	51
4.4.2Well Installation	51
4.5Physiography	52
5HISTORY	54
5.1Previous NI 43-101 Technical Reports	54
6GEOLOGICAL SETTING, MINERALISATION, AND DEPOSIT	56
6.1Geological Setting – Çöpler Deposit	57
6.1.1Geology – Çöpler Deposit	57
6.1.2Mineralisation – Çöpler Deposit	60
6.1.3Structure – Çöpler Deposit	64
6.2Geological Setting – Çakmaktepe Deposit	65
6.2.1Geology – Çakmaktepe Deposit	65
6.2.2Mineralisation – Çakmaktepe Deposit	66
6.3Geological Setting – Ardich Deposit	68
6.3.1Geology – Ardich Deposit	68
6.3.2Mineralisation – Ardich Deposit	69
6.4Geological Setting – Bayramdere Deposit	69
6.4.1Geology – Bayramdere Deposit	69
6.4.2Mineralisation – Bayramdere Deposit	69
6.5Geological Setting – Regional Prospects and Targets	69
6.5.1Geology – Çöpler Saddle	70
6.5.2Geology – Meşeburnu and Elmadere	71

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6.5.3Geology – Mavialtin Porphyry Belt Prospects	71
6.6Deposit Types	73
7EXPLORATION	74
7.1Exploration – Çöpler Deposit	74
7.1.1Geological Mapping – Çöpler Deposit	74
7.1.2Geochemical Sampling – Çöpler Deposit	75
7.1.3Geophysics – Çöpler Deposit	75
7.2Exploration – Çakmaktepe Deposit	76
7.2.1Geological Mapping – Çakmaktepe Deposit	76
7.2.2Geochemical Sampling – Çakmaktepe Deposit	76
7.3Exploration – Ardich Deposit	77
7.3.1Geological Mapping – Ardich Deposit	77
7.3.2Geochemical Sampling – Ardich Deposit	77
7.4Drilling	78
7.4.1Drilling – Çöpler Deposit	78
7.4.2Drilling – Çakmaktepe Deposit	81
7.4.3Drilling – Ardich Deposit	82
7.4.4Drilling – Mavialtin Porphyry Belt Prospects	84
7.4.5Grid Coordinate Systems	84
7.4.6Collar and Down-hole Surveys	84
8SAMPLE PREPARATION, ANALYSES, AND SECURITY	87
8.1Sample Collection	87
8.1.1Reverse Circulation Drilling Sample Collection	87
8.1.2Diamond Drilling Sample Collection	88
8.1.3Drillhole Logging and Data Collection	88
8.2Sample Preparation	89

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8.2.1Reverse Circulation Sample Preparation	89
8.2.2Diamond Drilling Sample Preparation	89
8.3Sample Analysis	89
8.4Sample Security	90
8.5QA/QC Procedures	90
8.6QP Opinion	90
9DATA VERIFICATION	91
9.1Çöpler Deposit Data Verification	91
9.2Çakmaktepe Deposit Data Verification	91
9.3Ardich Deposit Data Verification	92
9.3.1Data Verification – Ardich	92
9.3.2Collar Location – Ardich	92
9.3.3Down-hole Surveys – Ardich	92
9.3.4Geology, Density, and Geotechnical Logs – Ardich	93
9.3.5Assays – Ardich	93
9.3.6Witness Samples – Ardich	93
9.3.7Quality Assurance / Quality Control (QA/QC) Results – Ardich	94
9.3.8Discussion – Ardich	96
9.4Bayramdere Deposit Data Verification	96
9.5QP Opinion	96
10MINERAL PROCESSING AND METALLURGICAL TESTING	97
10.1Oxide Ore for Heap Leaching	97
10.1.1Testwork – Çöpler Oxide	97
10.1.2Testwork – Çakmaktepe Oxide	97
10.1.3Testwork – Ardich Oxide	97
10.1.4Testwork – Bayramdere Oxide	97

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10.1.5Heap Leach Gold Recovery	98
10.2Sulfide Ores	99
10.2.1Historical Testwork – Çöpler Sulfide	99
10.2.2Sulfide Mineralogy	100
10.2.3Direct Cyanidation	101
10.2.4Flotation Testwork	101
10.2.5Testwork – Comminution	102
10.2.6Testwork – POX	103
10.2.7Testwork: Pyrite Recovery from Copper-Rich Ores	103
10.2.8Overall Circuit Performance	104
10.3Mineral Processing and Metallurgical Discussion	107
10.4QP Opinion	108
11MINERAL RESOURCES ESTIMATES	109
11.1Çöpler Deposit	109
11.1.1Çöpler Mineral Resource Estimate – Key Assumptions	109
11.1.2Çöpler Base Indicator Model	111
11.1.3Çöpler Domains	111
11.1.4Çöpler Geological Model	111
11.1.5Çöpler Data Summary	112
11.1.6Çöpler Exploratory Data Analysis	112
11.1.7Çöpler Top Cutting	114
11.1.8Çöpler Drillhole Compositing	115
11.1.9Çöpler Variography	115
11.1.10Çöpler Resource Model Estimation	115
11.1.11Çöpler Sulfur Model	115
11.1.12Çöpler Gold and Other Metal Models	116

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11.1.13Çöpler Density Model	117
11.1.14Çöpler Oxidation Model	118
11.1.15Çöpler Model Validation	119
11.1.16Çöpler Mineral Resource Classification	119
11.1.17Çöpler Model Validation	121
11.1.18Çöpler Assessment of Reasonable Prospects of Eventual Economic Extraction	121
11.1.19Çöpler Deposit Mineral Resource Tabulation	121
11.2Çakmaktepe	121
11.2.1Çakmaktepe Domains	121
11.2.2Çakmaktepe Data Summary	123
11.2.3Çakmaktepe Exploratory Data Analysis	123
11.2.4Çakmaktepe Top Cutting	124
11.2.5Çakmaktepe Resource Model Estimation	125
11.2.6Çakmaktepe Density Model	125
11.2.7Çakmaktepe Model Validation	126
11.2.8Çakmaktepe Comparison to Production Data	126
11.2.9Çakmaktepe Mineral Resource Classification	127
11.2.10Çakmaktepe Assessment of Reasonable Prospects of Eventual Economic Extraction	128
11.2.11Çakmaktepe Mineral Resource Tabulation	128
11.3Ardich	128
11.3.1Ardich Geological Model	129
11.3.2Ardich Structural Interpretation	129
11.3.3Ardich Lithological Interpretation	132
11.3.4Ardich Mineralisation	132
11.3.5Ardich Database Extract	134

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11.3.6Ardich 2021 Resource Modelling Dataset Summary	134
11.3.7Ardich Exploratory Data Analysis	135
11.3.8Ardich Core Recovery	140
11.3.9Ardich Top Cutting	141
11.3.10Ardich Drillhole Compositing	143
11.3.11Ardich Resource Model Estimation	145
11.3.12Ardich Density Model	149
11.3.13Ardich Resource Classification	152
11.3.14Ardich Model Validation	155
11.3.15Ardich Change of Support	156
11.3.16Ardich Assessment of Reasonable Prospects of Eventual Economic Extraction	156
11.3.17Ardich Mineral Resource Tabulation	156
11.4Bayramdere Deposit	156
11.4.1Bayramdere Domains	157
11.4.2Bayramdere Geological Model	157
11.4.3Bayramdere Data Summary	158
11.4.4Bayramdere Drillhole Compositing	158
11.4.5Bayramdere Top Cutting	159
11.4.6Bayramdere Cell Model	159
11.4.7Bayramdere Estimation Method	159
11.4.8Bayramdere Density Model	159
11.4.9Bayramdere Resource Classification	160
11.4.10Bayramdere Validation	161
11.4.11Bayramdere Assessment of Reasonable Prospects of Eventual Economic Extraction	161
11.4.12Bayramdere Mineral Resource Tabulation	161

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11.5Assessment for Reasonable Prospects for Economic Extraction	161
11.5.1Gold Mineral Resources	162
11.5.2Çöpler District Initial Assessment	164
11.6Initial Assessment Inputs	175
11.7Mineral Resources Statement	192
11.8Comparison with Previous Estimates	195
11.9QP Opinion	195
11.10Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects	196
12MINERAL RESERVES ESTIMATES	197
12.1Summary	197
12.2Mineral Reserves Statement	197
12.3Comparison with Previous Estimates	200
12.4Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects	201
13MINING METHODS	202
13.1Geotechnical	202
13.1.1Pit Slope Stability – Çöpler	202
13.1.2Review of 2021 Geotechnical Studies	202
13.1.3RQD Model	209
13.1.4Pit Slope Design Parameters	210
13.1.5Mine Operations Monitoring and Management	210
13.1.6Geotechnical Domains	211
13.1.7Pit Dewatering	212
13.2Mine Plan	212
13.2.1Ore Definition	213
13.2.2Ore Cut-off Grades	216

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13.2.3Pit Design	217
13.2.4Waste Dump and Stockpile Design	218
13.2.5Ore Stockpiles, Rehandle and Blending	222
13.2.6Grade Control	224
13.3Mine Production Schedule	224
13.3.1Scheduling Assumptions	225
13.3.2Production Schedule	226
13.3.3Processing Schedule	228
14PROCESSING AND RECOVERY METHODS	231
14.1Sulfide Ore Processing	231
14.1.1Sulfide Plant Performance	233
14.1.2Sulfide Plant Description	233
14.2Oxide Heap Leach Processing	241
14.2.1Oxide Heap Leach Performance	242
15INFRASTRUCTURE	244
15.1Introduction	244
15.1.1Existing Infrastructure	245
15.1.2Flotation Building	246
15.2Site Water Management	246
15.2.1Hydrology Background	246
15.2.2Site-Wide Surface Water Hydrology	247
15.2.3Surface Water Management Structures	247
15.2.4Fresh Water Supply	247
15.2.5Potable Water Treatment	248
15.2.6Waste Management	248
15.3Power to Site	249

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15.4Emergency Backup Power	249
15.5Communications	249
15.6Site Roads	249
15.7Plant Fire Protection System	249
15.8Site Water Management	250
15.8.1Hydrology Background	250
15.9Heap Leach Facility	250
15.9.1Heap Leach Pad Development	250
15.10Tailings Storage Facility	251
15.10.1TSF Development and Summary of Current Operations	251
15.10.2Site Classification	252
15.10.3Monitoring and Inspection	258
15.10.4TSF Design	258
15.10.5Seismic Deformation Evaluation	259
15.10.6Tailings Consolidation and Capacity	260
15.10.7TSF Schedule Assumptions	260
15.10.8Further Work	261
16MARKET STUDIES	262
16.1Marketing and Meta Prices	262
16.2Contracts	262
16.3QP Opinion	262
17ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS	263
17.1Environmental Studies and Material Impacts	263
17.2Physical Features	264
17.2.1Land Use	265

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17.2.2Biological Features	266
17.3Social and Community Plans	269
17.4Mine Closure	270
17.4.1Closure Cost Estimate Assumptions – Waste Rock Dumps	270
17.4.2Closure Cost Estimate Assumptions – Pits	270
17.4.3Heap Leach Pad	271
17.4.4Tailings Storage Facility	271
17.4.5Other	272
17.4.6Monitoring	272
17.4.7Closure Planning	272
17.4.8Construction Management	272
17.4.9Human Resources	272
17.4.10Closure Schedule	272
17.4.11Further Work	273
17.5Sustainability	273
17.5.1Stakeholder Engagement	273
17.5.2Health and Safety	274
17.5.3Training and Development	274
17.5.4Industrial Relations	274
17.5.5Diversity and Inclusion	275
17.5.6Sustainable Community Development	275
17.5.7Environmental Management	277
17.5.8Water Risk	277
17.5.9Energy and Climate Change	277
17.5.10Tailings Dam Management	277
17.5.11Water Management	278

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17.5.12Cyanide Management	278
17.5.13Biodiversity	278
17.5.14Air Quality	279
17.6QP Opinion	279
18CAPITAL AND OPERATING COSTS	280
18.1Capital Costs	280
18.2Operating Costs	281
18.3Mining Cost Summary	282
18.4Processing and Infrastructure Cost Summary	282
18.5General and Administration Cost Summary	283
19ECONOMIC ANALYSIS	285
19.1Economic Assumptions	285
19.1.1Metal Prices	285
19.1.2Taxation	285
19.1.3Royalties	285
19.1.4QP Opinion on Inputs	286
19.2Reserve Case Economic Analysis Results	287
19.2.1Project Cash Flow	292
20ADJACENT PROPERTIES	294
21OTHER RELEVANT DATA AND INFORMATION	295
22INTERPRETATION AND CONCLUSIONS	296
23RECOMMENDATIONS	298
23.1Mineral Resources	298
23.2Mineral Reserves	299
24REFERENCES	300
25RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT	303

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TABLES

Table 1.1CDMP21TRS Reserve Case Results Summary	5
Table 1.2Gold Price Sensitivity	5
Table 1.3CDMP21TRS Results Summary	16
Table 1.4Gold Price Sensitivity	16
Table 1.5Summary of CDMP21TRS Mineral Resources Estimates Exclusive of Mineral Reserves as at 31 December 2021 Based on $1,750/oz Gold Price	18
Table 1.6Summary of Cut-off Values, Metallurgical Recoveries, and SSR Ownership of CDMP21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on Gold Price $1,750/oz, Silver Price $22.00/oz Ag and Copper Price $3.95/lb	19
Table 1.7Summary of CDMP21TRS Mineral Reserves Estimates (as at 31 December 2021) Based on $1,350/oz Gold Price	21
Table 1.8Summary of Cut-off Values, Metallurgical Recoveries, and SSR Ownership of CDMP21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price	22
Table 1.9Summary of LOM Average Operating Costs	30
Table 1.10CDMP21TRS Reserve Case Results Summary	31
Table 1.11CDMP21TRS Reserve Case Before and After-Tax NPV	32
Table 1.12CDMP21TRS Reserve Case Cash Costs	34
Table 1.13CDMP21TRS Economic Analysis Metal Price Assumptions	34
Table 3.1Granted Licences and Operating Permits	44
Table 3.2Gold Royalty Rates	47
Table 7.1Number of Geochemical Samples within the Çakmaktepe Deposit	77
Table 7.2Number of Geochemical Samples within the Ardich Deposit	77
Table 7.3Drilling History – Çöpler Deposit	80
Table 7.4Drilling History – Çakmaktepe Deposit	81
Table 7.5Drilling History – Ardich Deposit	83
Table 7.6Drilling History – Mavialtin Porphyry Belt Prospects	86

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Table 10.1Çöpler Gold Recovery Assumptions for Heap Leaching of Oxide	98
Table 10.2Çakmaktepe Gold Recovery Assumptions for Heap Leaching of Oxide (incl. Bayramdere)	98
Table 10.3Ardich Gold Recovery Assumptions for Heap Leaching of Oxide	99
Table 10.4Gold Deportment in Flotation Separated Streams	101
Table 10.5Gold POX Recovery Model Parameters	106
Table 10.6Commissioning and Ramp-up Allowances	107
Table 11.1Çöpler Block Model Parameters	111
Table 11.2Çöpler Top Cuts for Au	114
Table 11.3Çöpler Top Cuts for Non-Au Elements, Applied Globally	114
Table 11.4Çöpler Upper and Lower Density Limits by Lithology	117
Table 11.5Density Values Assigned to the Çöpler Cell Model by Lithology and Vertical Depth Below Surface	117
Table 11.6Çakmaktepe Top Cuts for Au, Cu, and Ag	124
Table 11.7Çakmaktepe Cell Model Parameters	125
Table 11.8Çakmaktepe Upper and Lower Density Limits by Lithology	126
Table 11.9Density Values Assigned to the Çakmaktepe Cell Model by Lithology	126
Table 11.10Interpreted Fault Trends	130
Table 11.11Ardich Drilling History in 2021 Resource Modelling Dataset	134
Table 11.12Key Element Statistics of Uncomposited Drillhole Data (length weighted)	135
Table 11.13Au Summary Statistics by Lithology based on Uncomposited Samples (length weighted)	136
Table 11.14Sulfur Summary Statistics by Lithology based on Length Weighted Uncomposited Samples	140
Table 11.15Au Top Cut Values and Effect on Statistics	142
Table 11.16Drillhole Au Statistics for Uncomposited (Raw) and Composited (Comp.) Sample Data – Jasperoid	144

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Table 11.17Drillhole Au Statistics for Uncomposited (Raw) and Composited (Comp.) Sample Data – Listwanite	145
Table 11.18Ardich Cell Model Prototype Parameters	145
Table 11.19Ardich Search Volume Criteria for Grades Estimation in All Domains	148
Table 11.20Gold Variogram Parameters	149
Table 11.21Sulfur Variogram Parameters	149
Table 11.22Ardich Search Volume Criteria for Density Estimation in All Domains	151
Table 11.23Bayramdere Cell Model Prototype Parameters	159
Table 11.24Bayramdere Density Values for Mineralisation Domains	160
Table 11.25Bayramdere Density Values for Lithology Domains	160
Table 11.26Summary of Key Parameters Used in 2021 Conceptual Pit Shell at Çöpler	162
Table 11.27Summary of Key Parameters Used in Conceptual Pit Shell at Çakmaktepe	163
Table 11.28Summary of Key Parameters Used in Conceptual Pit Shell at Ardich	163
Table 11.29Summary of Key Parameters Used in Conceptual Pit Shell at Bayramdere	164
Table 11.30l Assessment Case Process Types	168
Table 11.31Heap Leach Recoveries	168
Table 11.32Copper Concentrate Recoveries	168
Table 11.33Pyrite Concentrate Recoveries	168
Table 11.34POX Plant Recoveries	168
Table 11.35Operating Costs	169
Table 11.36Copper Concentrator and NaSH Capital Cost	169
Table 11.37Other Capital Costs	169
Table 11.38nomic Analysis Metal Prices Assumptions	170
Table 11.39Transport and Treatment Charges	170
Table 11.40Payable Metal Assumptions	170

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Table 11.41Capital and Reclamation Costs	172
Table 11.42Average Operating Costs Unit Rates	172
Table 11.43Summary of LOM Average Operating Costs	172
Table 11.44Operating Costs	175
Table 11.45Copper Concentrator and NaSH Capital Cost	176
Table 11.46Other Capital Costs	176
Table 11.47Çöpler Initial Assessment MII Case Production Schedule	178
Table 11.48Çöpler Initial Assessment MII Case Copper Concentrator Processing Schedule	179
Table 11.49Çöpler Initial Assessment MI Case Production Schedule	180
Table 11.50Çöpler Initial Assessment MI Case Copper Concentrator Processing Schedule	181
Table 11.51Initial Assessment MII Case Cash Flow	182
Table 11.52Initial Assessment MI Case Cash Flow	183
Table 11.53Initial Assessment Results Summary – MII Case	185
Table 11.54Initial Assessment Before and After-Tax NPV – MII Case	186
Table 11.55Initial Assessment Results Summary – MI Case	190
Table 11.56Initial Assessment Before and After-Tax NPV – MI Case	191
Table 11.57Summary of CDMP21TRS Mineral Resources Estimates Exclusive of Mineral Reserves as at 31 December 2021 Based on $1,750/oz Gold Price	193
Table 11.58Summary of Cut-off Values, Metallurgical Recoveries, and SSR Ownership of CDMP21TRS Mineral Resources Estimates Exclusive of Mineral Reserve (as at 31 December 2021) Based on Gold Price $1,750/oz, Silver Price $22.00/oz Ag and Copper Price $3.95/lb	194
Table 11.59EOY 2020 Mineral Resources Summary – Exclusive of Mineral Reserves (as at 31 December 2020)	195
Table 12.1Summary of CDMP21TRS Mineral Reserves Estimates (as at 31 December 2021) Based on $1,350/oz Gold Price	198
Table 12.2Summary of Cut-off Values, Metallurgical Recoveries, and SSR Ownership of CDMP21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price	199

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Table 12.3EOY 2020 Mineral Reserves Summary as at 31 December 2020	200
Table 13.1Golder Intact Strength Estimates	203
Table 13.2Çöpler Mine Pit Slope Parameters	210
Table 13.3Heap Leach – Gold Recoveries	213
Table 13.4Heap Leach – Silver Recoveries	213
Table 13.5Heap Leach – Copper Recoveries	214
Table 13.6Oxide Operating Costs	214
Table 13.7Plant Throughput Limits	214
Table 13.8POX – Gold Recovery Parameters	215
Table 13.9Sulfide Operating Costs	215
Table 13.10Au Cut-off Grade Revenue and Realisation Assumptions	216
Table 13.11Internal Au Cut-off Grades	217
Table 13.12Waste Rock Dump (WRD) Design Factor of Safety (FOS)	221
Table 13.13Waste Rock Geochemical Classification	222
Table 13.14CDMP21TRS Reserve Case Mining Schedule	227
Table 13.15CDMP21TRS Reserve Case Production Schedule	230
Table 16.1CDMP21TRS Economic Analysis Metal Price Assumptions	262
Table 18.1Capital and Reclamation Costs	281
Table 18.2Average Operating Costs Unit Rates	281
Table 18.3Summary of LOM Average Operating Costs	281
Table 18.4Other Capital Costs	284
Table 19.1CDMP21TRS Economic Analysis Metal Price Assumptions	285
Table 19.2Gold Royalty Rates	286
Table 19.3CDMP21TRS Reserve Case Results Summary	288
Table 19.4CDMP21TRS Reserve Case Before and After-Tax NPV	288
Table 19.5CDMP21TRS Reserve Case Cash Costs	290

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Table 19.6CDMP21TRS Reserve Case Metal Price Assumptions	291
Table 19.7CDMP21TRS Reserve Case Gold Price Sensitivity	291
Table 19.8CDMP21TRS Reserve Case Cost Sensitivity	292
Table 19.9CDMP21TRS Reserve Case Project Cash Flow	293

FIGURES

Figure 1.1Project Location Map	1
Figure 1.2CDMP21TRS Reserve Case Boundary	4
Figure 1.3Structural Setting of Anatolia	7
Figure 1.4CDMP21TRS Ultimate Pit Designs	23
Figure 1.5CDMP21TRS Reserve Case Mining Production	24
Figure 1.6Process Flow Sheet for Sulfide Plant	25
Figure 1.7Flotation Block FlowDiagram	26
Figure 1.8Process Flow Sheet for Oxide Ore Heap Leach	27
Figure 1.9Reserve Case After-Tax Cash Flow	32
Figure 1.10CDMP21TRS Reserve Case Processing	33
Figure 1.11CDMP21TRS Reserve Case Gold Production	33
Figure 3.1Location of the Project	40
Figure 3.2Çöpler Project Licence and Surrounding Licences (UTM Grid)	41
Figure 3.3Ownership	42
Figure 3.4Tenure Layout Plan	45
Figure 4.1Average Monthly Rainfall for Çöpler Project Area	50
Figure 4.2Groundwater Wells	52
Figure 6.1Geological Setting of the Çöpler District	56
Figure 6.2Geological and Structural Map of the Çöpler District	57
Figure 6.3Çöpler Deposit Geological Map	59

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Figure 6.4Çöpler Deposit Porphyry Vein Mineralisation	61
Figure 6.5Çöpler Deposit Intermediate Sulfidation Epithermal Mineralisation	62
Figure 6.6Çöpler Deposit Porphyry Vein Mineralisation	63
Figure 6.7Simplified Schematic of the Çöpler Deposit Structures (cross-section)	65
Figure 6.8Geological Map of the Çakmaktepe and Ardich Deposits	66
Figure 6.9Geological Mapping within Çakmaktepe Central Pit	67
Figure 6.10Schematic of Ardich Geological Setting with Mineralisation Examples	68
Figure 6.11Çöpler District Exploration Projects	70
Figure 7.1Çöpler Deposit Map of Alteration Minerals	75
Figure 7.2Çakmaktepe Deposit – Example East Pit Geological and Structural Map	76
Figure 7.3Drillhole Collar Location Plan – Çöpler Deposit	79
Figure 7.4Drillhole Collar Location Plan – Ardich, Çakmaktepe, and Bayramdere Deposits	82
Figure 7.5Drillhole Collar Location Plan – Ardich	83
Figure 7.6Drillhole Collar Location Plan – Mavialtin Porphyry Belt Prospects	85
Figure 10.1Feed SS% – Mass Pull Relationship	102
Figure 10.2Metasediment Gold Recovery Results and Model	105
Figure 10.3Diorite Gold Recovery and Model	105
Figure 10.4Manganese Diorite Gold Recovery and Model	106
Figure 11.1Çöpler Model Zones	110
Figure 11.2Çöpler Lithology – Marble Zone, Cross-Section 459,700 mE (looking west)	112
Figure 11.3Çöpler Drill Core Showing Colour Change from Oxide to Sulfide	119
Figure 11.4Projected Plan View of Çöpler Resource Classification	120
Figure 11.5Çakmaktepe Model Domains (oblique view)	122

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Figure 11.6Ardich Geology Schematic (long-section)	129
Figure 11.7Ardich Structural Framework: Interpreted Faults and Resultant FAULTZONs	131
Figure 11.8Schematic Example Cross-Section of Ardich Lithology Model (looking north-west)	132
Figure 11.9Ardich Drill Collar Location Plan	135
Figure 11.10AU_PPM Box and Whisker Plot by Lithology (MODLITH)	137
Figure 11.11Jasperoid AU_PPM Box and Whisker Plot by FAULTZON	138
Figure 11.12Example Cross-Section (looking north-west) through FAULTZONs 19.1 and 10.2 showing AU_PPM Differences in Adjacent FAULTZONs	139
Figure 11.13S_PCT Box and Whisker Plot by MODLITH	140
Figure 11.14Histogram of Raw Sample Lengths	143
Figure 11.15Plan View of Jasperoid Resource Classification	153
Figure 11.16Plan View of Listwanite Resource Classification	154
Figure 11.17Plan View of Upper Dolomite Resource Classification	155
Figure 11.18Bayramdere Geology Schematic Section	158
Figure 11.19Çöpler Project Initial Assessment Simplified Process Flow Diagram	176
Figure 11.20Çöpler Preliminary Copper Concentrator Process Flow Diagram	167
Figure 11.21Çöpler Plan Initial Assessment Pit Shell, Resource Shell, and Reserve Pit Design	177
Figure 11.22Çöpler Long-section Initial Assessment Pit Shell, Resource Shell, and Reserve Pit Design	177
Figure 11.23Initial Assessment After-Tax Cash Flow – MII Case	186
Figure 11.24Initial Assessment Processing – MII Case	187
Figure 11.25Initial Assessment Gold Production – MII Case	187
Figure 11.26Initial Assessment MII Case Gold Production	188
Figure 11.27Initial Assessment After-Tax Cash Flow – MI Case	191
Figure 11.28Initial Assessment Processing – MI Case	191

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Figure 11.29Initial Assessment Gold Production – MI Case	192
Figure 13.1Rock Mass Quality Mapping Locations (2015–2017)	204
Figure 13.2Rock Mass Quality Mapping of Main Pit	205
Figure 13.3Geotechnical Units	206
Figure 13.4Locations of Seepage and Ponding	208
Figure 13.5Çöpler VWP Data	209
Figure 13.6Typical 15 m Pre-Split at Çöpler Mine	211
Figure 13.7Ultimate Pit Designs – CDMP	218
Figure 13.8CDMP21TRS Reserve Case Site Plan	220
Figure 13.9CDMP21TRS Reserve Case Mining Production	225
Figure 13.10CDMP21TRS Reserve Case Processing Schedule	229
Figure 13.11CDMP21TRS Reserve Case Gold Production and Recovery	229
Figure 14.1Çöpler Process Flow Sheet for Sulfide Plant	231
Figure 14.2Flotation Block Flow Diagram	232
Figure 14.3Gold Recovery and Throughput Comparison	233
Figure 14.4Process Flow Sheet for Sulfide Plant	235
Figure 14.5Heap Leach Process Flow Sheet	243
Figure 15.1Çöpler Project Plan	244
Figure 15.2Mine Water Supply Well Locations	248
Figure 15.3Phase 4 – Top of Embankment and Impoundment Grade	253
Figure 15.4Phase 5 – Top of Embankment and Impoundment Grade	254
Figure 15.5Phase 6 – Top of Embankment and Impoundment Grade	255
Figure 15.6Phase 7 – Top of Embankment and Impoundment Grade	256
Figure 15.7TSF 2 Impoundment Grading	257
Figure 17.1Current Land Use Types and Cadastral Map for the Çöpler Project	267
Figure 17.2Land Use Capability Classes (LUCC)	268

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Figure 19.1CDMP21TRS Reserve Case After-Tax Cash Flow	287
Figure 19.2CDMP21TRS Reserve Case Processing	289
Figure 19.3CDMP21TRS Reserve Case Gold Production and Recovery	290

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1 EXECUTIVE SUMMARY

1.1 Introduction

The Çöpler District Master Plan 2021 Technical Report Summary (CDMP21TRS) is an independent Technical Report Summary that has prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300) for SSR Mining Inc. (SSR), on the Çöpler project (the Project).

The Project is located in east central Turkey, 120 km west of the city of Erzincan, in Erzincan Province, 40 km east of the iron-mining city of Divriği (one-hour drive), and 550 km east of Turkey’s capital city, Ankara (Figure 1.1). The nearest urban centre, İliç, (approximate population 3,800), is located approximately 6 km north-east of the Çöpler mine.

Figure 1.1 Project Location Map

SSR, 2020

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SSR is a gold mining company with four producing assets located in the USA, Turkey, Canada, and Argentina, and with development and exploration assets in the USA, Turkey, Mexico, Peru, and Canada. SSR is listed on the NASDAQ (NASDAQ:SSRM), the Toronto Stock Exchange (TSX:SSRM), and the Australian Stock Exchange (ASX:SSR).

The Çöpler project is owned and operated by Anagold Madencilik Sanayi ve Ticaret Anonim Şirketi (Anagold). SSR controls 80% of the shares of Anagold, Lidya Madencilik Sanayi ve Ticaret A.Ş. (Lidya), controls 18.5%, and a bank wholly owned by Çalık Holdings A.Ş., holds the remaining 1.5%. Exploration tenures surrounding the Project area and mining at Çakmaktepe are subject to joint venture agreements between SSR and Lidya that have varying interest proportions. SSR controls 50% of the shares of Kartaltepe Madencilik Sanayi ve Ticaret Anonim Şirketi (Kartaltepe) and 30% of Tunçpinar Madencilik Sanayi ve Ticaret Anonim Şirketi (Tunçpinar). The remaining 50% of shares for the Kartaltepe and 70% of shares of Tuncpinar are controlled by Lidya.

The CDMP21TRS Qualified Persons (QPs) have reviewed the supplied data and information and accept this information as being accurate and complete and suitable for use in the CDMP21TRS. Information and data supplied by SSR that were outside the areas of expertise of the QPs and was relied upon when forming the findings and conclusions of this report are detailed in Section 25. Any individual or entity referenced as having completed work relevant to the CDMP21TRS, but not identified therein as a QP, does not constitute a QP.

The CDMP21TRS should be construed in light of the methods, procedures, and techniques used to prepare the CDMP21TRS. Sections or parts of the CDMP21TRS should not be read in isolation of, or removed from their original context.

The key features of the CDMP21TRS are:

•Updated Mineral Reserves on the Çöpler deposits.

•A new Mineral Reserve on the Ardich deposit.

The Mineral Reserves are supported by feasibility study level work on the currently operated pits at the Çöpler and Çakmaktepe deposits, the brownfield Ardich deposit, and the oxide heap leach facility and sulfide plant in the Reserve Case.

Mineral Resources are discussed and disclosed in the separate Technical Report Summary titled: Çöpler District Mineral Resource 2021 Technical Report Summary (2021MR). The 2021MR includes an Initial Assessment prepared to demonstrate the further economic potential of the Mineral Resources at the Çöpler deposit with the inclusion of a copper concentrator.

The Initial Assessment is preliminary in nature, it includes Inferred Mineral Resources that are considered too speculative geologically to have modifying factors applied to them that would enable them to be categorised as Mineral Reserves, and there is no certainty that this economic assessment will be realised. The Initial Assessment from the 2021MR is described in Section 11.5.2.

The Mineral Resource for the Ardich deposit has benefited from additional drilling and an updated model in 2021. The mining of Ardich requires development of a new open pit that is approximately 6 km east of the current Çöpler pit and 1 km north of the Çakmaktepe pits.

A plan showing facilities location and the boundary of the Reserve Case is shown in Figure 1.2.

The key production and economic analysis from the CDMP21TRS are shown in Table 1.1.

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The economic analysis uses long-term metal price assumptions of $1,600/oz gold, $21.00/oz silver, and $3.40/lb copper. These prices are based on a review of consensus price forecasts from financial institutions and similar studies recently published.

The economic analysis was prepared on a 100% project basis using the Reserve Case production schedule, operating, and capital assumptions on an annual basis. The Reserve Case production includes 22.6 Mt at 1.69 g/t Au oxide ore processed by heap leaching and 52.9 Mt at 2.33 g/t Au processed in the sulfide plant on a 100% project basis. Total production is 75.4 Mt at 2.14 g/t Au. Total gold production is 4.4 Moz. Mining at the Çöpler pit is completed in 2029 and at Ardich in 2034. Oxide heap leach stacking is completed in 2034, while sulfide processing will continue from stockpiles until 2042.

The Reserve Case results include:

•After-tax net present value (NPV) at a 5% real discount rate is $1.73 billion

•Mine life of 21 years

An internal rate of return (IRR) is not reported as the operation is cash positive in each year of the mine plan until closure. The Reserve Case average all-in sustaining cost (AISC) is $966/oz gold. Key results of the CDMP21TRS Reserve Case are shown in Table 1.1.

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Figure 1.2 CDMP21TRS Reserve Case Boundary

Anagold, 2022

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Table 1.1 CDMP21TRS Reserve Case Results Summary

Item	Unit	Reserve Case
Oxide Processed
Heap Leach Quantity	kt	22,557
Au Feed Grade	g/t	1.69
Sulfide Processed
Quantity Milled	kt	52,892
Au Feed Grade	g/t	2.33
Total Processed
Processed	kt	75,448
Au Feed Grade	g/t	2.14
Total Gold Produced
Oxide – Gold	koz	765
Sulfide – Gold	koz	3,604
Total – Gold	koz	4,369
Oxide – Gold Recovery	%	61
Sulfide – Gold Recovery	%	91
5-Year Annual Average
Average Gold Produced	kozpa	278
Free Cash Flow	$Mpa	158
Total Cash Costs (CC)	$/oz gold	880
All-in Sustaining Costs (AISC)	$/oz gold	1,071
Key Financial Results
Life-of-Mine (LOM) CC	$/oz gold	803
LOM AISC	$/oz gold	966
Site Operating Costs	$/t treated	45.91
After-Tax NPV5%	$M	1,732
Mine Life	years	21

5-Year Annual Average is for the period 1 January 2021 through 31 December 2026

The after-tax net present value (NPV) sensitivity to metal price variation is shown in Table 1.2 for gold prices from $1,000–$2,000/oz.

Table 1.2 Gold Price Sensitivity

After-Tax NPV5% (M)	Long-Term Gold Price(/oz)
			1,350		1,600		1,750		2,000
CDMP21TRS Reserve Case	769	1,115		1,370		1,732		1,939		2,252

All values are in US Dollars.

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1.2 Mineral and Surface Rights

Anagold holds the exclusive right to engage in mining activities within the Çöpler project area. Anagold holds six granted licences covering a combined area of approximately 16,600 ha. Mineral title is held in the name of Anagold. Kartaltepe holds eight licences covering approximately 9,200 ha. The total near-mine tenement package is approximately 25,800 ha. Anagold currently holds sufficient surface rights to allow continued operation of the mining operation in the CDMP21TRS Reserve Case.

1.3 Accessibility, Climate, Local Resources, Infrastructure and Physiography

The Project is serviced by road and rail networks. The mine is accessed from the main paved highway between Erzincan and Kemaliye. The Project area is in the Eastern Anatolia geographical district of Turkey. Mining operations are conducted year-round. The climate is typically continental with cold wet, winters and hot dry, summers.

1.4 History

The Çöpler region has been subject to gold and silver mining dating back at least to Roman times. The Turkish Geological Survey (MTA) carried out regional exploration work in the early- 1960s that was predominately confined to geological mapping. In 1964, a local Turkish company started mining for manganese, continuing through until closing in 1973. Unimangan Manganez San A.Ş. (Unimangan) acquired the property in January 1979 and re-started manganese production, continuing until 1992.

In 1998, Anatolia Minerals Development Ltd (Anatolia) identified several porphyry-style gold–copper prospects in east central Turkey and applied for exploration licences for these prospects. During this work, Anatolia identified a prospect in the Çöpler basin. This prospect and the supporting work were the basis for a joint venture agreement for exploration with Rio Tinto and Anatolia and in January 2004, Anatolia acquired the interests of Rio Tinto and Unimangan.

In August 2009, a joint venture agreement between Anatolia and Lidya was executed.

In February 2011, Anatolia merged with Avoca Resources Limited, an Australian company, to become Alacer Gold Corp. (Alacer). In September 2020, Alacer merged with SSR.

Technical Reports have been prepared on the Project, in accordance with NI 43-101 Standards for Disclosure for Mineral Projects, since 2003. The previous Technical Report on the Project, issued in 2020, described a Reserve Case plus a Preliminary Economic Assessment (PEA) on the Ardich deposit.

1.5 Geological Setting and Mineralisation

The Project is located near the northern margin of a complex collision zone that lies between the Pontide Belt / North Anatolian Fault, the Arabian Plate, and the East Anatolian Fault, which bounds several major plates. The region underwent crustal thickening related to the closure of a single ocean, or possibly several oceanic and micro-continental realms, in the late Cretaceous to early Tertiary period. Figure 1.3 illustrates the broad structural setting of the Anatolia region of Turkey. The Çöpler project area is located between Divriği and Ovacık.

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Figure 1.3 Structural Setting of Anatolia

Anagold, 2020

The gold, silver, and copper mineralisation of economic interest at Çöpler occurs in a porphyry-related epithermal environment, with most of the gold mineralisation concentrated in six distinct areas in the deposit: Main, Main West, Main East, Manganese, Marble, and West. The mineralisation is considered to be related to fluids associated with diorite intrusions at depth and generally manifests as three closely related mineralisation styles across the six areas:

•Low-grade porphyry vein mineralisation.

•Intermediate sulfidation epithermal mineralisation.

•Iron skarn and carbonate replacement mineralisation.

Oxidation of hypogene mineralisation has resulted in the formation of gossans, massive manganese oxide, and goethitic / jarositic assemblages hosting fine-grained free gold. The oxidised cap is underlain by primary and secondary sulfide mineralisation. Çöpler is a geologically complex system due to structural disturbance and multiple-stage diorite intrusions. The initial mineralisation concept model, based on geochemistry of an epithermal system overlying a copper–gold porphyry dome, continues to hold true with current modelling.

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1.6 Exploration

Exploration completed since Anatolia’s involvement in the Çöpler project commenced in 1998 has included:

•Geological and reconnaissance mapping.

•Rock chip, grab, soil, channel, and stream sediment geochemical sampling.

•Ground geophysical surveys including ground magnetic, complex resistivity / induced polarisation (IP), time domain IP, and controlled source audio-frequency magneto-telluric (CSAMT) surveys.

•A regional helicopter-borne geophysical survey.

•Reverse circulation (RC) and diamond core (DD) drilling programmes.

•Acquisition of satellite imagery.

Other related work has included:

•Mining technical studies.

•Geotechnical and hydrogeological studies.

•Environmental and social baseline studies.

•Studies in support of project permitting.

•Metallurgical testwork and metallurgical studies.

•Condemnation evaluations.

The principal exploration technique at the Çöpler project has been RC and DD drilling, conducted in several campaigns starting in 2000. Initially, exploration was directed at evaluating the economic potential of the near-surface oxide mineralisation for the recovery of gold by either heap leaching or conventional milling techniques.

1.7 Drilling

Drilling at the Çöpler deposit commenced in 2000, and since that time a total of 2,635 holes have been drilled for 373,561.9 m. A total of 68 DD holes have been completed in 2021 (18,491.8 m).

•Infill drilling programmes have been conducted since 2007 to improve confidence in the short-term mine planning.

•Drilling in 2014 focused on mineralisation confirmation with a twin-hole programme. The Çöpler deposit continues to be tested using RC and DD drilling as production proceeds.

•Development drilling continued in 2015 by improving sample coverage at depth in the Manganese Zone and along structural boundaries in the Main Zone. In addition to the drilling of in situ mineralisation, a stockpile drilling programme began in December 2015 to confirm sulfide stockpile ore grade, grade distribution, and mineralogy.

•Drilling in 2016–2020 mainly focused on target generation to supplement the amount of oxide material in production. This was focused on the Main Zone, the West pit, and the Saddle areas. These drilling programmes aimed to test continuation of the main gold-bearing structures based on a re-interpretation of the Çöpler structural and mineralisation settings. In-pit drilling campaigns continue.

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•Drilling at the Çöpler deposit in 2020–2021 focused on confirmation of copper mineralisation. A total of 199 DD holes were drilled for 41,521.7 m in this period.

Drilling at Çakmaktepe commenced in 2012 and has resulted in the definition of three distinct mineralised zones: East, Central, and North. As production proceeded within the Çakmaktepe Central and East pits, additional targets were generated to provide pushback options around the pit design. A total of 136 DD holes have been completed since 2019 to test for continuation of the Çakmaktepe mineralisation to the north and the east.

Since the initial discovery of mineralisation at Ardich, Anagold has undertaken several drilling programmes to better-define the geological model and to attempt to increase resource inventories. Anagold has completed 531 DD holes for 111,004.35 m at Ardich from late-2017 to 31 December 2021, including holes for metallurgical testing and hydrogeological studies. Drillholes AR01–AR427 are included in an updated geological model for Ardich, developed in late-2021.

Drilling at Bayramdere commenced in 2007 as part of the near-mine exploration strategy. Since that time 120 holes have been drilled at Bayramdere for a total of 10,734.2 m.

1.8 Sampling Method, Approach and Analyses

From 2004 through late-2012, drillhole samples were prepared at ALS İzmir, Turkey (ALS İzmir) and analysed at ALS Vancouver, Canada (ALS Vancouver), (collectively ALS Global).

From late-2012 through 2014, samples were prepared and analysed at ALS İzmir.

In 2015, samples were prepared and analysed at the SGS laboratory in Ankara, Turkey (SGS).

From 2015 to current, ALS İzmir is being used as the main laboratory and samples are being prepared and analysed there. Umpire analysis was completed by ACME Mineral Laboratories (ACME) in Ankara, Turkey.

ALS İzmir has ISO 9001:2008 certification, and ALS Vancouver is ISO/IEC 17025:2005 accredited for precious and base metal assay methods. SGS is certified to ISO 9001:2008 and OHSAS 18001. ACME is part of the Bureau Veritas (BV) group, globally certified to ISO9001:2008.

ALS Global and SGS are specialist analytical testing service companies; both are independent of Anagold.

Sampling and quality assurance and quality control (QA/QC) programmes have been in place for all RC and DD drilling conducted since the first drill programme. The QA/QC programme is currently still in use, although the insertion rates have been modified over time.

Anagold operates an on-site laboratory at Çöpler for assay of production samples. The on-site laboratory is certified to ISO 17025:2017 but is not independent. It is primarily used in grade control.

1.9 Data Verification

Data verification procedures are well-established at the Project. Routine ongoing checking of all data is undertaken prior to being uploaded to the database. This is followed by campaign-based independent data verification audits at milestone stages throughout data collection programmes.

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For drillhole data, verification includes the checking of Topcon differential global positioning system (DGPS) collar coordinates relative to topographic surveys, checking of down-hole surveys relative to adjacent readings and planned dip and azimuth of the hole, checking logged data entries to ensure they are consistent with log key sheets, cross-checking a subset of assay data with the original laboratory reports, and submission of and review of QA/QC data.

The QA/QC programme has historically consisted of a combination of QA/QC sample types that are designed to monitor different aspects of the sample preparation and assaying process: Blanks are routinely inserted in order to identify the presence of contamination through the sample preparation process; a variety of CRM standards are routinely inserted in order to monitor and measure the accuracy of the assay laboratory results over time; Field duplicates are routinely inserted as a means of monitoring and assessing sample homogeneity and inherent grade variability and to enable the determination of bias and precision between sample pairs; laboratory duplicates are inserted as a means of testing the precision of the laboratory measurements; and inter-laboratory pulp duplicates are submitted to alternative independent laboratory to assess for bias or drift. The rate of submission has been modified over time but is currently 3%–5% for blanks, CRMs, and duplicates, and 5%–10% for field duplicates.

None of the verification programmes have identified material issues with the supporting data.

1.10 Metallurgical Testwork

1.10.1 Oxide Testwork

The heap leaching facilities were commissioned at the Çöpler mine site in late-2010 and have operated continuously since that time. Oxide heap leach operations were continuing at the CDMP21TRS effective date.

Metallurgical testwork on Çakmaktepe oxide material for heap leaching has been undertaken at the on-site metallurgical laboratory, initially under the supervision of Kappes, Cassiday & Associates. The initial testwork in 2015 undertook bottle roll and column leach tests. The results are comparable with the Çöpler oxide ore, with similar behaviour and leach kinetics. Subsequently, Çakmaktepe oxide ore has been heap leached along with Çöpler oxide ore. Oxide column testwork on oxide ore continues at the on-site laboratory.

Metallurgical testwork on Ardich oxide material for heap leaching has been undertaken at McClelland laboratories and supervised by Metallurgium. An initial testwork programme, including bottle roll and column leach, was carried out in 2019. This initial programme identified two distinct domains with respect to gold recovery based on sulfur content: <1% and 1%–2%. The column testwork results indicated that the listwanite, dolomite, and jasperoid lithologies have physical properties amenable to heap leaching. This initial test programme was followed up with further testwork in 2020.

Analysis of the results of the metallurgical testwork and a review of the existing recovery models for use in economic analysis were undertaken in 2020. This was done for the oxide and sulfide processing, including the flotation circuit. The resulting recoveries have been used in the economic analysis for the CDMP21TRS.

Oxide gold recoveries vary by lithology for Çöpler in the range 62.3%–78.4%, at Çakmaktepe the range is 61%–80%. At Ardich, the testwork suggest recoveries will vary in the range 40%–73%. The average oxide recovery in the CDMP21TRS Reserve Case is 61%.

1.10.2 Sulfide Testwork

The sulfide process plant commenced commissioning in Q4’18. The plant consists principally of a pressure oxidation (POX) leach followed by a cyanide leach to recover gold.

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Significant testwork had been conducted on sulfide ores prior to commissioning of the sulfide plant, with pilot plant testwork campaigns and a significant number of batch variability tests on POX / cyanide leach completed.

Whilst a POX / cyanide leach circuit was implemented, significant work had also been undertaken on flotation of the gold-bearing sulfides as a process route, although ultimately this option was not selected for development. Flotation of a partial stream of the plant feed was considered to maximise the available capacity of the plant, including the POX autoclave and available oxygen supply. Further flotation testwork demonstrated that the addition of a small flotation plant into the existing sulfide process route would allow optimisation and maximisation of already installed capacities.

The testwork indicates that sulfur recovery through flotation is estimated to be 75% to concentrate with a corresponding 55% gold recovery. Flotation tails gold recovery is estimated at 43%.

The current determination of POX gold recovery is based on assessment of results for the pilot testwork programmes undertaken prior to commencement of operations and benchmarked with the existing operating data. An equation has been derived to calculate gold recovery by material type for all ore that is subject to POX; this includes direct POX feed and flotation concentrate. The CDMP21TRS Reserve Case average sulfide gold recovery is 91%.

1.11 Mineral Resources

Mineral Resource sections have not been changed from the 2021MR and remains the most current study work available. The CDMP21TRS QPs have reviewed and accepted this information for use in the CDMP21TRS.

Mineral Resources for the Project have been estimated using industry best practices and conform to the requirements of S-K-1300.

The resource model for Ardich has been updated in 2021 and is reported in detail here.

All other resource models are unchanged since the Çöpler District Master Plan 2020 (CDMP20TR), and the reader is directed to that report for the more detail on those resource models, with only summaries included here.

1.11.1 Resource Modelling

1.11.1.1 Çöpler Deposit

The Çöpler deposit includes four mine areas: Main, Manganese, Marble, and West. The current Çöpler resource model, which was constructed by Anagold personnel, was completed in February 2016.

The cut-off date for the drillholes database was 15 July 2015. The data extract contained 1,957 drillholes with a total of 297,798.2 m of drilling. Of this, a total of 1,880 drillholes have collar coordinates within the extents used to construct the resource model. In general, the drillhole spacing ranged from 5–60 m, averaging approximately 20 m. Most drillholes are either vertical or inclined at 60°.

Wireframes were constructed for the four main geological units: diorite, metasediment, marble, and manganese-rich diorite. Drillhole data and surface mapping were developed into 3D solids that represent the major rock types using implicit modelling techniques. This process included generating contact surfaces used to define the division boundaries that represent the geological faults and lithological contacts.

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The resource estimation method at Çöpler was developed to address the variable nature of the gold mineralisation while honouring the bi-modal distribution of the sulfur mineralisation that is critical for mine planning (material with a total sulfur grade <2% is sent to the heap leach while material with total sulfur grade ≥2% is sent to the sulfide stockpile for eventual processing at the POX plant). Since no obvious correlations were observed between gold and total sulfur, gold and sulfur were domained and estimated separately. Gold showed little correlation with lithology and was domained by mining areas to reflect the different trends of the mineralisation that commonly follow structures and/or lithological contacts. Due to the strong correlation between sulfur content and lithology, sulfur was domained by lithology. However, since each lithology may contain <2% S and ≥2% S material, each lithology was additionally separated into <2% S and ≥2% S sub-domains.

Probability assigned constrained kriging (PACK) was used to estimate the gold content of the mineralisation within an expanded mineralised wireframe. A probabilistic envelope was generated within the expanded gold shape to define the limits of the economic mineralisation. The wireframe and probabilistic envelope were used to prevent potentially economic assays from being ‘smeared’ into non-economic zones, and conversely to restrict waste assays from diluting the potentially economic mineralisation. Two PACK cell models were constructed for gold. The first (low-grade gold) model was applied to <2% S material that can be processed by heap leaching, and the second (high-grade gold) model was later applied to ≥2% S material to be processed by the POX plant.

Once constructed, the gold models were calibrated to historical production data, categorised by sulfur content (<2% S and ≥2% S), and mining area. Estimates were classified into Mineral Resource categories based on drillhole density and data quality.

Density values were assigned to the cell model based on lithological domain and depth below the surface.

1.11.1.2 Çakmaktepe Deposit

The Çakmaktepe deposit is located 6 km east of the current Çöpler pit and includes four areas: North, Central, East, and South-east. The current Çakmaktepe resource model, which was constructed by Anagold personnel, was completed in February 2020.

The drillhole dataset used to develop the February 2020 resource model contained a total of 1,109 holes with a drilling date range of September 2007 through October 2019. The total drilled metres input into the modelling was 119,001.1 m. Original sample lengths are predominately 1 m in length with some 2 m sampling across areas presumed to be waste. The mean sample length was 1.02 m. The shortest interval was 0.1 m with maximum length 3.1 m. Composited samples 5 m in length were used for statistical analysis, construction of interpretation boundaries, and grade estimation.

Mineralisation at Çakmaktepe follows structural controls and designated lithological contact orientations. Mineralised zones often incorporate multiple lithological units along the boundary rather than being hosted by a single rock type. For this reason, grade shells were constructed for gold and copper to allow estimation concordant with the mineralised zones instead of being controlled by samples residing within a single lithological unit. Mineralised trends were honoured in 3D with no specific grade cut-off used to bound the mineralised shapes. The resulting shapes for gold and copper are lenticular with thicknesses ranging from 5–40 m. On average, thicknesses are of the order of 6 m.

Sulfur grades correlate with lithological units: higher sulfur values are associated with diorite and metasediment, and lower sulfur values are in association with gossan, jasperoid, ophiolite, and marble.

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A single geological cell model with 5 m x 5 m x 5 m parent cells was constructed to include the four deposit areas. Gold, silver, copper, sulfur, and carbon were estimated using inverse distance interpolation (ID) weighted to the power of three (ID3) and 5 m drillhole composites. Gold, copper, and silver were estimated using grade shells as hard boundaries. Sulfur and carbon estimates were constrained by modelled lithological units. All grade shell boundaries for metal estimates were treated as hard. Domains were treated as soft boundaries allowing the selection of samples from nearby domains.

Density values were assigned to the cell model based on lithological domain.

1.11.1.3 Ardich Deposit

The Ardich deposit is located 1.5 km north of Çakmaktepe and includes two areas: Main and East. The current Ardich resource model, which was constructed by OreWin, was completed in 2021.

The drillhole dataset used to develop the January 2022 resource model contained a total of 427 diamond core drillholes with a drilling date range of September 2017 through May 2021. The total drilled metres for this Ardich dataset was 87,038.25 m. Original sample lengths are predominately 1 m (77.5% of the samples). The shortest assayed interval was 0.2 m, the maximum length 3.8 m, and the mean sample length was 1.19 m. Samples were composited to 1 m length for use in statistical analysis, construction of interpretation boundaries, and grade estimation.

The Ardich Mineral Resource estimate was based on a 3D geological solids model developed within constraining fault blocks. High-angle faults crosscut the deposit creating rotated structural blocks that have moved up and down relative to each other. There are 25 distinct fault block domains in the 2021 model.

The main lithological units: ophiolite, listwanite, dolomite, jasperoid, and cataclasite, are disrupted by the faults. Owing to the offsets at fault boundaries and the variable thicknesses of lithologies from one fault block to the next, the lithological interpretations have been completed separately for each fault block.

Discrete domains for grade estimation are defined by the fault block and lithology interpretations. As the amount of drill data increased, the understanding of the structural and lithological domains has developed such that a total of 221 unique fault block / lithology domains exist in the 2021 model.

Gold distribution is related to the lithological contact zones and structural intersections. These zones tend to be narrow and localised. Mineralised trends generally follow the orientations of the structural features, further nuanced at the lithological contacts as they change within each of the fault blocks. Control of the gold estimation in the model is accomplished with the use of the fault block and lithology domains as hard boundaries to (a) limit the samples informing estimation in each lithological unit to only those of the same fault block, and (b) to orientate local search neighbourhoods within each domain (dynamic anisotropy). Unlike previous Ardich models, grade shells were not used to constrain estimation in the 2021 model.

A cell model with 10 m x 10 m x 5 m parent cells was constructed to cover the entire Ardich deposit. Sub-celling to 2.5 m x 2.5 m x 1.25 m was permitted to honour interpreted boundaries. Further sub-celling to a minimum of 0.25 m was permitted at the topographic surface. Estimation of a suite of 13 grades (including Au with and without top cuts) and density was undertaken using ordinary kriging. A nearest neighbour estimation of Au was completed for validation purposes.

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1.11.1.4 Bayramdere Deposit

The Bayramdere deposit is located approximately 6.3 km east of the Çöpler mine and 5 km south-east of Iliç. It is within the Kartaltepe Mining Licence 7083. This licence is an operational licence and is 50% SSR-owned.

The Bayramdere mineralisation has an overall strike length of approximately 300 m. Mineralisation is localised within three stacked, shallow-dipping lodes that vary in depth between 30–40 m below topography. Mineralisation appears to be open to the east and south.

A resource model for Bayramdere was completed in 2016. Separate mineralisation domains were created for gold, silver, copper, and sulfur. In the creation of mineralised domains, a minimum mining width of 2.5 m was used based on anticipated open pit mining methods. Grade estimation was limited to the interpreted domains. Outside the mineralised domains a ‘mineralised waste’ estimate was completed. Lithological domains were used for estimates outside of the mineralisation domains. Ordinary kriging was used to estimate gold, silver, and copper mineralisation into parent cells of 10 m x 10 m x 5 m size with sub-celling permitted to 2 m x 2 m x 1 m to better honour the domain boundaries.

Density was assigned as a default for each of the mineralisation and lithological domains.

Although a small deposit, Bayramdere is relatively high-grade and can support a high-stripping ratio to access mineralisation.

1.11.2 Reasonable Prospects for Eventual Economic Extraction

Mineral Resources in the Çöpler District Mineral Resource 2021 Technical Report Summary (2021MR) were assessed for reasonable prospects for eventual economic extraction by one of two ways. For existing operations, by reporting only material that fell within conceptual pit shells based on metal prices of $1,750/oz for gold, or as otherwise specified, and an Initial Assessment has been prepared to demonstrate the further economic potential of the Mineral Resources at the Çöpler deposit with the inclusion of a copper concentrator. The Initial Assessment is preliminary in nature, it includes Inferred Mineral Resources that are considered too speculative geologically to have modifying factors applied to them that would enable them to be categorised as Mineral Reserves, and there is no certainty that this economic assessment will be realised. The Initial Assessment from the 2021MR is described in Section 11.5.2 of the CDMP21TRS.

1.11.3 2021MR Initial Assessment

Mineral Resources were originally discussed and disclosed in the 2021MR. The 2021MR is an independent Technical Report Summary prepared to provide a preliminary technical and economic study of the economic potential of the Çöpler District mineralisation to support the disclosure of Mineral Resources. The 2021MR includes an Initial Assessment that has two cases:

•Initial Assessment MII Case - the base case that includes Measured, Indicated and Inferred Mineral Resources in the analysis and

•Initial Assessment MI Case - this case includes only Measured and Indicated Mineral Resources

The Çöpler project currently has two processing methods:

•Sulfide process plant

•Heap leach oxide processing facility

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The sulfide plant includes the crushing, grinding, flotation and pressure oxidation to produce gold and small amounts of silver. The heap leach facility produces gold and small amounts of silver and copper.

The scenario for the 2021MR Initial Assessment analysis includes additional processing options to recover copper from the sulfide Mineral Resource. The two processing options are:

•Copper concentrator producing a copper concentrate and a pyrite concentrate, and

•Sodium hydrosulfide (NaSH) copper recovery circuit to be installed in the current sulfide plant.

The copper concentrator would make a copper concentrate for sale to smelters and a pyrite concentrate to be fed into the autoclaves in the sulfide plant. The pyrite concentrate would have a high gold content and also provide sulfur as a source of fuel for the autoclaves. The copper concentrator capacity is 1.8 Mtpa.

The Çöpler Mineral Resource has been selected for the 2021MR Initial Assessment analyses, as the other Mineral Resources at the Project do not contain significant amounts of copper.

The 2021MR Initial Assessment has been prepared to demonstrate economic potential of the Mineral Resources at the Çöpler deposit. The Initial Assessment is preliminary in nature, it includes Inferred Mineral Resources that are considered too speculative geologically to have modifying factors applied to them that would enable them to be categorised as Mineral Reserves, and there is no certainty that this economic assessment will be realised.

The estimates of cash flows have been prepared on a real basis with a base date of Q4’21 and a mid-year discounting is used to calculate NPV.

The key production and economic analysis from the CDMP21TRS Reserve Case and the 2021MR Initial Assessment Cases are shown in Table 1.3. The after-tax net present value (NPV) sensitivity to metal price variation is shown in Table 1.4 for gold prices from $1,000–$2,000/oz.

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Table 1.3 CDMP21TRS Results Summary

Item	Unit	CDMP21TRS Reserve Case	Initial Assessment MI Case	Initial Assessment MII Case
Oxide Processed
Heap Leach Quantity	kt	22,557	39,874	41,792
Au Feed Grade	g/t	1.69	1.28	1.26
Sulfide Processed
Quantity Milled	kt	52,892	61,158	59,654
Au Feed Grade	g/t	2.33	2.32	2.45
Cu Concentrator Processed
Quantity Milled	kt	–	9,987	24,939
Au Feed Grade	g/t	–	0.48	0.50
Cu Feed Grade	%	–	0.21	0.20
Total Gold Produced
Oxide – Gold	koz	765	1,026	1,068
Sulfide – Gold	koz	3,604	4,135	4,078
Cu Concentrator – Gold	koz	–	85	222
Total – Gold	koz	4,369	5,247	5,368
Total Copper Production	Mlb	0.02	148	164
5-Year Annual Average
Average Gold Produced	kozpa	278	297	300
Free Cash Flow	$Mpa	158	165	165
Total Cash Costs (CC)	$/oz gold	880	753	761
All-in Sustaining Costs (AISC)	$/oz gold	1,071	932	938
Key Financial Results
Life-of-Mine (LOM) CC	$/oz gold	803	775	783
LOM AISC	$/oz gold	966	921	924
Site Operating Costs	$/t treated	45.91	44.13	43.79
After-Tax NPV5%	$M	1,732	1,867	2,004
Mine Life	years	21	22	22

5-Year annual average is for the period 1 January 2022 through 31 December 2026

Table 1.4 Gold Price Sensitivity

After-Tax NPV5% ($M)	Long-Term Gold Price(/oz)
Case	1,000	1,350	1,600	1,750	2,000
CDMP21TRS Reserve Case	769	1,370	1,732	1,939	2.252
Initial Assessment MI Case	771	1,459	1,867	2,111	2,474
Initial Assessment MII Case	859	1.579	2,004	2,259	2,642

All values are in US Dollars.

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1.11.4 Mineral Resources Estimates

Mineral Resources are reported exclusive of Mineral Reserves and have been summarised by project, Mineral Resource classification, and oxidation state in Table 1.5. Mineral Resources are presented showing only the SSR attributable proportion.

Table 1.6 shows the cut-off values, metallurgical recoveries, and SSR ownership percentage associated with the Mineral Resources.

Mineral Resources have been classified in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300) and were estimated by Sharron Sylvester BSc (Geology), RPGeo AIG (10125), employed by OreWin Pty Ltd as Technical Director – Geology. Mineral Resources are presented on a project basis and have an effective date of 31 December 2021.

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Table 1.5 Summary of CDMP21TRS Mineral Resources Estimates Exclusive of Mineral Reserves as at 31 December 2021 Based on $1,750/oz Gold Price

Mineral Resource Classification		SSR Tonnage<br>(kt)		Grades						Contained Metal
	Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)		Gold<br>(koz)		Silver<br>(koz)			Copper<br>(klb)
Çöpler Mine – Oxide
Measured		65		1.39		4.67		0.16		3	10		225
Indicated		21,739		0.84		2.30		0.16		589	1,610		77,646
Measured + Indicated		21,803		0.84		2.31		0.16		592	1,619		77,871
Inferred		28,017		0.90		6.87		0.13		813	6,192		78,353
Çöpler Mine – Sulfide
Measured		121		0.83		3.72		0.18		3	15		472
Indicated		37,667		1.06		3.66		0.19		1,286	4,428		158,692
Measured + Indicated		37,788		1.06		3.66		0.19		1,289	4,442		159,164
Inferred		39,838		1.24		13.60		0.17		1,585	17,418		145,512
Çakmaktepe – Oxide
Measured		–		–		–		–		–	–		–
Indicated		1,671		1.55		8.33		–		83	447		–
Measured + Indicated		1,671		1.55		8.33		–		83	447		–
Inferred		602		0.85		4.04		–		16	78		–
Ardich – Oxide
Measured		2,101		1.46		3.06		0.01		99	207		350
Indicated		6,786		0.99		2.45		0.01		215	534		1,147
Measured + Indicated		8,887		1.10		2.59		0.01		314	741		1,497
Inferred		8,980		1.29		3.17		0.02		372	914		3,306
Ardich – Sulfide (Incl. sulfide and sulfide-with-Cu)
Measured		180		5.94		8.41		0.01		34	49		34
Indicated		953		2.05		3.61		0.02		63	111		383
Measured + Indicated		1,133		2.67		4.37		0.02		97	159		416
Inferred		2,209		2.75		4.47		0.01		195	317		367
Bayramdere – Oxide
Measured		–		–		–		–		–	–		–
Indicated		72		2.34		20.82		–		5	48		–
Measured + Indicated		72		2.34		20.82		–		5	48		–
Inferred		4		2.17		19.95		–		0	3		–
CDMP21 Mineral Resources – Oxide Subtotal
Measured		2,166		1.46		3.11		0.01		102	217		575
Indicated		30,267		0.92		2.71		0.12		894	2,640		78,793
Measured + Indicated		32,433		0.95		2.74		0.11		995	2,857		79,368
Inferred		37,603		0.99		5.95		0.10		1,202	7,188		81,659
CDMP21 Mineral Resources – Sulfide Subtotal
Measured		301		3.88		6.52		0.08		38	63		506
Indicated		38,620		1.09		3.66		0.19		1,349	4,538		159,075
Measured + Indicated		38,921		1.11		3.68		0.19		1,387	4,602		159,580
Inferred		42,047		1.32		13.12		0.16		1,780	17,735		145,879
CDMP21 MINERAL RESOURCES – OVERALL TOTAL (Exclusive of Mineral Reserves)
Measured		2,467		1.76		3.53		0.02		139	280		1,081
Indicated		68,887		1.01		3.24		0.16		2,243	7,178		237,867
Measured + Indicated		71,354		1.04		3.25		0.15		2,382	7,458		238,948
Inferred		79,650		1.16		9.73		0.13		2,982	24,923		227,538

1. Mineral Resources are reported based on 31 December 2021 topography surface.

2. Mineral Resources are reported exclusive of Mineral Reserves.

3. Mineral Resources are reported showing only the SSR attributable tonnage portion. Çöpler Mineral Resources are located on ground held 80% by SSR, Çakmaktepe and Bayramdere Mineral Resources are located on ground held 50% by SSR, and approximately 96% of Ardich Mineral Resources are located on ground held 80% by SSR, with the remainder located on ground 50% held by SSR.

4. Oxide definitions: At Çöpler: oxide is defined as material <2% total sulfur and sulfide material is ≥2% total sulfur. At Ardich and Çakmaktepe, oxide is comprised of low-sulfur (LS) oxide (<1% total sulfur) and high-sulfur oxide (≥1% and <2% total sulfur). At Bayramdere: oxide is defined as material <2% total sulfur.

5. Sulfide definitions: At Ardich, sulfide is comprised of standard sulfide material (≥2% total sulfur) and sulfide-with-Cu material (sulfide with Cu>0.10%). There is no sulfide material at Çakmaktepe or Bayramdere.

6. At Çöpler and Ardich: sulfide cut-off uses an NSR value in $/t based on gold price $1,750/oz, silver price $22.00/oz Ag and copper price $3.95/lb with allowances for payability, deductions, transport, and royalties.

7. The point of reference for Mineral Resources is the point of feed into the processing facility.

8. All Mineral Resources in the CDMP21TRS were assessed for reasonable prospects for eventual economic extraction by reporting only material that fell within conceptual pit shells ($1,400/oz for gold and $19/oz for silver for Bayramdere, and $1,750/oz for gold, $22/oz for silver for all other projects).

9. Totals may vary due to rounding.

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Table 1.6 Summary of Cut-off Values, Metallurgical Recoveries, and SSR Ownership of CDMP21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on Gold Price $1,750/oz, Silver Price $22.00/oz Ag and Copper Price $3.95/lb

Mineral Resource Classification		SSR Tonnage<br>(kt)		Grades				Cut-off Value/s	Metallurgical Recovery (%)	SSR Ownership (%)
	Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)
Çöpler Mine – Oxide
Measured		65		1.39		4.67	0.16	0.19–0.24 g/t Au	62.3–78.4	80
Indicated		21,739		0.84		2.30	0.16
Measured + Indicated		21,803		0.84		2.31	0.16
Inferred		28,017		0.90		6.87	0.13
Çöpler Mine – Sulfide
Measured		121		0.83		3.72	0.18	34.88/t NSR or>0.10% Cu and 7.68/t NSR	Au 55–91<br><br>Ag 10–45<br><br>Cu 84–98	80
Indicated		37,667		1.06		3.66	0.19
Measured + Indicated		37,788		1.06		3.66	0.19
Inferred		39,838		1.24		13.60	0.17
Çakmaktepe – Oxide
Measured		–		–		–	–	0.36–0.76 g/t Au	38.0–80.0	50
Indicated		1,671		1.55		8.33	–
Measured + Indicated		1,671		1.55		8.33	–
Inferred		602		0.85		4.04	–
Ardich – Oxide
Measured		2,101		1.46		3.06	0.01	0.23–0.41 g/t Au	40.0–73.0	75
Indicated		6,786		0.99		2.45	0.01
Measured + Indicated		8,887		1.10		2.59	0.01
Inferred		8,980		1.29		3.17	0.02
Ardich – Sulfide (Incl. sulfide and sulfide-with-Cu)
Measured		180		5.94		8.41	0.01	36.25/t NSR or>0.10% Cu and 9.05/t NSR	Au 55–91<br><br>Ag 10–45<br><br>Cu 84–98	78
Indicated		953		2.05		3.61	0.02
Measured + Indicated		1,133		2.67		4.37	0.02
Inferred		2,209		2.75		4.47	0.01
Bayramdere – Oxide
Measured		–		–		–	–	0.35–0.50 g/t Au	75	50
Indicated		72		2.34		20.82	–
Measured + Indicated		72		2.34		20.82	–
Inferred		4		2.17		19.95	–
CDMP21 Mineral Resources – Oxide Subtotal
Measured		2,166		1.46		3.11	0.01	As Above	As Above	75
Indicated		30,267		0.92		2.71	0.12
Measured + Indicated		32,433		0.95		2.74	0.11
Inferred		37,603		0.99		5.95	0.10
CDMP21 Mineral Resources – Sulfide Subtotal
Measured		301		3.88		6.52	0.08	As Above	As Above	78
Indicated		38,620		1.09		3.66	0.19
Measured + Indicated		38,921		1.11		3.68	0.19
Inferred		42,047		1.32		13.12	0.16
CDMP21 MINERAL RESOURCES – OVERALL TOTAL (Exclusive of Mineral Reserves)
Measured		2,467		1.76		3.53	0.02	As Above	As Above	76
Indicated		68,887		1.01		3.24	0.16
Measured + Indicated		71,354		1.04		3.25	0.15
Inferred		79,650		1.16		9.73	0.13

All values are in US Dollars.

1. Mineral Resources are reported based on 31 December 2021 topography surface.

2. Mineral Resources are reported exclusive of Mineral Reserves.

3. Mineral Resources are reported showing only the SSR attributable tonnage portion. SSR Ownership is an average based on location of Mineral Resources (gold) relative to licenses: Çöpler and part of Ardich are on Anagold 80:20 ground on which SSR holds 80% rights, and Çakmaktepe, Bayramdere and the remainder of Ardich are on Kartaltepe 50:50 ground on which SSR holds 50% rights.

4. Totals and Ardich ownership percentages are weighted averages.

5. Oxide definitions: At Çöpler: oxide is defined as material <2% total sulfur and sulfide material is ≥2% total sulfur. At Ardich and Çakmaktepe, oxide is comprised of low-sulfur (LS) oxide (<1% total sulfur) and high-sulfur oxide (≥1% and <2% total sulfur). At Bayramdere: oxide is defined as material <2% total sulfur.

6. Sulfide definitions: At Ardich, sulfide is comprised of standard sulfide material (≥2% total sulfur) and sulfide-with-Cu material (sulfide with Cu>0.10%). There is no sulfide material at Çakmaktepe or Bayramdere.

7. At Çöpler and Ardich: sulfide cut-off uses an NSR value in $/t based on gold price $1,750/oz, silver price $22.00/oz, and copper price $3.95/lb with allowances for payability, deductions, transport, and royalties.

8. The point of reference for Mineral Resources is the point of feed into the processing facility.

9. All Mineral Resources in the CDMP21TRS were assessed for reasonable prospects for eventual economic extraction by reporting only material that fell within conceptual pit shells ($1,400/oz for gold and $19/oz for silver for Bayramdere, and $1,750/oz for gold, $22/oz for silver for all other projects).

10. Totals may vary due to rounding.

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1.12 Mineral Reserves Estimates

Mineral Reserves have been summarised by project, Mineral Reserve classification, and oxidation state in Table 1.7 and in Table 1.8.

Mineral Reserves have been classified in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300) and were estimated by Bernard Peters BEng (Mining), FAusIMM (201743), employed by OreWin Pty Ltd as Technical Director - Mining. Mineral Reserves are presented on a SSR attributable tonnage portion basis and have an effective date of 31 December 2021.

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Table 1.7 Summary of CDMP21TRS Mineral Reserves Estimates (as at 31 December 2021) Based on $1,350/oz Gold Price

Mineral Reserve Classification		SSR Tonnage<br>(kt)		Grades						Contained Metal
	Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)		Gold<br>(koz)		Silver<br>(koz)		Copper<br>(klb)
Çöpler Mine – Oxide
Proven Mineral Reserve		–		–		–		–		–		–	–
Probable Mineral Reserve		1,763		1.22		11.17		0.13		69		633	5,043
Probable – Stockpile		–		–		–		–		–		–	–
Total Mineral Reserve		1,763		1.22		11.17		0.13		69		633	5,043
Çöpler Mine – Sulfide
Proven Mineral Reserve		326		2.02		6.69		–		21		70	–
Probable Mineral Reserve		28,662		2.13		4.96		–		1,964		4,571	–
Probable – Stockpile		9,974		2.25		–		–		720		0	–
Total Mineral Reserve		38,962		2.16		3.70		–		2,705		4,641	–
Çakmaktepe Mine – Oxide
Proven Mineral Reserve		–		–		–		–		–		–	–
Probable Mineral Reserve		137		1.26		10.91		–		6		48	–
Probable – Stockpile		6		2.69		–		–		1		–	–
Total Mineral Reserve		143		1.32		10.49		–		6		48	–
Ardich – Oxide Reserve
Proven Mineral Reserve		5,279		1.73		2.00		0.00		293		339	158
Probable Mineral Reserve		10,287		1.73		1.91		0.01		572		631	2,231
Probable – Stockpile		–		–		–		–		–		–	–
Total Mineral Reserve		15,566		1.73		1.94		0.01		866		970	2,389
Ardich – Sulfide
Proven Mineral Reserve		1,428		5.67		11.08		–		260		509	–
Probable Mineral Reserve		1,632		3.10		4.15		–		163		218	–
Probable – Stockpile		–		–		–		–		–		–	–
Total Mineral Reserve		3,060		4.30		7.38		–		423		726	–
CDMP21 Mineral Reserves – Oxide Subtotal
Proven Mineral Reserve		5,279		1.73		2.00		0.00		293		339	158
Probable Mineral Reserve		12,187		1.65		3.35		0.03		647		1,312	7,274
Probable – Stockpile		6		2.83		–		–		1		0	0
Total Mineral Reserve		17,472		1.68		2.94		0.02		941		1,651	7,432
CDMP21 Mineral Reserves – Sulfide Subtotal
Proven Mineral Reserve		1,754		4.99		10.26		–		281		579	–
Probable Mineral Reserve		30,294		2.18		4.92		–		2,127		4,788	–
Probable – Stockpile		9,974		2.25		–		–		720		0	–
Total Mineral Reserve		42,022		2.32		3.97		–		3,128		5,367	–
CDMP21 MINERAL RESERVES – OVERALL TOTAL
Proven Mineral Reserve		7,033		2.54		4.06		0.00		574		918	158
Probable Mineral Reserve		42,481		2.03		4.47		0.01		2,774		6,100	7,274
Probable – Stockpile		9,980		2.25		–		–		721		0	0
Total Mineral Reserve		59,494		2.13		3.67		0.01		4,069		7,018	7,432

1. Mineral Reserves are reported based on 31 December 2021 topography surface.

2. The Mineral Reserves were scheduled based on end of August 2021topography surface. Small differences between the Mineral Reserve statement and the production schedule may occur.

3. Mineral Reserves are reported showing only the SSR attributable tonnage portion. Çöpler and part of Ardich are on Anagold 80:20 ground on which SSR holds 80% rights, and Çakmaktepe and the remainder of Ardich are on Kartaltepe 50:50 ground on which SSR holds 50% rights.

4. Mineral Reserve cut-offs are based on $1,350/oz gold price; average oxide recoveries are 61% and average sulfide recoveries are 91%.

5. The point of reference for Mineral Reserves is the point of feed into the processing facility.

6. Cut-off values are shown in Table 1.8. All cut-off values include allowance for royalty payable. There are no credits for silver or copper in the cut-off calculations.

7. There is no Çakmaktepe sulfide Mineral Reserve or Bayramdere Mineral Reserve.

8. Economic analysis has been carried out using a long-term gold price of $1,600/oz.

9. Totals may vary due to rounding.

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Table 1.8 Summary of Cut-off Values, Metallurgical Recoveries, and SSR Ownership of CDMP21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price

Mineral Reserve Classification		SSR Tonnage<br>(kt)		Grades				Cut-off Value/s<br><br>(g/t Au)		Metallurgical Recovery<br><br>(%)	SSR Ownership<br><br>(%)
	Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)
Çöpler Mine – Oxide
Proven Mineral Reserve		–		–		–	–	–		–	–
Probable Mineral Reserve		1,763		1.22		11.17	0.13	0.47–0.59		62.3–78.4	80
Probable – Stockpile		–		–		–	–	–		–	–
Total Mineral Reserve		1,763		1.22		11.17	0.13	0.47–0.59		62.3–78.4	80
Çöpler Mine – Sulfide
Proven Mineral Reserve		326		2.02		6.69	–	1.05		85	80
Probable Mineral Reserve		28,662		2.13		4.96	–
Probable – Stockpile		9,974		2.25		–	–
Total Mineral Reserve		38,962		2.16		3.70	–
Çakmaktepe Mine – Oxide
Proven Mineral Reserve		–		–		–	–	–		–	–
Probable Mineral Reserve		137		1.26		10.91	–	0.52–0.71		14–80	50
Probable – Stockpile		6		2.69		–	–
Total Mineral Reserve		143		1.32		10.49	–
Ardich – Oxide
Proven Mineral Reserve		5,279		1.73		2.00	0.00	0.44–0.80		40–73	77
Probable Mineral Reserve		10,287		1.73		1.91	0.01
Probable – Stockpile		–		–		–	–	–		–	–
Total Mineral Reserve		15,566		1.73		1.94	0.01	0.44–0.80		40–73	77
Ardich – Sulfide
Proven Mineral Reserve		1,428		5.67		11.08	–	1.11		83	78
Probable Mineral Reserve		1,632		3.10		4.15	–		72
Probable – Stockpile		–		–		–	–	–		–	–
Total Mineral Reserve		3,060		4.30		7.38	–	1.11		83	75
CDMP21 Mineral Reserves – Oxide Subtotal
Proven Mineral Reserve		5,279		1.73		2.00	0.00	0.44–0.80		14–80	77
Probable Mineral Reserve		12,187		1.65		3.35	0.03		77
Probable – Stockpile		6		2.83		–	–	0.52–0.71		14–80	50
Total Mineral Reserve		17,472		1.68		2.94	0.02	0.44–0.80		14–80	77
CDMP21 Mineral Reserves – Sulfide Subtotal
Proven Mineral Reserve		1,754		4.99		10.26	–	1.05–1.11		83–85	78
Probable Mineral Reserve		30,294		2.18		4.92	–		79
Probable – Stockpile		9,974		2.25		–	–		80
Total Mineral Reserve		42,022		2.32		3.97	–		79
CDMP21 MINERAL RESERVES – OVERALL TOTAL
Proven Mineral Reserve		7,033		2.54		4.06	0.00	0.44–1.11		14–85	77
Probable Mineral Reserve		42,481		2.03		4.47	0.01		79
Probable – Stockpile		9,980		2.25		–	–		80
Total Mineral Reserve		59,494		2.13		3.67	0.01		78

1. Mineral Reserves are reported based on 31 December 2021 topography surface.

2. The Mineral Reserves were scheduled based on end of August 2021topography surface. Small differences between the Mineral Reserve statement and the production schedule may occur.

3. Mineral Reserves are reported showing only the SSR attributable tonnage portion. SSR Ownership is an average based on location of Mineral Reserves (gold) relative to licenses: Çöpler and part of Ardich are on Anagold 80:20 ground on which SSR holds 80% rights, and Çakmaktepe and the remainder of Ardich are on Kartaltepe 50:50 ground on which SSR holds 50% rights. Totals and Ardich ownership percentages are weighted averages.

4. Mineral Reserve cut-offs are based on $1,350/oz gold price; average oxide recoveries are 61% and average sulfide recoveries are 91%.

5. The point of reference for Mineral Reserves is the point of feed into the processing facility.

6. All cut-off grades include allowance for royalty payable. There are no credits for silver or copper in the cut-off grade calculations.

7. There is no Çakmaktepe sulfide Mineral Reserve or Bayramdere Mineral Reserve.

8. Economic analysis has been carried out using a long-term gold price of $1,600/oz.

9. Totals may vary due to rounding.

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1.13 Mining Method

Open pit mining at the Çöpler project is carried out by a mining contractor and managed by Anagold. The mining method is a conventional open pit method with drill and blast and utilising excavators and trucks operating on bench heights of 5 m. The mining contractor provides operators, line supervisors, equipment, and ancillary facilities required for the mining operation. Anagold provides management, technical, mine planning, engineering, and grade control functions for the operation.

Anagold currently operates a sulfide process plant and an oxide heap leach facility. Costs are based on the actual operational costs and the Project budget assumptions.

Production schedules and costs have been updated based on current site performance and contracts.

Pit designs for the Çöpler pit were reviewed and updated in 2021. The Ardich pit designs were prepared in 2021 and 2022. The pit designs included in the CDMP21TRS Reserve Case are shown in Figure 1.4. The CDMP21TRS Reserve Case mining production on a 100% project basis is shown in Figure 1.5.

Figure 1.4 CDMP21TRS Ultimate Pit Designs

Anagold, 2022

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Figure 1.5 CDMP21TRS Reserve Case Mining Production

OreWin, 2022

1.14 Recovery Methods

1.14.1 Sulfide Plant

The sulfide plant commenced commissioning in Q4’18. The basic flow sheet is shown in Figure 1.6 and comprises:

•Crushing and ore handling

•Grinding

•Acidulation

•Pressure oxidation

•Iron / arsenic precipitation

•Counter current decantation (CCD)

•Gold leach, carbon adsorption, and detoxification

•Carbon desorption and refining

•Neutralisation and tailings

•Tailing storage facility (TSF)

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The sulfide plant performance from Q4’18 up to Q4’21, including commissioning and ramp-up, has achieved greater-than-design throughputs and approaches design gold recovery for the ore types processed.

Figure 1.6 Process Flow Sheet for Sulfide Plant

Anagold, 2020

The incorporation of a new flotation circuit in the existing sulfide plant to upgrade sulfide sulfur (SS) to fully utilise POX autoclave oxidation capacity is complete and commissioning commenced in January 2021. This addition to the sulfide plant is incorporated between grinding and acidulation, as shown in Figure 1.7, by taking a bleed / slip stream from the grinding thickener feed, floating gold-bearing sulfides, rejecting acid-consuming carbonates and returning the sulfide concentrate to the grinding thickener to be combined with direct POX feed. The gold not recovered to concentrate that remains in the flotation tails is directed to the gold leach circuit feed to recover this remaining gold, albeit at lower gold recoveries than ore that is treated through the POX autoclave circuit.

This will increase overall plant maximum throughput rate to 400 tonnes per hour (tph), allowing the grinding and POX circuit to operate at their maximum demonstrated capacities. The grinding circuit maximum volumetric flow throughput will increase from an original design limit of 306 tph to 400 tph, fully utilising latent capacity within the crushing and grinding circuit. The flotation plant is designed to operate in the throughput range of 50–150 tph to produce a concentrate that will supplement the feed ore SS to maximise autoclave SS up to 13.75 tph at a maximum autoclave feed rate of 280 tph.

Operating performance of the autoclaves indicates that higher than design oxygen utilisations efficiencies are possible, which may allow greater than 13.75 tph SS to be treated. This oxygen utilisation efficiency along with increased oxygen availability is upside to the CDMP21TRS Reserve Case.

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Figure 1.7 Flotation Block Flow Diagram

Anagold, 2020

1.14.2 Oxide Ore Heap Leach Processing

In the CDMP21TRS Reserve Case production is predominantly from sulfide ore. The maximum oxide ore placed in any year is 4.0 Mt for a total production of 22.5 Mt.

The oxide heap leaching and associated facilities were commissioned in the second half of 2010 and initial gold production was achieved in Q4’10. The process was originally designed to treat approximately 6.0 Mtpa of ore by three-stage crushing (primary, secondary, and tertiary) to 80% passing 12.5 mm, agglomeration, and heap leaching on a lined heap leach pad with dilute alkaline sodium cyanide solution. Gold is recovered through a carbon in column (CIC) system, followed by stripping of metal values from carbon, electrowinning and smelting to yield a doré (containing gold and silver) suitable for sale. Control of copper in leach solutions is undertaken in a sulfidisation, acidification, recovery, and thickening (SART) plant, which also regenerates cyanide. The oxide ore heap leach process flow sheet is shown in Figure 1.8.

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Figure 1.8 Process Flow Sheet for Oxide Ore Heap Leach

Anagold, 2021

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1.14.3 Project Infrastructure

1.14.3.1 Infrastructure

The facility infrastructure supports the mine and process areas of oxide heap leach and sulfide plant. The existing infrastructure, and the tailings storage and heap leach pad area once the planned expansion is complete, will be sufficient for the current Mineral Reserves. The infrastructure for the addition of flotation to the sulfide plant will be supported by the existing facility infrastructure with some components modified to meet the addition of the flotation circuit.

The current leach pad consists of four phases designed to accommodate approximately 58 Mt of oxide ore heap with a nominal maximum heap height of 100 m above the pad liner. An additional two phases (phase 5 and phase 6), with a capacity of 20 Mt will be added to accommodate the oxide to be mined from Ardich.

The tailings storage facility (TSF) is developed and constructed in stages. The development of TSF 1 includes seven phases. TSF 1 phase 3 construction has been completed and approval for use was granted in October 2021 by the Ministry of Environment, Urbanisation and Climate Change (MoEUCC). Ongoing work in ensuring sufficient long-term capacity for storage of tailings has been undertaken. Studies by Anagold have determined that the effect of the addition of the flotation circuit to the sulfide plant would result in an increase in the solids content and improvement in the final settled density based on an increase in the rate of tailings consolidation.

TSF 1 has sufficient storage capacity (70.8 Mt) to accommodate the CDMP21TRS tailings. Scoping level investigations have identified additional TSF sites. An adjacent site, TSF 2, has been the subject of a PFS level study and can provide approximately 20 Mt of net additional tails storage capacity, if required in the future. A detailed design of TSF 2 has been substantially progressed. In November 2021 an application project package was submitted to the MoEUCC. Project design and approval finalisation is expected by Q3’22.

1.15 Market Studies

The markets for gold and silver doré are readily accessed and available to gold producers. Currently, 100% of the gold and silver is delivered to the Istanbul Gold Refinery. Copper precipitate is currently produced from the SART plant and sold into local markets in Turkey. The sulfide plant does not currently include a copper circuit and the analysis of copper recovery at Çöpler has been considered in the Initial Assessment as part of the 2021MR.

1.16 Environmental and Permitting

The Çöpler mining and processing operations involve open pit mining from multiple pits, construction of multiple waste rock dumps (WRD) to accommodate mined materials, processing of oxide ores and placement on a heap leach pad, and processing of sulfide ores with placement of tailings in a TSF. These activities and facilities are carried out on treasury, pasture, and forestry lands.

In addition to the direct impacts on the involved lands, the operations impact the surrounding lands and the local communities. Physical impacts may include changes to local surface and groundwater (including potential pollution), air quality impacts particularly from dust, and increased noise and vibration from mining and processing activities.

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Operation of the Çöpler mining and processing facilities, and subsequent mining at Çakmaktepe, have been investigated and authorised by means of a series of Environmental Impact Assessments (EIAs), with positive decisions obtained from the Turkish Ministry of Environment, Urbanisation and Climate Change. These EIAs include specific actions designed to address all material impacts of the mining and processing operations. Anagold has remained in compliance with all aspects of the EIA and operating permits throughout the history of the Project.

The original 2008 EIA obtained on 16 April 2008 included three main open pits (Manganese, Marble, and Main), five WRDs, a heap leach pad, a processing plant, and a TSF. The 2008 project description involved only the oxide resources.

The Çöpler mine started its open pit and heap leach operation in 2010 and first gold was poured in December 2010. Additional EIA investigations have been submitted and approved, as required, to support on-going mining and processing operations, including:

•EIA to allow operation of a mobile crushing plant, approved 10 April 2012.

•EIA to allow waste dump capacity expansion, oxide capacity expansion to 23,500 tpd and a SART plant, approved 17 May 2012.

•EIA to allow the sulfide plant and heap leach area expansion, approved 24 December 2014.

•EIA to allow the Çakmaktepe satellite pits expansion, approved 26 January 2017.

•EIA to allow a Çakmaktepe capacity increase, approved 9 August 2018.

•EIA to allow a second capacity expansion at Çöpler, including heap leach pads 5 and 6, TSF expansion, and operation of a flotation plant, approved 7 October 2021.

In addition, pending EIA processes include:

•EIA to allow second capacity increase on the Çakmaktepe EIA to include initial mining from Ardich in the EIA project description file. The EIA project description file was submitted in October 2020 and a Public Hearing was held in November 2020. All public institutions gave positive feedback regarding the report and the approval process is ongoing with the MoEUCC.

Following the EIA positive decisions, additional licences and permits were required to be issued by government agencies consistent with the Turkish governing laws and regulations. These include land access permits (pasture and forestry), operational environmental licences and permits, and workplace opening and operating permits, licences, and certificates.

1.17 Capital and Operating Costs

Capital and operating cost estimates have been developed based on the current project costs, the mine and process designs, and discussions with potential suppliers and contractors. The sulfide growth costs include the capital cost for the flotation circuit. The estimated capital costs are to a feasibility level of accuracy and include a contingency of 10%.

1.17.1 Capital Costs

Capital costs have been split into growth and sustaining costs. The sustaining costs also include the reclamation costs for closure.

Growth capital costs in the Reserve Case includes costs for:

•Ardich establishment and mine development

•Heap leach phase 5 and phase 6

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•Road relocation, studies, and project management

•Explosives magazine relocation

Sustaining capital in the Reserve Case includes costs for:

•TSF expansion

•Project team

•Technical services

•Administration

•Assay laboratory

•Mining

•IT

•Sulfide and oxide processing

•Environment

•Mineral / lands rights

•Health and safety

•Security

•Supply chain

•Reclamation

The CDMP21TRS Reserve Case capital costs in 2022 for heap leach is $7.8M and for POX is $28.53M. Total capital over the life-of-mine (LOM) including reclamation and closure is $626M.

1.17.2 Operating Costs

Operating costs were estimated based on current site cost performance and contract costs, including actual operational costs for labour, consumables, contracts, and the Anagold budget assumptions. Operating costs have a base date of Q4’21 with no allowance for escalation. LOM average operating costs are shown in Table 1.9.

Table 1.9 Summary of LOM Average Operating Costs

Cost	Total LOM<br>($M)	5-Year Average<br>per year<br>($/t ore)	LOM Average<br>per year<br>($/t ore)
Mining	766	14.98	10.15
Process	2,225	27.79	29.49
Site Support and G&A	473	7.14	6.27
Operating Costs	3,464	49.91	45.91

Mining costs include waste stripping costs

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1.18 CDMP21TRS Reserve Case

The economic analysis was prepared on a 100% project basis using the Reserve Case production schedule, operating, and capital assumptions on an annual basis. The CDMP21TRS Reserve Case production includes 22.6 Mt at 1.69 g/t Au oxide ore processed by heap leaching and 52.9 Mt at 2.33 g/t Au processed in the sulfide plant on a 100% project basis. Total production is 75.4 Mt at 2.14 g/t Au. Total gold production is 4.4 Moz. Mining at the Çöpler pit is completed in 2029 and at Ardich in 2034. Oxide heap leach stacking is completed in in 2034, while sulfide processing will continue from stockpiles until 2042. The Reserve Case results include:

•After-tax NPV at a 5% real discount rate is $1.73 billion

•Mine life of 21 years

An IRR is not reported as the operation is cash positive in each year of the mine plan until closure. The Reserve Case average all-in sustaining cost (AISC) is $966/oz gold. Key results of the Reserve Case economic analysis are shown in Table 1.10.

Table 1.10 CDMP21TRS Reserve Case Results Summary

Item	Unit	Reserve Case
Oxide Processed
Heap Leach Quantity	kt	22,557
Au Feed Grade	g/t	1.69
Sulfide Processed
Quantity Milled	kt	52,892
Au Feed Grade	g/t	2.33
Total Processed
Processed	kt	75,448
Au Feed Grade	g/t	2.14
Total Gold Produced
Oxide – Gold	koz	765
Sulfide – Gold	koz	3,604
Total – Gold	koz	4,369
Oxide – Gold Recovery	%	61
Sulfide – Gold Recovery	%	91
5-Year Annual Average
Average Gold Produced	kozpa	278
Free Cash Flow	$Mpa	158
Total Cash Costs (CC)	$/oz gold	880
All-in Sustaining Costs (AISC)	$/oz gold	1,071
Key Financial Results
Life-of-Mine (LOM) CC	$/oz gold	803
LOM AISC	$/oz gold	966
Site Operating Costs	$/t treated	45.91
After-Tax NPV5%	$M	1,732
Mine Life	years	21

5-Year Annual Average is for the period 1 January 2021 through 31 December 2026

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The after-tax cash flow is shown in Figure 1.9. The NPV results for before and after-tax over a range of discount rates is shown in Table 1.11. The sulfide and oxide production profiles are shown in Figure 1.10 and gold production in Figure 1.11. Cash costs are shown in Table 1.12.

Figure 1.9 Reserve Case After-Tax Cash Flow

OreWin, 2022

Table 1.11 CDMP21TRS Reserve Case Before and After-Tax NPV

Discount Rate	Before-Tax NPV<br>($M)	After-Tax NPV<br>($M)
Undiscounted	2,729	2,555
5%	1,824	1,732
10%	1,322	1,268
12%	1,185	1,140

A financial model was prepared using the Reserve Case production schedule and operating and capital assumptions on an annual basis. The assumptions for taxes and royalties were provided by Anagold. The corporate tax rate in Turkey is 23% in 2022 but will revert to 20% from 2023. The royalty rate for precious metals under Turkish Mining Law is variable and tied to metal prices. As Çöpler ores are processed on site, the applicable royalty rate is subject to a further 40% reduction for certain qualifying operating costs. The average royalty calculated as a proportion of gross revenue in the CDMP21TRS Reserve Case is approximately 4.9%.

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Figure 1.10 CDMP21TRS Reserve Case Processing

OreWin, 2022

Figure 1.11 CDMP21TRS Reserve Case Gold Production

OreWin, 2022

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Table 1.12 CDMP21TRS Reserve Case Cash Costs

Description		Reserve Case
Mining and Rehandle		766
Process, Freight, and Refining		2,031
Site Support		393
Royalties		353
Total Production Costs		3,543
Total Cash Costs (CC)		803
Sustaining Capital		442
Fixed Lease Payments		192
Site G&A		81
All-in Sustaining Costs (AISC)		4,257
	/oz gold

All values are in US Dollars.

Process, Freight, and Refining includes by-product credits and excludes fixed lease costs.

Royalties are calculated in the period incurred and applied to cash flow in the subsequent year.

Metal prices were estimated after analysis of consensus industry metal price forecasts and metal prices used in other comparable studies. The metal prices used for the economic analysis are shown in Table 1.13.

Table 1.13 CDMP21TRS Economic Analysis Metal Price Assumptions

Metal Price	Units	2022	2023	2024	2025	Long- Term
Gold	$/oz	1,800	1,740	1,710	1,670	1,600
Silver	$/oz	24.00	23.00	22.00	21.00	21.00
Copper	$/lb	4.00	3.80	3.80	3.80	3.40

The estimates of cash flows have been prepared on a real basis with a base date of Q4’21 and a mid-year discounting is used to calculate NPV. All monetary figures have a base date of Q4’21 with no allowance for escalation and are expressed in US dollars (US$) unless otherwise stated.

1.19 Interpretation and Conclusions

Mineral Resources were originally discussed and disclosed in the 2021MR. Some content from the 2021MR report has been used in the CDMP21TRS. Sections relating to the Mineral Resource have not been changed from the 2021MR, as that remains the most current study work available.

Mineral Resources and Mineral Reserves in the CDMP21TRS are reported in accordance with subpart 1300 of US Regulation S-K Mining Property Disclosure Rules (S-K 1300).

The CDMP21TRS has identified additional Mineral Resources and additional Mineral Reserves when compared to prior studies.

The 2021MR Initial Assessment has demonstrated that there is significant economic potential that may be derived from the copper in the Çöpler Mineral Resource. Given this economic potential it is then concluded that it is valid to report the Mineral Resources using the Mineral Resource metal prices and the Mineral Resource pit shell.

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Further study and analysis will be required to advance the understanding of this potential.

1.20 Recommendations

Key recommendations from the CDMP21TRS are:

•Continue to update and evaluate the Çöpler District Master Plan as the existing Mineral Resources and Mineral Reserves are updated and as new prospects are advanced.

•Undertake infill drilling at Çöpler and update the copper Mineral Resource estimate.

•Prepare further studies of the copper recovery options.

•Conduct Geotechnical reviews and re-evaluation of updated pit designs.

•Optimisation of the sulfide flotation circuit, POX, and process operation.

•Metallurgical testwork on future oxide, sulfide, and copper ore sources.

•Optimisation of the oxide heap leach circuit.

•Optimisation of the mining rates to increase gold production.

•Stockpile reconciliation and management studies.

•Review and adapt ore control and stockpiling strategies to optimise recovery and throughput and maximise gold production.

•Reconcile monthly blend and gold production with predictive modelling.

•Continue exploration drilling at Ardich.

•Conduct geotechnical studies of Ardich.

•Conduct reconciliation studies of Çöpler.

•Update Çöpler and Ardich resource models and estimates.

•Further study of the 2021MR Initial Assessment Cases and advance to next stage of study:

•Geotechnical studies

•EIA and permitting

•Blasting studies

•Metallurgical studies

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2 INTRODUCTION

The CDMP21TRS has been prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300).

Exploration tenures surrounding the Project area and mining at Çakmaktepe are subject to joint venture agreements between SSR and Lidya that have varying interest proportions. SSR controls 50% of the shares of Kartaltepe Madencilik Sanayi ve Ticaret Anonim Şirketi (Kartaltepe) and 30% of Tunçpinar Madencilik Sanayi ve Ticaret Anonim Şirketi (Tunçpinar). Lidya holds the remaining 50% of Kartaltepe and 70% of the Tunçpınar. Ownership percentages of the Mineral Resources are shown in Table 1.6 and of the Mineral Reserves in Table 1.8.

In most cases, the parent company will be referred to as SSR throughout this Technical Report Summary even though it may have been Alacer or Anatolia at the time referenced in the report. Anagold remains the operating company for the Çöpler project and is the entity that undertakes the day-to-day work for the Project.

SSR has reported that the total cost of the gross mineral properties, and plant and equipment as of 31 December 2021 was $2,761.6M.

2.1 Terms of Reference

The CDMP21TRS is an independent Technical Report Summary (TRS) on the Çöpler project, prepared for SSR by the CDMP21TRS Qualified Persons (QPs) as part of the strategy for expansion of the Çöpler project. The TRS is based on information and data supplied to the QPs by SSR and other parties where necessary. Any individual or entity referenced as having completed work relevant to the CDMP21TRS, but not identified therein as a QP, does not constitute a QP. The CDMP21TRS QPs have reviewed the supplied data and information and accept this information as being accurate and complete and suitable for use in the CDMP21TRS. The primary source of data for the CDMP21TRS is the Çöpler District Master Plan 2021.

Mineral Resources were originally discussed and disclosed in Çöpler District Mineral Resource 2021 Technical Report Summary (2021MR). The 2021MR is an independent Technical Report Summary prepared to provide a preliminary technical and economic study of the economic potential of the Çöpler District mineralisation to support the disclosure of Mineral Resources. Some content from the 2021MR report has been used in the CDMP21TRS.

Information and data supplied by SSR that were outside the areas of expertise of the QPs and was relied upon when forming the findings and conclusions of this report are detailed in Section 25.

The QPs have used their experience and industry expertise to produce the estimates and approximations in the CDMP21TRS. It should be noted that all estimates and approximations contained in the CDMP21TRS will be prone to fluctuations with time and changing industry circumstances.

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The purpose of the CDMP21TRS is to report the Mineral Resources and Mineral Reserves for the Project. This report is a Feasibility Study (FS) that represents forward-looking information. The forward-looking information includes metal price assumptions, cash flow forecasts, projected capital and operating costs, metal recoveries, mine life and production rates, and other assumptions used in the FS. Readers are cautioned that actual results may vary from those presented. The factors and assumptions used to develop the forward-looking information, and the risks that could cause the actual results to differ materially are presented in the body of this report under each relevant section.

The conclusions and estimates stated in the CDMP21TRS are to the accuracy stated in the CDMP21TRS only and rely on assumptions stated in the CDMP21TRS. The results of further work may indicate that the conclusions, estimates and assumptions in the CDMP21TRS need to be revised or reviewed.

The CDMP21TRS is intended to be used by SSR, subject to the terms and conditions of its contract with OreWin. Recognising that SSR has legal and regulatory obligations, OreWin has consented to the filing of the CDMP21TRS with US SEC. Except for the purposes legislated, any other use of this report by any third party is at that party's sole risk.

A list of the references used to prepare the CDMP21TRS is provided in Section 24.

2.2 Qualified Persons

The following people served as the QPs as defined in subpart 1300 of US Regulation S-K Mining Property Disclosure Rules (S-K 1300):

•Bernard Peters, BEng (Mining), FAusIMM (201743), employed by OreWin Pty Ltd as Technical Director – Mining, was responsible for the overall preparation of the CDMP21TRS and, the Mineral Reserve estimates, Sections 1 to 4; Sections 5 and 6; Section 13; and Sections 15 to 27.

•Sharron Sylvester, BSc (Geol), RPGeo AIG (10125), employed by OreWin Pty Ltd as Technical Director – Geology, was responsible for the preparation of the Mineral Resources, Sections 1 to 4; Sections 7 to 12; Section 14; and Sections 25 to 27.

2.3 Qualified Persons Property Inspection

Site visits were performed as follows:

•Mr Bernard Peters visited the Project 13–17 May 2019, 15–21 September 2019, 14–18 October 2019, 18–21 November 2019, and 27 February to 4 March 2020. The site visits included briefings from geology and exploration, mine, processing, environmental, permitting, and corporate personnel, site inspections of current and future areas for mining and plant and infrastructure, and discussions with other consultants. In addition, several visits to SSR’s head office in Denver Colorado were undertaken during the same timeframe for the purpose of project-related meetings.

•Sharron Sylvester visited the Project 13–17 May 2019, 15–21 September 2019, 14–18 October 2019, 18–21 November 2019, and 27 February to 4 March 2020. The site visits included briefings from geology and exploration, mine, processing, environmental, permitting, and corporate personnel, site inspections of current and future areas for mining and plant and infrastructure, and discussions with other consultants. In addition, several visits to SSR’s head office in Denver, Colorado were undertaken during the same timeframe for the purpose of project-related meetings. Visits to analytical laboratories were planned to be undertaken but not completed due to global travel restrictions related to Covid-19.

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2.4 Units and Currency

This Technical Report Summary uses metric measurements except where otherwise noted. The currency used is US dollars ($) unless otherwise stated.

2.5 Effective Dates

The report has several effective dates, as follows:

•Effective date of the Technical Report Summary: 31 December 2021

•Date of drillhole database close-out for the Çöpler Mineral Resource estimate: 15 July 2015

•Date of drillhole database close-out for the Çakmaktepe Mineral Resource estimate: 31 October 2019

•Date of drillhole database close-out for the Ardich Mineral Resource estimate: 29 May 2021

•Effective date of Mineral Resources: 31 December 2021

•Effective date of Mineral Reserves: 31 December 2021

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3 PROPERTY DESCRIPTION

3.1 Location

The CDMP21TRS is an independent Technical Report Summary, in accordance with S-K 1300, prepared for SSR on the Çöpler project (the Project), located in Turkey. The project consists of several mining licences covering Mineral Resources on the Çöpler, Çakmaktepe, Ardich, and Bayramdere deposits, Mineral Reserves on the Çöpler and Çakmaktepe open pit mines, oxide and sulfide processing facilities, and supporting infrastructure.

The Çöpler project is in east central Turkey, 120 km west of the city of Erzincan, in Erzincan Province, 40 km east of the iron-mining city of Divriği (one hour drive), and 550 km east of Turkey’s capital city, Ankara. The nearest urban centre, Iliç, (approximate population 3,800), is located approximately 6 km east of the current Çöpler pit. Figure 3.1 illustrates the location of the Project within the country of Turkey and indicates the deposit’s proximity to surrounding communities.

The Çöpler project uses the European 1950 (E1950) datum coordinate system, which is a Turkish Government requirement. The Project is in UTM6 zone 37N of the E1950 coordinate system, and its centroid is situated at approximately 459,975 mE and 4,364,420 mN and has an approximate elevation of 1,160 m above mean sea level (mamsl).

The Çöpler mining operations are located 900 m south-west of the Iliç district centre, 650 m south of the Bahçe neighbourhood, 250 m south of the Çöpler village, and 180 m north of the Sabırlı village. The project site lies within the licence areas numbered 847, 49729, and 20067313 (Figure 3.2), which have been granted by the General Directorate of Mining and Petroleum Affairs (MAPEG).

The Çakmaktepe satellite mining operation is located 6 km east of the current Çöpler pit and 1.5 km south of Iliç. The Çakmaktepe pits are located within Kartaltepe Licence 1054. Ore mined at Çakmaktepe is hauled and treated at the Çöpler facilities.

The Çöpler operation’s currently permitted Environmental Impact Assessment (EIA) boundary incorporates 1,747 ha, whereas the footprint of the mine units covers a combined 1,089 ha. The currently permitted Çakmaktepe EIA boundary incorporates 290 ha. Pending approval, Çakmaktepe EIA boundary will increase to 360 ha with the second capacity increase.

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Figure 3.1 Location of the Project

Anagold, 2020

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Figure 3.2 Çöpler Project Licence and Surrounding Licences (UTM Grid)

Anagold, 2021

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3.2 Ownership

SSR controls the Çöpler project through a series of companies that own the licence areas. The company structure that links SSR to the Çöpler project is shown in Figure 3.3.

Figure 3.3 Ownership

1 Lidya holds 18.5% of this entity and Bank Kombetare Tregtare SHA, a bank wholly owned by Çalık Holdings A.Ş., holds the remaining 1.5%.

2 Lidya holds the remaining 50% of Kartaltepe and 70% of the Tunçpınar.

The Çöpler project (the Project) is owned and operated by Anagold Madencilik Sanayi ve Ticaret Anonim Şirketi (Anagold). SSR controls 80% of the shares of Anagold, Lidya Madencilik Sanayi ve Ticaret A.Ş. (Lidya), controls 18.5%, and a bank wholly owned by Çalık Holdings A.Ş., holds the remaining 1.5%.

Ownership percentages of the Mineral Resource are shown in Table 1.6 and of the Mineral Reserves in Table 1.8.

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The license that hosts the Çöpler deposit, including the Mineral Resources and Mineral Reserves, is wholly owned by Anagold. Çakmaktepe is wholly-owned by Kartaltepe. Ardich, Mavialtin, Bayramdere, Aslantepe, and Findiklidere have areas owned by both Anagold and Kartaltepe.

The Initial Assessment has only analysed Mineral Resources located on the Anagold licence.

3.3 Mineral Tenure

Anagold holds the exclusive right to engage in mining activities within the Çöpler project area. Anagold holds six granted licences (Table 3.1) covering a combined area of approximately 16,600 ha. Mineral title is held in the name of Anagold. Kartaltepe holds eight licences covering approximately 9,200 ha. The total near-mine tenement package is approximately 25,800 ha. Anagold currently holds sufficient surface rights to allow continued operation of the mining operation in the CDMP21TRS Reserve Case. The major licence boundaries are shown in Figure 3.4.

The granted licences include two clay borrow pit licences, numbered 76817 and 76818.

The Çöpler mine and associated infrastructure are hosted within the triangular-shaped concession 847. Anagold has received approval from the Mining Affairs Committee to grant extensions to the three Anagold licences that had expired (76817, 76818, and 50237). Licenses 76817 and 76818 have been extended to 15 July 2029 and Licence 50237 has been extended until 21 March 2028.

Anagold has confirmed that charges and administrative expenses due to the Turkish Ministry of Energy and Natural Resources, Directorate General of Mining and Petroleum Affairs (MAPEG) have been paid, and all Anagold licences were in good standing as of 31 December 2021.

The mined Çakmaktepe pits are all on Kartaltepe Licence 1054. Bayramdere prospect is on Kartaltepe Licence 7083.

The three expired Kartaltepe licences (200707602, 200707605, and 200707606) were combined, and an operation project was prepared and submitted to receive an operation licence. The process continues. Kartaltepe maintains ownership of these licences during this process.

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Table 3.1 Granted Licences and Operating Permits

Province	Town	Village	Registration No.	Licence No.	Licence Area (ha)	Licence Type	Licence Group	Operation Permit	Operation Permit Area (ha)	Licence Issue Date	Licence Expiry Date	Licensee	Project
Erzincan	İliç	Çöpler	1027313	847	941.92	Operation	IV (Metallic)	Au+Ag+Cu+Hg Mn	Au+Ag+Cu+Hg: 941.92 Mn: 941.92	6/11/1986	6/11/2026	Anagold	Çöpler-Çöpler Saddle
Erzincan	İliç	Çöpler	2384036	49729	13,747.51	Operation	IV (Metallic)	Au+Ag+Cu+Mo	909.50	4/08/2016	4/08/2026	Anagold	Ardich-Çöpler Saddle-Kiziltepe-Meseburnu
Erzincan	İliç	Ortatepe	2386272	50237	600.00	Operation	IV (Metallic)	Au	18.07	21/03/2018	21/03/2028	Anagold	Elmadere-Demirmagara
Erzincan	İliç	Sabırlı	3095732	20067313	1,184.91	Operation	IV (Metallic)	Au+Ag+Cu	216.41	25/10/2021	25/10/2031	Anagold	Çakmaktepe Se-Ardich
Erzincan	İliç	Çöpler	3201587	76817	49.32	Operation	I-B (Brick Tile Clay)	Clay	6.68	15/07/2019	15/07/2029	Anagold	Clay Licence
Erzincan	İliç	Çöpler	3201588	76818	49.09	Operation	I-B (Brick Tile Clay)	Clay	49.09	15/07/2019	15/07/2029	Anagold	Clay Licence
Total (ha)					16,572.75
Erzincan	Kemaliye	Kabataş	2450158	57004	1,564.69	Operation	IV (Metallic)	Au+Cu	931.87	2/09/2018	2/09/2023	Kartaltepe	Mavidere
Erzincan	Kemaliye		3129489	200707602	1,572.23	Pending Operation	IV (Metallic)	–	–	2/08/2007	2/08/2012	Kartaltepe	Mavidere
Erzincan	Kemaliye		3129490	200707605	577.92	Pending Operation	IV (Metallic)	–	–	2/08/2007	2/08/2012	Kartaltepe	Mavidere
Erzincan	Kemaliye		3129496	200707606	1,818.11	Pending Operation	IV (Metallic)	–	–	2/08/2007	2/08/2012	Kartaltepe	Mavidere
Erzincan	İliç		1032544	58473	606.60	Operation	IV (Metallic)	Fe+Cu	7.54	16/11/2017	16/11/2027	Kartaltepe	Findiklidere
Erzincan	İliç	Yakuplu	1032719	7083	1,756.55	Operation	IV (Metallic)	Au+Ag+Cu+Fe Cr	Au+Ag+Cu+Fe: 175.00 Cr: 607.47	2/04/2011	2/04/2021	Kartaltepe	Bayramdere-Aslantepe-Saridere
Erzincan	İliç	Yakuplu	1027026	1054	660.87	Operation	IV (Metallic)	Au+Ag+Cu+Fe	359.33	30/07/2017	30/07/2027	Kartaltepe	Çakmaktepe
Erzincan	İliç	Ortatepe	2003094	7161	642.68	Operation	IV (Metallic)	Fe	214.65	7/05/2013	7/05/2023	Kartaltepe	Ortatepe
Total (ha)					9,199.65

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Figure 3.4 Tenure Layout Plan

Anagold, 2020

3.4 Surface Rights

Anagold currently holds sufficient surface rights to support the CDMP21TRS Reserve Case oxide heap leach mining operations and sulfide processing and tailings disposal.

3.5 Taxation

The Turkish government implemented a temporary taxation rate increase from 20% to 22% for the period 2018 through 2020, 25% for 2021, and 23% for 2022. From 2023 onwards, the effective tax rate is expected to return to 20%.

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The CDMP21TRS economic analysis applies a corporate tax rate of 20%.

For tax purposes, a 20% accelerated depreciation rate is applicable for both the oxide and sulfide capital. The depreciation period is 10 years for general mining equipment, if not specifically defined by the tax office.

Investment incentive certificates are available for investments that promote economic development. Investment incentive certificates can be classified as strategic in specific circumstances and such certificates provide additional incentives. Anagold received a strategic incentive certificate for the sulfide process plant. An investment incentive certificate generates credits that offset corporate income taxes generated by the investment. The amount of investment credits generated from the investment incentive certificate is based on eligible capital expenditures. The investment credits generated by the strategic investment incentive certificate reduce the corporate tax rate to a minimum of 2% in a given tax period until Q4'23, thereafter it is assumed subsequent non-strategic investment incentive certificates will be available and the minimum rate will be 4%. Incentive tax credits can be carried forward to future tax periods indefinitely until exhaustion. Incentive tax credits and other tax pools are determined in the local currency, Turkish Lira, and subject to devaluation and revaluation as fluctuations against the US dollar occur. The cash flow model is prepared on a constant Turkish Lira basis.

Value-added tax (VAT) in Turkey is levied at 18% and the Project is eligible for the Turkish exemptions for mining projects and mining equipment purchases. In the CDMP21TRS assumes the cash flows are not subject to VAT.

Import duties are not included in the capital cost estimate for mining related imported equipment because they are exempted in the incentive certificates.

3.6 Royalties

Under Turkish Mining Law, the royalty rate for precious metals is variable and tied to metal prices. The Çöpler project is subject to a mineral production royalty that is based on a sliding scale to gold price and is payable to the Turkish government. In September 2020 a presidential decree was issued, increasing the prescribed royalty rates by 25%.

Table 3.2 details the relevant prescribed royalty rates along with the revised rates following the September 2020 presidential decree. The royalties are calculated on total revenue with deductions allowed for processing and haulage costs of ore. Revenue from by-products (silver and copper) is included in the total revenue used for royalty calculations.

The royalty rates outlined in Table 3.2 apply to gold production from heap leaching. Royalty rates are reduced by 40% for ore processed in country, as an incentive to process ore locally. As the Çöpler project produces its gold doré on site, the Çöpler project is eligible for a 40% reduction to the royalty rate for gold produced from POX processing.

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Table 3.2 Gold Royalty Rates

Metal Price(/oz Gold)	Prescribed Royalty <br>Rate<br><br>(%)	Royalty After 40% In-Country Processing Incentive<br>(%)
From
0	1.25	0.50
800	2.50	1.00
900	3.75	1.50
1,000	5.00	2.00
1,100	6.25	2.50
1,200	7.50	3.00
1,300	8.75	3.50
1,400	10.00	4.00
1,500	11.25	4.50
1,600	12.50	5.00
1,700	13.75	5.50
1,800	15.00	6.00
1,900	16.25	6.50
2,000	17.50	7.00
2,100	18.75	7.50

All values are in US Dollars.

The Çöpler project effective life-of-mine (LOM) royalty rate based on the financial model metal price assumptions and applicable deductions is approximately 4.9%.

Other than the royalty payments, there are no other known back-in rights, payments, or other agreements and encumbrances to which the property is subject.

3.7 Environmental Liabilities

There are no known existing environmental liabilities for the Çöpler project, except for Anagold’s obligation for ultimate reclamation and closure.

3.8 Permits

The EIA permitting for the Çöpler mine oxide ore was completed in April 2008 with the issuance of an EIA positive certificate. All the necessary operation permits have already been obtained for the oxide inventory. These include:

•Explosive storage permit

•Permit for water abstraction from groundwater sources

•EIA positive certificate for power transmission line construction

•Environmental permits and licences

•Land acquisition permits for forest areas and pasturelands

•Workplace opening permit

•Hazardous workplace permit

•Operating permits.

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The EIA permitting process for the Sulfide Expansion Project was commenced on 7 April 2014 and completed with the receipt of an ’EIA Positive Statement‘ on 24 December 2014. In addition to an EIA approval, other permits required for the Sulfide Expansion Project involved an expanded workplace opening permit, additional operating permits, and land acquisition permits for forest areas and pasture lands.

Additional EIA studies conducted, and environmental permits received for the Çöpler mine since the start of the gold mining operations are as follows:

•EIA permit, dated 10 April 2012, for the operation of mobile crushing plant.

•EIA permit, dated 17 May 2012, for the capacity expansion involving:

•Increasing operation rate to 23,500 tpd.

•Increasing Çöpler waste rock dump (WRD) footprint area.

•Adding a sulfidisation, acidification, recovery, and thickening (SART) plant to the process in order to decrease the cyanide consumption due to the high copper content of the ore.

•EIA permit, dated 24 December 2014, for the capacity expansion involving:

•Sulfide plant expansion

•Heap leach area expansion

•EIA permit, dated 26 January 2017, for the Çakmaktepe satellite pits expansion.

•EIA permit, dated 9 August 2018, for the Çakmaktepe expansion for the new defined Central pit.

•EIA permit dated 7 October 2021 for the second capacity expansion involving:

•Heap leach pads 5 and 6

•TSF expansion

•Operation of floatation plant.

3.9 Other Significant Factors and Risks

SSR have advised that there are no other known significant risks that may affect access, title or the right or ability to perform mining related work on the Property.

Legal matters such as statutory and regulatory interpretations affecting the mine plan and environmental matters are outside the expertise of the QPs (see Section 25).

The CDMP21TRS QPs considers it reasonable to rely on SSR because SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

Additionally, the Project has been in operation for a number of years, and following a review of the current supplied information, the opinion of the QPs is the current plans appear adequate to address any issues related to environmental compliance, permitting, and local individuals or groups.

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4 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

4.1 Accessibility

The Çöpler project is accessed from the main paved highway between Erzincan and Kemaliye, crossing the Karasu River and passing by the village of İliç. From İliç there is an additional 4.5 km of road to reach the Çöpler mine site.

The Ankara to Erzincan railway line, operated by the Turkish State Railway Company, (TCDD), runs parallel to the south bank of the Karasu River and passes within 2 km north of the site at a point between the train stations at İliç and Bağıştaş. The railway line connects the site with Ankara and the west as well as with seaports to the north on the Black Sea, and to the south on the Mediterranean Sea. Overnight passenger sleeper cars are available between Erzincan and Ankara.

The reservoirs of the Bağıştaş I and II hydro-electric power plants (HEPP) are 350 m and 1,800 m away from the Çöpler mine site, respectively. The embankment of Bağıştaş I Dam originally covered a portion of the existing highway, railroad, and railroad station until these were relocated before dam construction was completed. Construction routes for the railroad and highway were located between the new Çöpler village and the Çöpler mine site. The bridge on the north-east side of İliç was relocated to further east of the embankment.

There are regular commercial airline flights from Istanbul and Ankara to the regional cities of Erzincan, Erzurum, Malatya, Elazığ, and Sivas. Driving from the regional cities to the Project site takes between two to four hours on paved highways. Driving from Ankara to the site takes approximately eight hours.

4.2 Local Resources and Infrastructure

The district of İliç has a population of approximately 3,800 inhabitants and is located approximately 6 km east of the current Çöpler pit. The district has a hospital, schools, municipal offices, a fire station, a police station, and a Gendarmerie post. The primary economic activity in the region is sheep herding for wool, meat, and dairy products. Other agricultural activities include bee keeping for honey production and, some wheat farming along the Karasu River. Additionally, there is some light manufacturing and grain milling performed in İliç.

The workforce for the Anagold exploration programmes has primarily included residents drawn from the local communities of Çöpler, İliç, and Sabırlı.

Turkish telecommunications are up to European standards. High-speed, fibre-optic internet access is in operation at the mine site.

Initially, electrical power at 380 V and 50 Hz, was available in İliç and at the mine site. This was upgraded to support the Project by the construction of a 40 km long 154 kV power line from the sub-station at Divriği to the mine site. The power supply was further upgraded when the hydroelectric dam near the mine site was commissioned. Çöpler is now connected to the national grid by a 6 km 154 kV powerline from the Bağıştaş sub-station.

Sufficient local fresh water supply exists to support the mining and processing operations. Ground water resources include seven production wells with a 25,728 m3/day extraction permit. Further information on project infrastructure is included in Section 15. Section 17.3 contains additional data on the Project social setting.

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4.3 Climate

Site climate data were developed during previous studies. No additional climate data were generated for the CDMP21TRS.

Mining operations are conducted year-round. The climate is typically continental with cold wet winters and hot dry summers. In winter, the night-time temperature can drop to –25°C although the average is usually a few degrees below freezing. The July temperature frequently exceeds +40°C but the climate is usually pleasantly warm outside of these extremes. The average monthly temperature ranges from +3.7°C for the coldest month of January to +23.9°C for August, the warmest month.

Most precipitation occurs in the winter and spring. Monthly average rainfall values are shown in Figure 4.1. The average annual rainfall for the site is 384.3 mm. Snowfall is common during the period mid-November through February, but with little, if any, accumulation. Snow depth assessments are based on the Divriği meteorological weather station, located 41 km west of the Project area, which shows maximum snow-pack depths at approximately 200 mm for 1985.

Figure 4.1 Average Monthly Rainfall for Çöpler Project Area

Anagold, 2016

The frost depth is less than 0.3 m, based on local information, with 0.5 m selected as the design frost depth limit.

The maximum wind speed recorded at the Divriği weather station in 2004 ranges from 15-25 m/s, with variable directions mainly from the north, south, and east.

4.4 Hydrogeology

SRK compiled and updated the Project conceptual hydrogeological model with new geological data, established a new numerical model and used it to evaluate the hydrogeology of the Project area.

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4.4.1 Existing Data Evaluation, Field Investigation, and Hydrogeology Conceptual Model

Within the regional hydrology area, lithological units are defined in three main classes according to their underground water transport and transmission properties. These units are:

•Impervious units.

•Low permeate units: such units contain some thin layers that are more permeable than other layers with small extensions and provide water through sources with a flow rate of less than 1 L/s.

•Conductive units and very permeable units: Munzur Formation limestone and Quaternary alluvium units.

The regional geology is a complex structural assemblage of fault-bounded blocks including the following rock types:

•Limestone: grey to blue-grey, fine-grained to recrystallised marbles. Much of the unit displays various degrees of karst development. Bedding within the unit is indistinct to massive. This limestone group is also named the Çöpler limestone in the vicinity of the area where Mineral Resources have been estimated.

•Metasediment: fine-grained argillite sequences consisting of interbedded siltstones, shale units, marls, and sandy siltstones. The thermal and hydrothermal impact to this unit from the intrusions resulted in the creation of the skarns and hornfels.

•Ophiolitic mélange: ophiolitic mélange consists of diabase and serpentinite units. Serpentinisation is non-uniform and appears to be best developed near major fault zones.

•Diabase: the diabase is located within the upper zone of the ophiolitic mélange. In general, joint surfaces are covered with calcite and iron oxide sealing. In places, the rock mass shows blocky textures embedded in a fine matrix.

•Diorite to granodiorite intrusions: beige and light brown, medium to coarse-grained plutons. This formation has intruded into the pre-existing argillite’s and Munzur limestone. This includes fine to medium-grained quartz, feldspar, biotite, and amphibole minerals.

•Skarn: the skarn zone is developed along the granodiorite contact with the limestone and ophiolitic mélange. This zone was developed under elevated pressure and temperature conditions during intrusion of the granodiorite mass. The skarn units are black to dark brown, silicified, moderately weathered and includes frequent solution cavities.

4.4.2 Well Installation

A total of 56 wells for groundwater observation, testing, and water supply purposes have been drilled. Forty-one of the wells were drilled prior to 2018, ten were for groundwater control and slope stability studies in 2018, two were for waste storage area observation purposes, and three were developed in 2018 as part of the sulfide expansion project for additional water supply. Hydrogeology wells drilled are shown in Figure 4.2.

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Figure 4.2 Groundwater Wells

Anagold, 2020

Groundwater is expected to be recharged through the infiltration of precipitation through secondary porosity in the bedrock terrain. Groundwater elevation data indicates that the flow direction is generally northward to the Karasu River through the Munzur limestone. During the resource drilling and subsequent monitoring well installation programmes, perched groundwater conditions were reported above the clay-altered intrusions. It is anticipated that the perched groundwater is present in restricted areas. The volume of water held in storage as perched groundwater is unknown.

Groundwater elevations at the Çöpler project range from 1,328.5 m at Well GMW-10 (southern end of the site) to 864.7 m at Well GMW-09 (northern end of the site). Observations of cavernous features (karst) during drilling and high-values of hydraulic conductivity from aquifer tests suggest an area of karst development in the limestone near the Karasu River, at boreholes GMW-09 and GMW-24. This was incorporated into the groundwater flow model as an area of high hydraulic conductivity near these wells and along the Sabırlı Fault.

4.5 Physiography

The Çöpler project is located in a roughly east–west oriented valley at altitudes of 1,100-1,300 m. The valley is surrounded by limestone mountains that rise to more than 2,500 m on the north and south sides of the Project area. These mountains are at the western end of the Munzur range, which rises to more than 3,300 m between Ovacık and Kemah.

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The region is sparsely vegetated, predominantly with semi-arid brush and scrub trees including dwarf oaks and junipers.

The following are the site data developed during previous studies for the design of the Project:

•Latitude: 39° 25’ North

•Longitude: 38° 32’ East

•Elevation: 1,150 mamsl

•Frost depth: 500 mm

•Snow load: 145 kg/m2

•Wind load: 40 m/sec, Exposure ‘C’

•Earthquake zone: second order, Ao = 0.20

•Atmospheric pressure (average): 880.5 millibars

•Maximum design temperature: +40°C

•Minimum design temperature: –25°C

•Annual rainfall: 384 mm

•Maximum snowfall depth: 200 mm (estimated)

•Design maximum 24-hour rainfall: 76 mm

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5 HISTORY

The region around the Çöpler project has been subject to gold and silver mining dating back at least to Roman times, with historical bullion production estimated at approximately 50 koz of gold. A copper-rich slag pile of approximately 2.5 kt is located at the western edge of the district and is believed to be waste from ancient production. Although the district contains copper mineralisation, there appears to have been little production targeting copper. There are several additional minor slag piles scattered around the property thought to be from ancient, small-scale gold and by-product copper production.

The Turkish Geological Survey (MTA) carried out regional exploration work in the early 1960s that was predominately confined to geological mapping. In 1964, a local Turkish company started mining for manganese, continuing through until closing in 1973 and producing approximately 7.3 kt of manganese ore during its active life. Unimangan Manganez San A.Ş. (Unimangan) acquired the property in January 1979 and re-started manganese production, producing 1–5 ktpa of ore until ceasing operations in 1992.

In 1998, Anatolia Minerals Development Ltd (Anatolia) identified several porphyry-style gold-copper prospects in east central Turkey and applied for exploration licences for these prospects. This work was based upon the earlier work by MTA in the 1960s. During this effort, Anatolia delineated a prospect in the Çöpler basin formed by an altered and mineralised granodiorite, intruded metasediment, and limestone. This prospect and the supporting work were the basis for a joint venture agreement for exploration with Rio Tinto.

During the period of the joint venture, exploration drilling of the Çöpler deposit was completed and a Mineral Resource estimate was developed with three mineralised zones: Main, Manganese, and Marble. In January 2004, Anatolia acquired sole control over the Project and maintained exclusivity until 2009, at which time a joint venture with Lidya was executed.

In February 2011, Anatolia merged with Avoca Resources Limited to form Alacer Gold Corp. (Alacer). In September 2020 Alacer merged with SSR.

Today the Çöpler project is owned and operated by Anagold Madencilik Sanayi ve Ticaret Anonim Şirketi (Anagold). SSR controls 80% of the shares of Anagold, Lidya Madencilik Sanayi ve Ticaret A.Ş. (Lidya), controls 18.5%, and a bank wholly owned by Çalık Holdings A.Ş., holds the remaining 1.5%.

In most cases the company will be referred to as SSR throughout the CDMP21TRS even though it may have been Alacer or Anatolia at the time referenced in the report.

The previous Technical Report was the 2020 Çöpler District Master Plan 2020 NI 43-101 Technical Report dated 27 November 2020.

The previous reporting of Mineral Resources and Mineral Reserves was in the SSR Annual Information Form. Those statements on Mineral Resources and Mineral Reserves have been used for comparison.

5.1 Previous NI 43-101 Technical Reports

The following Technical Reports have been filed on the Çöpler project (in chronological order):

•Watts, Griffis and McQuat Limited, 2003. Update of the Geology and Mineral Resources of the Çöpler Prospect, 1 May 2003.

•Independent Mining Consultants, Inc., 2005. Çöpler Project Resource Estimate, 19-October 2005.

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•Marek, J.M., Pennstrom, W.J., Reynolds, T., 2006. Çöpler Gold Project Feasibility Study, 30-May 2006.

•Marek, J.M., Moores, R.C., Pennstrom, W.J., Reynolds, T., 2007. Çöpler Gold Project, 2-March 2007, as amended 30 April 2007.

•Easton, C.L., Malhotra, D., Marek, J.M., Moores, R.C., and Pennstrom, W.J., 2008. Çöpler Gold Project East Central Turkey Preliminary Assessment Sulfide Ore Processing, 4-February 2008.

•Marek, J.M., Benbow, R.D., and Pennstrom, W.J., 2008. Çöpler Gold Project East Central Turkey, 5 December 2008 (amended and restated; supersedes 11 July 2008 version).

•Altman, K., Liskowich, M., Mukhopadhyay, D.K., and Shoemaker, S.J., 2011. Çöpler Sulfide Expansion Project Prefeasibility Study, 27 March 2011.

•Altman, K., Bascombe, L., Benbow, R.D., Mach, L., and Shoemaker, S.J., 2012. Çöpler Resource Update, Erzincan Province, Turkey, 30 March 2012.

•Altman, K., Bair, D., Bascombe, L., Benbow, R., Mach, L., and Swanson, B., 2013. Çöpler Mineral Resource Update, Erzincan Province, Turkey, 28 March 2013.

•Armstrong, D., Bascombe, L., Bohling, R., Kiel, R., Liskowich, M., Parker, H.M., Parshley, J., Seibel, G., and Swanson, B., 2014. Çöpler Sulfide Expansion Project Feasibility Study, Erzincan Province, Turkey, 29 July 2014.

•Bascombe, L., Benbow, R.D., Birch, R.G., Bohling, R., Francis, J., Khoury, C., Kiel, R., Liskowich, M., Marsden, J., Parker, H.M., Parshley, J., Seibel, G., and Statham, S., 2015. Çöpler Sulfide Expansion Project Feasibility Update, Erzincan Province Turkey, 27-March-2015.

•David, D., Kiel, R., Liskowich, M., Parshley, J., Marsden, J., Seibel, G., Parker, H., Bascombe, L., Benbow, R., Statham, S., Francis, J., and Smolonogov, S., 2016. Çöpler Mine, Erzincan Province, Turkey, 9 June 2016.

•OreWin Pty. Ltd., 2020. Çöpler District Master Plan 2020, 27 November 2020. (CDMP20TR)

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6 GEOLOGICAL SETTING, MINERALISATION, AND DEPOSIT

This section has not been changed from the 2021MR and remains the most current study work available.

The Çöpler district is located near the north margin of a complex collision zone and to the south of the prominent North Anatolian Fault Zone (Figure 6.1). The collision zone, and subsequent crustal thickening, is related to the closure of the northern branch of the Neotethys ocean, resulting from the northward subduction and coming together of the Pontides and Tauride Anatolide Block in the Late Cretaceous to Early Tertiary. In this intensely-deformed tectonic region, east–west trending imbricated structures were cut by north–north-east trending strike-slip faults during the Late Cretaceous to Paleogene period.

Figure 6.1 Geological Setting of the Çöpler District

İmer, 2012

The Çöpler district deposits, including Çöpler, Çakmaktepe, Ardich, and Bayramdere, are within the Tethyan mineral belt, a terrane stretching from Indo-China to Europe through Eurasia that contains economically significant gold, copper, and base metal deposits.

Three main rock assemblages are exposed in the Çöpler district (Figure 6.2):

•The first assemblage includes the Keban, Munzur, and Kemaliye Formations. These units are tectonically overlain by ophiolitic nappes (Ovacık Formation of Özgül and Turşucu 1984).

•The second assemblage includes Middle Eocene magmatic and sedimentary rocks.

•The third assemblage includes the Oligocene to Recent sedimentary Sivas basin rocks.

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Figure 6.2 Geological and Structural Map of the Çöpler District

Anagold, 2020

6.1 Geological Setting – Çöpler Deposit

6.1.1 Geology – Çöpler Deposit

The Çöpler deposit is centred on composite diorite to monzonite porphyry stocks that are part of the Eocene Çöpler Kabataş magmatic complex dated (by İmer et al., 2013) at:

•43.8 ± 0.3 Ma and 44.2 ± 0.2 Ma (from 40 Ar / 39 Ar analysis of igneous biotite), and

•44.1 ± 0.4 Ma (from igneous hornblende).

The magmatic rocks have intruded into both the Keban and Munzur Formations.

Rocks of the Permian to Upper Cretaceous Keban Formation shelf sequences vary in composition between siliciclastic and calcareous, with fine to medium-grained sandstone interbedded with mudstone, and locally thick sections of fine laminated mudstone. The sedimentary units are folded with a fold axis oriented at approximately 25→200 (plunge→plunge direction) resolved from bedding measurements in the Çöpler pits. Limestone of the Upper Triassic to Late Cretaceous (Upper Campanian) Munzur Formation structurally overlies the folded Keban Formation with the contact represented by cataclasite at the base of the Munzur Formation. Intense shearing of the underlying sedimentary rocks is observed, with top-to-south kinematics.

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Stratigraphically, the Munzur Formation overlies the Keban. However, mapping of the Munzur Formation to the north of Çöpler shows homoclinal structure with consistent bedding in the limestones (40 / 060, dip / dip-direction) indicating juxtaposition of structural blocks. The Munzur allochthon was thrusted onto Permo-Triassic metamorphic basement in the Late Cretaceous (Özgül and Turşucu 1984). This structural contact pre-dates Eocene Çöpler Kabataş intrusions, which appear to have intruded across the sheared contact between Keban Formation metamorphic rocks (Main Zone) and Munzur Formation limestone (Manganese Zone).

The Çöpler intrusion is a hornblende–quartz diorite-porphyry that shows strong argillic alteration. Some fresh outcrop occurs in the central part of the Main Zone and as remnants within the Manganese Zone. In its least-altered state, the diorite-porphyry is relatively pristine with well-preserved hornblende, biotite, and K-feldspar phenocrysts in a granular matrix of plagioclase and quartz with prominent magnetite. Flow alignment of the hornblende phenocrysts can be seen in places. Gradational transitions to argillic-altered rocks are evident in outcrop and drill core on a centimetre scale.

The primary control on the location of the Çöpler intrusion appears to have been the hornfels-carbonate contact. The contact of the Çöpler intrusion has a roughly rectilinear shape, suggesting control by pre-existing east–north-east trending faults, and by a set of north–north-west trending fractures. The north–north-west striking bedding may also have exerted a local control in the central part of the intrusion where many intrusive contacts are parallel to bedding and have a sill-like morphology. However, it is considered more likely that this reflects the north–north-west trending fracture control referred to above.

A pronounced ground magnetic anomaly is centred on the core of the porphyry, which has been modelled to reflect the potassically altered core of the stock-like barren porphyry system dipping steeply towards the south. In addition, there are several dykes and intrusive apophyses; most notably, a brecciated and strongly clay-altered intrusion centred on the Manganese Zone.

In the area of the Çöpler deposit, two dominant sets of faults are present. These faults are approximately parallel to the long axis of the deposit and are oriented east–north-east. These are referred to as longitudinal faults. The other set of faults are transverse to the longitudinal faults and referred to as cross-faults (Figure 6.3). The major cross-faults include from east to west; Manganese fault, Marble fault, Main Zone fault, and West fault.

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Figure 6.3 Çöpler Deposit Geological Map

Anagold, 2020

The longitudinal faults include the Northern Boundary fault (NBF), North Çöpler fault (NÇF), Central Çöpler fault, South-West Çöpler fault, and Southern Boundary fault (SBF). The Central and South-West Çöpler faults dip to the south and were previously thought to be the same fault.

Weathering has resulted in oxidation of the mineralisation close to surface. The oxidised cap is underlain by primary and secondary sulfide mineralisation. In addition to the gold–silver–copper mineralisation of economic interest, arsenic, lead, magnesium, manganese, mercury, and zinc are also present.

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6.1.2 Mineralisation – Çöpler Deposit

The gold, silver, and copper mineralisation of economic interest at Çöpler deposit area is exposed in four adjacent open pits from east to west: Manganese pit, Marble pit, Main pit, and West pit. The pits expose economic parts of the same orebody and the three eastern pits will likely join up as the mining progresses. The predominant rock types in the mine include limestone / marble, metamorphic rocks (mainly hornfels) and diorite-tonalite porphyry, locally with equigranular biotite-granodiorite intrusions. Supergene enrichment enhanced along syn-mineralisation and post-mineralisation structures plays an important role in localising high-grade gold mineralisation at lithological contacts, late-stage faults and shear zones, and fault / contact intersections.

Most of the gold mineralisation concentrated in six distinct areas in the deposit: Main, Main West, Main East, Manganese, Marble, and West. The mineralisation is considered to be related to fluids associated with diorite intrusions at depth and generally manifests as three closely related mineralisation styles across the six areas:

•Low-grade porphyry vein mineralisation.

•Intermediate sulfidation epithermal mineralisation.

•Iron skarn and carbonate replacement mineralisation.

6.1.2.1 Three Mineralisation Styles at the Çöpler Deposit

Low-Grade Porphyry Vein Mineralisation

Sub-economic porphyry copper–gold–molybdenum mineralisation is characterised by well-developed alteration zones that are complex and superimposed on each other. Late-stage porphyry mineralisation is hosted in diorite-tonalite porphyry as dominant sheeted veinlet arrays and as stockworks in metamorphic wall rocks and intruded into a gold-deplete diorite-porphyry system (Figure 6.4). Porphyry veinlets are best exposed in the Main pit since the volume of outcropping intrusions is much greater than in other areas of the mine. Early, irregular high-temperature quartz–chalcopyrite–magnetite veinlets are overprinted by ‘D’ veinlets with pyrite±quartz and symmetric feldspar-destructive phyllic halos (Figure 6.4). Dense ‘A’/’B’ veinlets occur as sheeted arrays and lesser stockworks in the intrusions but form well-developed dense stockworks in the surrounding metamorphic wall rocks (Figure 6.4). Late-stage anhydrite veinlets with pyrite and molybdenite appear to overprint the ‘D’ veins, (Tripp, 2017; internal company report).

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Figure 6.4 Çöpler Deposit Porphyry Vein Mineralisation

Anagold, 2020

Intermediate Sulfidation Epithermal Mineralisation

Intermediate sulfidation epithermal mineralisation is primarily observed in the Manganese pit as clusters of bright pink, banded, colloform, rhodochrosite base metal sulfide veins and breccia lodes, with a spatial association with elevated gold grades, (Figure 6.5). Carbonate base metal veins contain base metal sulfides sphalerite±galena±chalcopyrite in a gangue of calcite, ferroan dolomite, and/or rhodochrosite and realgar. In the Main pit, the base metal carbonate veins are coarsely crystalline whereas veins in the Manganese pit display brecciation, colloform banding, and locally quartz pseudomorphs of bladed calcite. The change in vein style suggests the Manganese pit represents a higher level position with respect to the mineralising system.

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Figure 6.5 Çöpler Deposit Intermediate Sulfidation Epithermal Mineralisation

Anagold, 2020

Iron Skarn and Carbonate Replacement Mineralisation

Iron skarn and related carbonate replacement oxide gold mineralisation developed along faults, shear zones, and within karstic spaces. It is observed as iron oxide-rich zones as well as gossan-like and jarosite units developed by oxidation of previous pyrite-rich mineralisation, (Figure 6.6). This replacement type mineralisation appears to be derived from previously formed distal skarn mineralisation. Development of gossan and jasperoid is potentially related to weathering of primary Eocene sulfide deposits in situ or remobilised from a nearby source.

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Figure 6.6 Çöpler Deposit Porphyry Vein Mineralisation

Anagold, 2020

6.1.2.2 Six Mineralisation Areas at the Çöpler Deposit

Main Zone Mineralisation

The Main Zone lies in the west portion of the Çöpler deposit and occupies a footprint of approximately 750 m north–south by 1 km east–west. Typical depths of mineralisation range to 200 m below surface. Disseminated quartz–pyrite–arsenopyrite epithermal veinlets are primarily hosted in diorite and metasediment with some marble-hosted mineralisation on the eastern margin of the zone. Oxidation and related oxide mineralisation extends to depths of approximately 40 m from surface, with the thickest oxidised zones proximal to ridges and thinning of strata in the intervening valleys.

Minor volumes of massive pyrite mineralisation occur within the Main Zone.

Main Zone West Mineralisation

Main Zone West is in the north-west corner of the Çöpler deposit at the contact between diorite, marble, and the basement metasediment. This mineralisation is hosted within narrow gossans located at the contact, and in sub-parallel veinlets containing disseminated sulfides within the marble and metasediment. Main Zone West has a strike length of approximately 750 m and is approximately 75 m wide.

Main Zone East Mineralisation

The Main Zone East represents a portion of the mineralisation lying between the Manganese Zone and Main Zone. The geology in this area is typified by narrow, weakly to moderately mineralised gossans located at the contact between the basement metasediment and the overlying marble. It is postulated that the gossan is sourced from the diorite located in the Manganese Zone and has been emplaced along the metasediment / marble contact as the diorite has crystallised.

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Manganese Zone Mineralisation

The Manganese Zone occupies the eastern end of the Çöpler deposit. This zone is approximately 650 m wide north–south by 650 m east–west. The pre-mining surface expression of this area consists predominately of marble. A moderately-sized intrusion of diorite occurs sub-surface. A large proportion of the Manganese Zone mineralisation is associated with the contact between this diorite and the surrounding marble. Mineralisation ranges from surface to approximately 400 m depth.

Free gold mineralisation occurs in the marble with minimal associated sulfides. Disseminated quartz–sulfide mineralisation occurs in clay-altered and brecciated diorites as well as locally carbonate-altered diorite. Moderate volumes of massive sulfide pyrite mineralisation occur within the Manganese Zone. It appears that ‘leachable’ mineralisation is a combination of free gold in marble and supergene oxidised mineralisation in both marble and diorite. Leachable oxide mineralisation occurs to +200 m below surface.

Marble Zone Mineralisation

The Marble Zone occurs in the south-eastern portion of the Çöpler deposit and is associated with a north-east striking fault contact between marble to the east and metasediment and intrusions to the west. The geology in this area is typified by large ‘plugs’ of gossan and diorite that have formed at the junctions between large-scale faults, where mineralising fluid flow has been considerable. The width of the Marble Zone is approximately 350 m, and the strike length is 300 m east–north-east. The depth of mineralisation ranges from surface to approximately 160 m below surface.

Mineralisation occurs as both disseminated sulfides in veinlets and massive sulfide along the marble contact. Oxidation has occurred along the north-east structure resulting in greater depths of oxidised mineralisation than that seen in the Main Zone.

West Zone Mineralisation

The West Zone occupies the westernmost portion of the Çöpler deposit and is located at the contact between the basement metasediment and the overlying limestone / marble, where a large-scale north-east trending fault is located. Mineralisation is present within veinlets containing disseminated sulfides, massive sulfides, and oxidised gossan. The West Zone has a strike length of approximately 700 m north-east and is approximately 150 m wide. Multiple narrow mineralised zones are present sub-parallel to the faulted contact and occur to a depth of approximately 150 m below surface.

6.1.3 Structure – Çöpler Deposit

The Çöpler deposit area demonstrates trans-tensional deformation. The extensional deformation in the area dominates over strike-slip motion as indicated by the lack of compressional structures and the presence of normal movement on all faults. Structurally, the Çöpler deposit occurs in a horst-like feature developed within a sinistral trans-tensional strike slip setting (Figure 6.7). The two boundary faults delimit the northern and southern extent of the gossan-like, oxidised, supergene, gold-bearing deposits. The northern and southern boundary faults are located almost at the present boundaries of the mine, and they dip away from the mine, thereby defining the horst geometry. In addition, the deposit is traversed by several cross-cutting normal faults (with or without strike-slip components) in various orientations that complicate but localise the geometry and position of oxidised ore (Kaymakçı, 2017, internal company reporting).

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Figure 6.7 Simplified Schematic of the Çöpler Deposit Structures (cross-section)

Anagold, 2020

6.2 Geological Setting – Çakmaktepe Deposit

6.2.1 Geology – Çakmaktepe Deposit

The Çakmaktepe deposit is made up of several mineralised zones (Figure 6.8). The deposit area mainly comprises Palaeozoic metamorphic rocks and marble belonging to the Keban Formation and Mesozoic platform carbonate such as the Munzur Formation limestone. All these units are tectonically overlain by ophiolitic mélange rocks. These ophiolitic rocks originated from the northern branch of the Neotethys ocean, the former position of which is delineated by the Ankara–Erzincan suture zone. The emplacement of the ophiolitic units took place at the end of the Upper Cretaceous with north to south motion.

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Figure 6.8 Geological Map of the Çakmaktepe and Ardich Deposits

Anagold, 2020

The youngest units include Eocene and younger magmatic rocks, volcaniclastics rocks and sedimentary units that unconformably overlie and seal the Munzur Formation limestone, its basement and the ophiolitic units. All these units are intruded by intermediate igneous rocks that are exposed mainly at the northern and western parts of the Munzur mountains and southern margin of the Sivas Basin.

Listwanite formed in structurally deformed areas by the percolation of CO2-rich fluids along the margins of ultramafic rocks within the ophiolite complex. Sulfidic jasperoid is present, a result of silica-sulfide metasomatism of Munzur Formation carbonate rocks. Both listwanite and jasperoid are important host rocks for gold and silver mineralisation.

6.2.2 Mineralisation – Çakmaktepe Deposit

The Çakmaktepe deposit is a structurally controlled gold–silver–copper deposit, displaying both epithermal and replacement mineralisation styles. Mineralisation is primarily associated with jasperoid and listwanite. At depth, mineralisation transitions below the base of oxidation to disseminated pyrite, vein sulfides, and massive sulfide horizons, generally occurring within shear zones, along shallow thrusts, in diorite sills, and on intrusion margins.

As with the Çöpler deposit, Çakmaktepe is thought to be the result of intrusive activity that generated suitable conditions for mineralisation of ophiolite, limestone, and hornfels lithologies (Figure 6.9). A complex system of faults enabled emplacement of diorite intrusions and transport of metalliferous fluids associated with the mineralising system. Steeply dipping, shear-hosted mineralisation characterises the deposits at Çakmaktepe North, whereas more shallowly dipping thrust-related mineralisation is characterised at Çakmaktepe East, Çakmaktepe South-East and Çakmaktepe Central. Key to each structurally associated style of mineralisation is the juxtaposition of ophiolites against limestone and hornfels.

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Figure 6.9 Geological Mapping within Çakmaktepe Central Pit

Anagold, 2020

The Çakmaktepe North area is strongly sheared with epithermal characteristics and grade associations with intrusive diorite dykes. The bulk of the mineralisation is structurally confined to a major sub-vertical shear zone (Main Shear). The Main Shear varies in width from 5–40 m, has been defined to a depth of 200–250 m below surface, and dips at approximately 70° to the east. Surface mapping and sampling have defined the mineralised extent of the shear as being over 1 km in length.

Mineralisation at Çakmaktepe North is not solely contained within the shear zone, but also occurs along flat thrust structures and lithological contacts cut by the shear zone. Contacts between ophiolite and limestone, limestone and hornfels, and all lithologies in contact with intrusive diorite sills and dykes are generally mineralised. The listwanite horizon is the most favourable host rock for gold mineralisation. Diorite intrusions show evidence of hydrothermal activity that either takes the form of massive iron-dominated replacement (magnetite, specular hematite, or pyrite) or sheeted crystalline quartz vein bearing jasperoid closer to diorite contacts.

Other mineralised zones within the Çakmaktepe deposit are referred to as ‘contact’ styles of mineralisation where iron, sulfur, gold, copper, and silver have been emplaced along thrust surfaces where ophiolite is next to limestone and metasediment. Epithermal veining and replacement alteration and textures are prevalent. Skarn and metasomatic mineralisation occur in contact with intrusive diorite dykes, sills, and stocks.

Oxide mineralisation at Çakmaktepe is predominantly characterised by silica–iron–carbonate-rich jasperoid, less-siliceous iron-rich gossan, and epithermal veined and brecciated limestone.

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6.3 Geological Setting – Ardich Deposit

6.3.1 Geology – Ardich Deposit

The Ardich deposit is located immediately to the north-west of the Çakmaktepe deposit (Figure 6.8). The north-western portion of Ardich and the Çakmaktepe North mineralised zone are near each other, as are the Ardich South-east and Çakmaktepe East mineralised zones. While there are some characteristic differences between Ardich and Çakmaktepe, the local geology is generally very similar.

The mineralisation at Ardich occurs at a higher stratigraphic level that that seen at Çakmaktepe. The emphasis at Ardich is on the ophiolitic mélange rocks that have been thrust into place on top of the basement metasediment and carbonate lithologies.

The local geology at Ardich is dominated by ophiolite, listwanite, and dolomite and limestone, with mineralisation occurring along low-angle thrust zones between ophiolite, listwanite, and dolomite and limestone (Figure 6.10). This occurs within a complex north-west trending structural zone that is cut by multiple high-angle faults that together result in multiple rotated fault blocks and mineralised zones.

The mineralisation at Ardich is considered to be related to fluids associated with diorite intrusions at depth, much like those observed at the Çöpler and Çakmaktepe deposits. Diorite dykes are present but not common at Ardich, unlike the adjacent Çakmaktepe deposit and nearby Çöpler deposit where diorite is a dominant lithology.

Figure 6.10 Schematic of Ardich Geological Setting with Mineralisation Examples

Anagold, 2020

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6.3.2 Mineralisation – Ardich Deposit

The mineralisation at Ardich is related to crystalline and chalcedonic quartz veins within the brecciated and silicified listwanite and dolomite zones. The mineralisation is predominantly in the form of oxide, with sulfide mineralisation confined to limited pyrite-rich jasperoid zones. Clay / gossan in jasperoid or limestone karstic boundaries also contain high-grade gold across Ardich.

Gold grades increase at dolomite / listwanite contacts and within silica-rich listwanite that acts as horizontal traps for higher grade gold-bearing mineralisation. Increases in gold grade can be seen along the lithological contacts. Elevated grades can exist within either contact lithology. Several drillholes show a very rapid down-hole change in gold grade from mineralised to unmineralised material, indicating that mineralisation is tightly constrained instead of disseminated across the deposit. Due to these relationships, the three-dimensional model indicates that the main mineralised zone is tabular and sub-horizontal.

6.4 Geological Setting – Bayramdere Deposit

6.4.1 Geology – Bayramdere Deposit

The Bayramdere deposit is an oxide gold and copper deposit with similar geological and mineralisation characteristics to the Çakmaktepe and Ardich deposits. The geology is characterised by ophiolite thrust over the limestone and dolomite, which are in turn intruded by granodioritic stocks. Gossans are generally observed as lenses and confined by normal faults.

The Bayramdere deposit is structurally controlled, displaying a replacement gold (minor copper, minor silver) mineralisation style. The deposit is dominantly represented by near-surface oxide mineralisation, primarily associated with iron-rich gossan.

The Bayramdere deposit is thought to be the result of intrusive activity that generated suitable conditions for mineralisation. A complex system of faults enabled emplacement of diorite intrusions and transport of metalliferous fluids associated with the mineralising system. Key to each structurally associated style of mineralisation is the juxtaposition of ophiolite against limestone (±hornfels) to create the right geochemical conditions for the deposition of gold and other metals.

6.4.2 Mineralisation – Bayramdere Deposit

The Bayramdere mineralisation is localised within three stacked, shallow-dipping zones that formed at the contact of limestone and ophiolite, with mineralisation having replaced limestone along the contacts. The limestone / ophiolite contacts are low-angle thrusts, typified by limestone wedges within a dominantly ophiolite stratigraphy. Mineralisation occurs within shallow iron-rich gossan horizons.

6.5 Geological Setting – Regional Prospects and Targets

Since 2000, Anagold exploration programmes within the Çöpler district have identified several new gold-dominant and copper–gold prospects. The gold-dominant regional prospects include the Çöpler Saddle and Elmadere. Copper–gold prospects are Aslantepe, Sarıdere, Findiklidere and Mavidere porphyries located within the Mavialtin Porphyry Belt (Figure 6.11) and the early exploration stage Meşeburnu porphyry located west of the Çöpler deposit.

Each of these prospects is discussed below.

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Figure 6.11 Çöpler District Exploration Projects

Anagold, 2022

6.5.1 Geology – Çöpler Saddle

The Çöpler Saddle prospect borders the western flank of the Çöpler mine. The Çöpler Saddle is associated with a shear zone defined as an arc-like structure that trends north–south for approximately 2 km, Figure 6.11. Along the shear zone, the geology is dominated by limestone, marble, and hornfels units that are in turn intruded by small-scale microdioritic to granodioritic stocks. These lithologies were subjected to silica-clay alteration with iron oxide developments along the local structures as well as clay-pyrite alteration. At the south of the zone, silica is mainly observed as jasperoid lenses, of approximately 2 m long and 1 m wide, which occur along the hornfels and marble contacts. At the centre of the zone, less silica is observed and larger gossan-like mineralised iron oxide bodies have formed.

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6.5.2 Geology – Meşeburnu and Elmadere

The Meşeburnu and Elmadere prospects (former Demirmağara project licence group) are located approximately 7 km south-west of the Çöpler deposit (Figure 6.11). The area is covered by ophiolites, limestone, and metamorphic rocks that are intruded by dioritic to granodioritic stocks. Three types of mineralisation have been identified in the area:

•Gold-bearing skarn and jasperoid occurrences along limestone and granodiorite contacts.

•Epithermal gold mineralisation developed along ophiolite, listwanite, and limestone structural contacts (referred to as Elmadere mineralisation).

•Meşeburnu copper–gold porphyry mineralisation.

Gold-bearing skarn and jasperoid occurrences were tested with drilling between 2001–2017, however only short gold-mineralised intervals were intersected. Mapping and sampling in Elmadere and Meşeburnu prospects are ongoing to define drilling targets.

6.5.3 Geology – Mavialtin Porphyry Belt Prospects

The Mavialtin Porphyry Belt is a structural corridor approximately 6–7 km wide and extending over approximately 20 km from the Çakmaktepe deposit in the north to the Mavidere porphyry deposit in the south (Figure 6.11). The Mavialtin Porphyry Belt contains the Mavidere, Findiklidere, Saridere, and Aslantepe porphyry copper–gold prospects.

6.5.3.1 Geology – Mavidere

The Mavidere porphyry copper–gold mineralisation is hosted by hornblende–biotite monzonite to monzogranite to granodioritic phases of a shallow porphyritic intrusive hosted by metamorphic and crystallised limestone. At the centre of the porphyry system, the intrusive phases were subjected to mainly potassic alteration with clay and minor sericite overprinting covering an area of 800 m x 400 m. The porphyry system appears to continue underneath the moraine cover to the east and south.

Previous exploration activities included:

•surface mapping,

•geochemistry (soil, rock, stream sediment sampling),

•geophysical studies (induced polarisation (IP) and surface magnetics), and

•Reverse circulation (RC) and diamond core (DD) drilling.

The prospect was first drilled in 2001, with 1,780 m at eight locations. In 2008, 22 additional holes were drilled totalling 7,761 m, with the preliminary results announced in 2009. From 2011 through 2013, 77 DD holes totalling 20,653.3 m and 68 RC holes totalling 7,512 m were completed. Field studies and mapping in 2018 identified additional mineralised zones, some of which were drill tested in 2018 and 2019. Drillhole MD06, drilled in 2019, returned a highly prospective intercept of 269.1 m at 0.34% Cu and 0.55 g/t Au from surface.

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6.5.3.2 Geology – Aslantepe

The geology of the Aslantepe porphyry copper–gold prospect is dominated by ophiolites thrusted over Jurassic to Cretaceous limestone, both of which are intruded by dioritic to granodioritic stocks and dykes. The Aslantepe intrusives outcrop in a narrow corridor subjected to propylitic, potassic, and clay alteration. The potassic zone is characterised by well-developed intense quartz–sulfide stockwork veinlets with secondary biotite, K-feldspar, and magnetite. In 2018, two additional DD holes were drilled at Aslantepe, with drillhole AT07 intersecting 63.9 m at 0.22% Cu and 0.45 g/t Au from 46.7 m down-hole. The mineralisation appears to be dipping underneath the ophiolites.

6.5.3.3 Geology – Sarıdere

The Sarıdere porphyry copper–gold prospect is covered by metamorphic limestone and ophiolite, which are in turn intruded by tonalitic to granodioritic stocks. The prospect was initially identified by stream sediment and soil anomalies. In 2018 and 2019, exploration activities identified potassic-altered porphyry intrusive outcrops covering an area of approximately 800 m x 500 m, with a phyllic alteration halo around the potassic zone of 4.3 km x 0.6 km. Seven DD holes totalling 1,461.5 m were drilled from 2007 through 2013 at the margin of the porphyry system, testing the elevated soil geochemistry. These holes intersected short intervals of copper–gold mineralisation.

6.5.3.4 Geology – Fındıklıdere

The Findiklidere porphyry copper–gold prospect is covered by massive Jurassic to Cretaceous limestone, which has been over-thrusted by ophiolites on the eastern flank. These units were intruded by fine to medium-grained tonalitic to granodioritic intrusive stocks. The porphyry copper mineralisation is characterised by well-developed stockwork quartz–magnetite–pyrite veins with copper. Peripheral iron–copper–gold skarns are observed within the limestone. In 2018, the geology, structure, and alteration were re-mapped to better understand the porphyry potential of the prospect. Results of this field work indicated that the porphyry mineralisation was potentially continuing underneath the ophiolitic body to the south-west of the known porphyry mineralisation. In 2019, DD hole FD02 was drilled to test porphyry potential beneath the ophiolitic cover. The hole was mineralised over 234.4 m (down-hole) with some higher grade intervals such as 32.1 m at 0.84% Cu and 0.37 g/t Au from 13.4 m and 16.5 m at 1.27% Cu and 0.07 g/t Au from 139.5 m.

The abovementioned drilling results were announced within the exploration press release dated 14 February 2020.

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6.6 Deposit Types

Porphyry copper–gold systems host some of the most widely distributed mineralisation types at convergent plate boundaries, including porphyry deposits centred on intrusions; skarn, carbonate-replacement, and sediment hosted gold deposits in increasingly peripheral locations; and high to intermediate-sulfidation epithermal deposits.

The alteration and mineralisation in porphyry copper–gold systems are zoned outward from the stocks or dyke swarms, which typically comprise several generations of intermediate to felsic porphyry intrusions. Porphyry copper (± gold, ± molybdenum) deposits are centred on the intrusions, whereas carbonate wall rocks commonly host proximal copper–gold skarns, less common distal zinc–lead and/or gold skarns, and, beyond the skarn front, carbonate-replacement copper and/or zinc–lead–silver (± gold) deposits, and/or sediment-hosted (distal-disseminated) gold deposits. Peripheral mineralisation is less conspicuous in non-carbonate wall rocks but may include base metal-bearing or gold-bearing veins and mantos (Sillitoe, 2010). Skarn deposits are typically hosted in mineralogically simple fine-grained clastic and carbonate sedimentary rocks. Skarn mineralogy and metal content is largely dependent on the crystallisation history and genesis of associated plutons (Meinert et al., 2005).

The Çöpler district is located at the edge of a convergent plate boundary. It is characterised by a complex structural history and is associated with intermediate intrusive and carbonate-rich host lithologies. As such, porphyry copper–gold systems and related styles of mineralisation are appropriate models to be applied across the Çöpler district.

The Çöpler deposit consists of three major mineralisation types that are closely associated with each other: low-grade sub-economic porphyry copper–gold–molybdenum mineralisation characterised by well-developed alteration zones and stockwork quartz veins (Main Zone); intermediate sulfidation epithermal mineralisation observed in the Manganese Zone as clusters of bright pink, banded, colloform rhodochrosite base metal sulfide veins and breccia lodes; and iron–gold (± copper) skarn with related carbonate replacement gold mineralisation.

The setting, alteration mineralogy, and mineralisation characteristics of the Manganese Zone are somewhat consistent with an intermediate sulfidation epithermal system, as defined in Hedenquist et al., (2000).

Exploration programmes modelled on epithermal-style deposits have shown success in the Çöpler district. A multi-phase porphyry model with a barren trapping system and a possible mineralised porphyry underneath it is also applicable.

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7 EXPLORATION

This section has not been changed from the 2021MR and remains the most current study work available.

7.1 Exploration – Çöpler Deposit

Exploration of the Çöpler deposit has been conducted by Anagold and its predecessors since September 1998. Work completed has included:

•Geological and reconnaissance mapping.

•Rock chip, grab, soil, channel, and stream sediment geochemical sampling.

•Ground geophysical surveys including ground magnetic, complex resistivity / IP, time domain IP and controlled source audio-frequency magneto-telluric (CSAMT) surveys.

•A regional helicopter-borne geophysical survey.

•Reverse circulation (RC) and diamond core (DD) drilling programmes.

•Acquisition of satellite imagery.

•Mining technical studies.

•Geotechnical and hydrogeological studies.

•Environmental baseline studies.

•Studies in support of project permitting.

•Metallurgical testwork and studies.

•Condemnation evaluations.

The principal exploration technique used at Çöpler has been RC and DD drilling, conducted in multiple campaigns since 2000. Initially, exploration was directed at evaluating the economic potential of the near-surface oxide mineralisation for the recovery of gold by either heap leaching or conventional milling techniques.

In 2013, drilling occurred primarily in the western portion of the Main Zone and on the northern edge of the Çöpler deposit. Drilling during 2014 focused on verification of existing drilling results through a twin-hole programme. Drilling in 2015 provided data coverage at depth in the Manganese Zone, infill drilling in the Main Zone, and testing of low-sulfur mineralisation below the oxidation boundary.

Drilling continues to better define both the oxide and sulfide portions of the Çöpler deposit.

7.1.1 Geological Mapping – Çöpler Deposit

Surface mapping and sampling has been undertaken over the life of the Project, culminating in a detailed geological map of the Çöpler valley, shown in Figure 7.2.

Geological mapping is used in support of exploration vectoring, exploration activities, infrastructure locations, mine planning, and environmental monitoring. One of the aims of the mapping studies is to provide sufficient information to define mineralisation types and structural settings for the Çöpler deposits. Alteration zones, such as the high-temperature porphyry alteration preserved in the southern wall of the Main Zone (shown in Figure 7.1), were identified through detailed bench wall mapping during the target generation programmes.

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Figure 7.1 Çöpler Deposit Map of Alteration Minerals

Anagold, 2020

7.1.2 Geochemical Sampling – Çöpler Deposit

Extensive sampling programmes have been, and continue to be, conducted within the Çöpler area, leading to the identification of significant gold anomalies including the near-mine discovery of the Çöpler Saddle on the western flank of the Çöpler mine.

7.1.3 Geophysics – Çöpler Deposit

Various ground and airborne geophysical surveys have been conducted at the Çöpler deposit as well as across the wider Çöpler district since mid-2000. Surveys carried out include ground magnetic, complex resistivity / induced polarisation (IP), time domain IP, and CSAMT surveys, as well as a regional helicopter-borne aeromagnetic survey that included the broader Çöpler district.

Physical property measurements are collected regularly on outcrops and DD core samples, including magnetic susceptibility, resistivity, and chargeability. Additionally, four samples from DD hole CDD067 were sent to Systems Exploration in Australia for a detailed physical property analysis.

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7.2 Exploration – Çakmaktepe Deposit

The Çakmaktepe deposit and surrounding mineralised zones were identified by stream sediment samples with elevated gold geochemistry.

Drilling at Çakmaktepe started in 2012. The recent drilling (2019 onwards) has been designed to improve the known Mineral Resources identified at Çakmaktepe North. Data collected to date includes magnetic geophysical surveys, outcrop and bench wall mapping, rock and soil sampling, and both RC and DD drilling.

7.2.1 Geological Mapping – Çakmaktepe Deposit

The first geological mapping study in the area was conducted in 2000.

Mapping in 2014–2016 focused on deposit-wide surface geology definition at a scale of 1:1,000, reducing to 1:500 scale for the Çakmaktepe Mineral Resource area. The establishment of a network of drill tracks and pads on the sides of hills and ridges resulted in new rock exposures that have been subjected to detail geological mapping. Mapping included the collection of lithological, alteration, geochemical, and structural data.

An additional mapping study within the Çakmaktepe deposit was initiated as the Çakmaktepe operation advanced in late-2018. Details from the bench walls were collected and integrated into the drillhole dataset (mapping example shown in Figure 7.2). This has resulted in a more-accurate geological model for further pit extension exploration drilling.

Figure 7.2 Çakmaktepe Deposit – Example East Pit Geological and Structural Map

Anagold, 2020

7.2.2 Geochemical Sampling – Çakmaktepe Deposit

Geochemical sampling programmes at Çakmaktepe were initiated in 2014 and included rock chip and soil sampling (Table 7.1). Geochemical sampling was also used to define areas of alteration and mineralisation that inform additional detailed sampling surveys.

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Table 7.1 Number of Geochemical Samples within the Çakmaktepe Deposit

Year	Rock Chip Samples	Soil Samples
2014	661	341
2015	3,527	–
2016	356	270
2017	63	1,638
2019	540	–
2020	13	–
Total	5,160	2,249

A total of 5,160 rock chip samples have been collected from the Çakmaktepe deposit since 2014. During 2019, rock chip sampling extended into bench wall and haulage roadsides to help define the extents of the deposit more accurately.

Soil sampling programmes were initiated during the 2010 exploration programme. The deposit has been fully covered with a 50 m x 50 m sampling grid totalling 2,249 samples.

Stream sediment sampling was carried out on a regional scale as part of target generation programmes since 2002. A total of 851 sediment samples have been collected.

7.3 Exploration – Ardich Deposit

Exploration activities across the Ardich deposit began in 2017 and included geological mapping, geochemical sampling, and DD drilling programmes.

7.3.1 Geological Mapping – Ardich Deposit

The Ardich deposit was discovered in 2017 during detailed geological mapping and rock sampling programmes. Results of the mapping study highlighted the potential of the Ardich deposit and its extension to the south. The mineralisation identified to date continues approximately 4 km on a north westerly trend.

7.3.2 Geochemical Sampling – Ardich Deposit

Geochemical sampling programmes at Ardich have included rock chip / channel and soil sampling, (Table 7.2). Most of the geochemical sampling campaigns across the Ardich deposit were designed based on findings from the geological mapping programmes.

Table 7.2 Number of Geochemical Samples within the Ardich Deposit

Year	Rock Chip/Channel Samples	Soil Samples
2017	175	125
2018	912	–
2019	880	1,718
2020	140	–
Total	2,107	1,843

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A total of 2,107 rock chip / channel samples have been collected since 2017 from outcrops across the Ardich deposit. Rock chip / channel sampling has been the most representative surface sampling, collected directly from altered rock exposures. As the drilling programmes continue, newly opened drill tracks and pads give good access to new rock exposures that are subjected to rock sampling and geological mapping.

Soil sampling was completed in early-2000 as part of a regional geochemical reconnaissance programme, with early targets being potentially mineralised listwanite-capped faults. Anagold started regional systematic soil sampling on 200 m x 200 m grids to cover all tenements in 2011. At the Ardich deposit, a total of 1,843 soil samples were collected on a sampling grid of 50 m x 50 m, which was reduced to 25 m x 25 m in gold-anomalous areas in 2017–2019.

7.4 Drilling

All drillhole counts in this section include holes drilled for resource definition, geotechnical, and metallurgical purposes.

7.4.1 Drilling – Çöpler Deposit

The Çöpler deposit continues to be tested by reverse circulation (RC) and diamond core (DD) drilling. The details of drillholes utilised in this Mineral Resource update for the Çöpler deposit are presented in Table 7.3. Typically, the drillhole spacing at surface is a nominal 50 m, however, in some areas the drill spacing has been reduced to 25 m (Figure 7.3).

Step-out drilling at the Çöpler deposit has defined most of the lateral boundaries of the mineralisation. There has been additional development drilling, as well as condemnation drilling of areas planned for infrastructure during the last few years. In order to improve confidence in the short-range mine planning, infill drilling programmes have been conducted since 2007.

Drilling in 2014 focused on confirmation of the mineralisation with a twin-hole programme.

Development drilling continued in 2015 by improving sample coverage at depth in the Manganese Zone and along structural boundaries in the Main Zone. In addition to the drilling of in situ mineralisation, a stockpile drilling programme began in December 2015 to confirm sulfide stockpile ore grade, grade distribution, and mineralogy.

Drilling at Çöpler between 2016 and 2020 mainly concentrated on target generation to increase the amount of oxide material for the production portfolio. This was focused on the Main Zone, West Zone, and the Çöpler Saddle areas. More specifically, the programme aimed to test continuation of the main gold-bearing structures based on a re-interpretation of the Çöpler structural and mineralisation settings. In-pit drilling campaigns continue with extensive exploration programmes to define additional oxide gold potential.

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Figure 7.3 Drillhole Collar Location Plan – Çöpler Deposit

Anagold, 2022

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Table 7.3 Drilling History – Çöpler Deposit

Year		Hole Type		Number of Holes		Metres Drilled	Total Metres / Year
2000		DD		4		971.5	971.5
2001		DD		10		2,254.4	6,320.3
	RC		32		4,065.9
2002		DD		31		6,575.6	6,835.6
	RC		1		120.0
	Other		2		140.0
2003		DD		33		2,975.7	2,975.7
2004		DD		37		4,413.5	16,634.8
	RC		228		11,036.0
	Other		16		1,185.3
2005		DD		24		4,776.4	35,062.1
	RC		177		29,009.7
	Other		16		1,276.0
2006		DD		17		2,102.6	15,857.6
	RC		94		12,878.0
	Other		24		877.0
2007		DD		74		16,513.2	34,435.9
	RC		125		16,998.5
	Other		40		924.2
2008		DD		35		5,059.4	9,963.4
	RC		41		4,904.0
2009		DD		23		5,789.5	10,135.5
	RC		34		4,346.0
2010		DD		14		1,916.1	2,060.6
	RC		1		144.5
2011		DD		115		29,359.0	47,342.0
	RC		150		17,983.0
2012		DD		145		50,156.5	64,041.0
	RC		120		13,884.5
2013		DD		126		33,040.9	37,585.9
	RC		53		4,545.0
2014		DD		12		1,296.5	1,296.5
2015		DD		59		6,214.1	12,778.1
	RC		69		6,564.0
2016		DD		148		3,826.5	6,020.5
	RC		94		2,194.0
2017		DD		41		3,370.5	3,370.5
2018		DD		109		10,745.0	10,745.0
2019		DD		62		7,607.7	7,607.7
2020		DD		131		23,029.90	23,029.90
2021		DD		68		18,491.80	18,491.80
Total		RC		1,219		128,673.1
	DD		1,318		240,486.3
	Other		98		4,402.5
	All Types		2,635		373,561.9

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7.4.2 Drilling – Çakmaktepe Deposit

A total of 1,183 drillholes have been drilled at the Çakmaktepe deposit since 2012. This included 528 RC holes, 570 DD holes, and the remainder a mixture of RC and DD. As production proceeded within the Çakmaktepe Central and Çakmaktepe East pits, additional targets were generated to provide pushback options within the pit design. A total of 136 DD holes have been completed since 2019 to test for continuation of the Çakmaktepe deposit, Figure 7.4 and Table 7.4.

Table 7.4 Drilling History – Çakmaktepe Deposit

Year	Number of Drillholes	Drilled Metres
2012	21	2,287.5
2013	7	962.0
2014	162	15,976.7
2015	256	21,463.2
2016	485	64,108.6
2017	116	9,366.2
2019	75	5,919.4
2020	61	8,702.3
Total	1,183	128,785.9

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Figure 7.4 Drillhole Collar Location Plan – Ardich, Çakmaktepe, and Bayramdere Deposits

Anagold, 2022

7.4.3 Drilling – Ardich Deposit

A total of 531 DD holes have been drilled at the Ardich deposit since late-2017, Figure 7.5 and Table 7.5. After the initial discovery of the Ardich deposit, DD drilling programmes have continued to better-define the mineralisation and to improve the Mineral Resource estimates. Drilling to obtain samples for metallurgical testing and hydrogeological studies has also been undertaken at Ardich.

A total of 233 drillholes were included in the previously-announced Ardich Mineral Resource (CDMP20TR, drillholes AR1–AR233). Since the data cut-off date for the CDMP20TR Mineral Resource update, data has been obtained for an additional 129 drillholes (AR233–AR531).

A drillhole collar plot is shown in Figure 7.5, indicating the various generations of drilling.

The target of the post-2020 drilling has been two-fold:

•Infill drilling within the bounds of the 2020 resource model area.

•Step-out drilling to the west, south, and south-west of the 2020 resource area.

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Table 7.5 Drilling History – Ardich Deposit

Year	Number of Drillholes	Drilled Metres
2017	9	1,374.10
2018	91	14,216.40
2019	133	27,821.20
2020	147	35,146.65
2021	151	32,586.00
Total	531	111,004.35

Figure 7.5 Drillhole Collar Location Plan – Ardich

Anagold, 2022

Drillholes AR1 through AR427 have contributed to updated (2021) resource modelling for Ardich, which is discussed in Section 11.3. The 2021 update resulted not only in a larger inventory than that previously announced but is also a higher confidence inventory. The data cut-off date for updated Ardich resource model was 31 May 2021.

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7.4.4 Drilling – Mavialtin Porphyry Belt Prospects

Drilling within the Mavialtin Porphyry Belt first started in early-2000. Re-interpretation of historical drillholes and detailed mapping programmes resulted in the definition of new drill targets in subsequent years. A total of 353 holes have been completed between 2001–2020 at various targets within the Mavialtin Porphyry Belt, Figure 7.6 and Table 7.6.

7.4.5 Grid Coordinate Systems

The Çöpler project uses the European 1950 (E1950) datum coordinate system – this is a Turkish Government requirement.

The Çöpler project is in UTM6 zone 37N of the E1950 coordinate system. Until 2014, drillhole collars were surveyed by the mine surveyors in the E1950 UTM3 coordinate system and then converted to E1950 UTM6 before making them available to other personnel. The conversion from UTM3 to UTM6 was achieved by subtracting 1,746 m (–1,746 m) from the UTM3 northing coordinate and adding 17 m (+17 m) to the UTM3 easting coordinate. There is no rotation, scaling, or change in elevation between the E1950 UTM3 and E1950 UTM6 systems. Since March 2014, collar coordinates have been and are being collected in the ED1950 UTM6 coordinate system.

7.4.6 Collar and Down-hole Surveys

Up until 2014, drillhole collars were surveyed by Anagold surveyors using a Topcon differential global positioning system (DGPS) instrument. Approximately 4% of the drillholes up to 2014 have planned collar locations, rather than surveyed collar data. After 2014, the exploration department managed the collection of collar survey coordinates with the use of a differential GPS (DGPS). All collar survey data is checked prior to being stored within the corporate drillhole database.

Down-hole surveys are collected for all drillholes. Prior to 2009, down-hole surveys were undertaken using a Reflex Instruments Limited (Reflex) single shot down-hole camera. In 2009, a Reflex multi-shot down-hole camera was introduced on the Project. Drilling contractors upgraded to a Reflex EZ Trac tool for down-hole survey data collection through to the end of 2017, thereafter, to date the majority of the drillholes have been down-hole surveyed using Reflex S Process V2.5.0650 and Devico PeeWee. Survey measurements were taken every 10 m down-hole, and data provided with raw files to record quality assurance and quality control (QA/QC) for each survey.

The depth of the surveys varies between drillholes and is dependent on the depth and angle of the drillhole.

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Figure 7.6 Drillhole Collar Location Plan – Mavialtin Porphyry Belt Prospects

Anagold, 2022

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Table 7.6 Drilling History – Mavialtin Porphyry Belt Prospects

Project		Year		Number of Drillholes		Drilled Metres
Aslantepe		2014		15		2,278.7
	2018		2		440.3
	2020		1		400.8
Aslantepe Total				18		3,119.8
Bayramdere		2007		4		763.5
	2013		28		4,024.0
	2014		68		4,698.3
	2015		17		669.9
	2016		1		98.0
	2020		2		480.5
Bayramdere Total				120		10,734.2
Fındıklıdere		2008		4		1,085.3
	2012		15		5,132.0
	2013		4		1,091.2
	2014		3		825.5
	2019		5		2,501.5
	2020		5		2,121.8
Fındıklıdere Total				36		12,757.3
Sarıdere		2007		6		1,160.5
	2013		1		301.0
	2020		3		1,384.0
Sarıdere Total				3		2,845.5
Mavidere		2001		8		1,780.3
	2008		22		7,761.1
	2011		22		3,806.2
	2012		37		10,479.5
	2013		78		11,171.6
	2018		5		2,119.8
	2019		4		1,567.1
Mavidere Total				176		38,685.6

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8 SAMPLE PREPARATION, ANALYSES, AND SECURITY

This section has not been changed from the 2021MR and remains the most current study work available.

From 2004 to late-2012, samples were prepared at ALS İzmir, Turkey (ALS İzmir) and analysed at ALS Vancouver, Canada (ALS Vancouver), (collectively ALS Global). From late-2012 through 2014, samples were prepared and analysed at ALS İzmir. Samples in 2015 were prepared and analysed at the SGS laboratory in Ankara, Turkey (SGS). From 2015 to current, ALS İzmir is being used as the main laboratory and samples are being prepared and analysed there. Umpire analysis was completed by ACME Mineral Laboratories (ACME) in Ankara, Turkey.

SGS is certified to ISO 9001:2008 and OHSAS 18001, ALS İzmir has ISO-9001:2008 certification, and ALS Vancouver is ISO/IEC 17025:2005 accredited for precious and base metal assay methods. ACME is part of the Bureau Veritas (BV) group, globally certified to ISO 9001:2008.

ALS Global and SGS are specialist analytical testing service companies, both independent of SSR.

Samples from the 2000–2003 drilling programme were submitted to OMAC Laboratories Limited (OMAC) in Loughrea, Ireland. ALS Global assumed ownership of OMAC in 2011.

Detailed sampling and quality assurance and quality control (QA/QC) procedures for reverse circulation (RC) and diamond core (DD) drilling were instigated and have been in use since the first drill programme. The QA/QC procedures have been retained by Anagold, although the insertion rates have been modified for some of the later programmes.

Anagold operates an on-site laboratory for assay of production samples. The on-site laboratory is certified to ISO 17025:2017 but is not independent. It is primarily used for the analysis of grade control samples.

8.1 Sample Collection

8.1.1 Reverse Circulation Drilling Sample Collection

Historically, RC drilling was completed with a 4.5–4.75 inch (11.4–12.0 cm) diameter down-the-hole hammer drill rig. RC cuttings were passed through a cyclone with a 10 inch (25.4 cm) port for sample collection. RC drill intervals were 1 m in length and cuttings for the entire 1 m sample interval were collected from the cyclone underflow in large, reinforced plastic bags. Prior to 2015, RC samples were split using a Jones splitter.

Since 2015, RC drilling has been completed with a nominal 5.25 inch face sampling hammer with centre sample return to a rig-side mounted sampling system. The sampling system included of a cyclone, sending 1 m samples to a rotary cone splitter. The rotary cone sample splitter was adjusted to maintain a representative sample volume. RC chip samples, to a weight of 3–5 kg, were collected in calico bags for analysis. All sample bags are clearly numbered and labelled with the drillhole name and sample number. Residual samples were collected in PVC bags and stored in a bag farm for six months in case re-logging, duplicate sampling, metallurgical sampling, or follow-up QA/QC was required.

The rig sampler sieves a small portion of the residual sample from the large plastic bag and places the sieved portion in a plastic chip tray to provide a sample for logging and as an enduring geological record. The plastic chip trays are photographed.

RC drilling is generally only used above the water table. The water table is closer to the surface in the northern region of the Main Zone, and for that reason, diamond drilling is the preferred method in this zone.

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The following QA/QC samples are collected during the RC sampling process:

•Certified Reference Materials (CRMs) are inserted into each sample batch at a rate of two CRMs in every 40 samples (1-in-20 insertion rate).

•Prior to 2015, blank samples were inserted into each batch at a rate of one blank in every 60 samples (1-in-60 insertion rate). Since 2015, this has been changed to a 1-in-30 insertion rate.

•Field duplicate samples are collected by splitting an RC sample twice to collect two independently numbered samples of the same interval or selecting a quarter of the remnant core. Historically, field duplicates were collected and inserted into the sample job at a rate of 1-in-40 samples. In 2015, field duplicate insertion rates were increased to 1-in-20.

8.1.2 Diamond Drilling Sample Collection

Up until 2017, the diamond drilling undertaken on the Project has generally been HQ or NQ diameter. HQ core has a nominal diameter of 63.5 mm while NQ has a nominal size of 47.6 mm. Approximately 90% of the DD core drilled at Çöpler and Çakmaktepe is HQ. Some drillholes are started with HQ and then reduced in size to NQ further down the hole.

Of the more recent drilling at Ardich, approximately 60% was completed with HQ core, and the remainder was mostly PQ sized core (very few holes were NQ core). PQ core has a nominal diameter of 85 mm.

Drill core is boxed at the rig by the driller and transported to the sample preparation facility on site for logging by Anagold exploration staff.

Logging includes the collection of lithological, alteration, and structural information. Since 2017, drill core has also undergone a detailed geotechnical logging process including a detailed ‘mining rock mass rating’ to ‘rock mass rating’ system. In addition, core samples are collected every 10 m to undertake point load IS50 testing for uniaxial compressive strength (UCS).

Diamond core that is competent is sawn in half longitudinally with a diamond saw at the core yard. Core that is broken or rubbly is sampled using a spatula to take approximately half the sample. Half the core is placed in a sample bag and the remaining half is returned to the correct position in the core tray. Sample numbers are assigned, and sample tags are placed in the sample bags and recorded in the master sample list. Sample intervals are typically 1 m down-hole.

Prior to 2015, QA/QC samples were collected routinely during the sampling process. CRMs were inserted into each sample job at a rate of 1-in-20. Blank samples were inserted into each sample job at a rate of 1-in-60. Field duplicate samples were collected by cutting the remaining half core portion into two and selecting one quarter of the remaining sample to be submitted as the field duplicate. Field duplicates are collected and inserted into the sample job at a rate of 1-in-40 samples. From 2015 onwards, the field duplicate insertion rate was increased to 1-in-20.

8.1.3 Drillhole Logging and Data Collection

RC chip samples are collected by field staff for review by the logging geologist. Similarly, core samples are metre marked by field staff in preparation for the logging geologist.

Drill core is subjected to detail logging using Anagold geological codes and logging formats. Information captured includes lithology, structure, alteration, mineralisation, and geotechnical data on veining, joint frequency, and joint sets.

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Until September 2019, all geological data was recorded onto hard-copy logs and then transcribed into text files, using data-loading templates, ready for loading into the corporate relational SQL database. Since September 2019, hard copy logs have been replaced with data loading templates on touchpads with direct links to the company server. Files located on the server are uploaded into the corporate database regularly following appropriate checking of the data entry.

Until 2017, the SQL drilling database was managed by the Anagold exploration team located at the Çöpler mine site. Thereafter, the exploration database is controlled and managed by the Anagold exploration team located at the head office in Ankara.

8.2 Sample Preparation

8.2.1 Reverse Circulation Sample Preparation

The majority of historical RC sample preparation was completed at ALS İzmir. From late-2012 through to the end of 2013, pulp samples weighing approximately 150 g were sent to ALS Vancouver. All samples in 2014 were generated and analysed by ALS İzmir. In 2015, samples were sent to SGS for preparation and assay. Since 2015, ALS Global is being used as the main laboratory.

8.2.2 Diamond Drilling Sample Preparation

The majority of historical DD sample preparation was completed at ALS İzmir. From late-2012 through to the end of 2013, pulp samples weighing approximately 150 g were sent to ALS Vancouver. All samples in 2014 were generated and analysed by ALS İzmir. In 2015, samples were sent to SGS for preparation and assay. Since 2015, ALS Global is being used as the main laboratory.

8.3 Sample Analysis

In the period 2004–2014, samples analysed for Au at ALS Vancouver used method Au-AA25, which is a fire assay of a 30 g sample followed by atomic absorption spectroscopy (AAS). The lower and upper detection limits are 0.01 g/t Au and 100 g/t Au respectively. Samples that returned Au grades above the upper detection limit were re-analysed using the gravimetric method Au-GRA21.

Analysis of an additional 33 elements was performed using the ALS Global method ME-ICP61, which involves a four-acid (perchloric, nitric, hydrofluoric, and hydrochloric acid) digestion (four-acid digest), followed by inductively coupled plasma-atomic emission spectroscopy (ICP-AES). Ag, Cu, Pb, Zn, and Mn are among the 33 elements analysed by this method.

In 2015, samples sent to SGS were analysed using the Au fire assay method FAA303, which also uses a 30 g sample and ICP-AES. Detection limits are 0.01 g/t Au. When content was detected above 3 g/t Au, method FAG303 using a gravimetric finish was added.

A 36 element analysis was performed at SGS with ICP40B method, which involves a four-acid digest (4A) followed by analysis via inductively coupled plasma-optical emission spectroscopy (ICP-OES).

From 2016 to recent, samples have been sent to ALS İzmir. Until 2019, Au-AA23 method was used, involving a fire assay of a 30 g sample followed by AAS with the lower and upper detection limits being 0.01 g/t Au and 10 g/t Au respectively. Samples that returned grades above the upper Au detection limit were re-analysed using the gravimetric method Au-GRA21. Since 2019, Au-AA24 method with a 50 g sample and lower detection limit of 0.005 g/t Au has been used. For Au grades above the upper detection limit, gravimetric method Au-GRA22 with a 50 g sample is used.

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8.4 Sample Security

Drill core and RC chips are transported to the core storage facility by either the drilling company personnel or Anagold geological staff. Once at the facility, the samples are kept in a secure location while logging and sampling is conducted. The DD core storage facility is enclosed by a fence and gate that is locked at night and when the geology staff are absent. When samples are transported off site, a commercial carrier is used.

8.5 QA/QC Procedures

The QA/QC programme has historically consisted of a combination of QA/QC sample types that are designed to monitor different aspects of the sample preparation and assaying process.

Blanks consist of non-mineralised samples that are submitted in order to identify the presence of contamination through the sample preparation process. Prior to 2015, blank samples comprised of commercially available pulp samples. As pulp blanks require neither crushing nor pulverising, they are of limited value in terms of identifying contamination through those aspects of the sample preparation process. Therefore, commencing in 2015, the pulp samples were switched to a coarse quartz material that would allow for better monitoring of sample contamination. Blank samples have been inserted routinely into all sample batches. If a blank sample returns an assay grade above an acceptable limit, contamination from a previous mineralised sample is assumed to have occurred at either the crushing or pulverisation stage. The first sample in a drillhole is typically a blank, after which blanks are inserted into the sample batch at a nominal rate of 1-in-60 samples. The insertion rate was updated and for the period 2015–2020 to approximately 1-in-30 for diamond drillholes.

CRM samples are inserted into sample submissions in order to monitor and measure the accuracy of the assay laboratory results over time. CRMs have been inserted into sample submissions at a nominal rate of 1-in-30. The frequency was increased from 3% to 5% in 2015. Several different CRMs have been selected for use at varying Au and Cu grades over the life of the Project. Pulp blanks have been used to determine the accuracy of assay results at very low-grades, and as such are inserted using the same logic as CRMs. The combined insertion rate of pulp blanks and CRMs is a nominal 1-in-20 samples. For the period 2015–2020, the combined rate is approximately 1-in-25.

Field duplicates are used as a means of monitoring and assessing sample homogeneity and inherent grade variability and enable the determination of bias and precision between sample pairs. Field duplicates have been routinely inserted into both RC and DD sample submissions since drilling began. DD field duplicates are generated by cutting the residual half core sample into halves again and submitting one of the resultant quarters of core as the field duplicate. RC field duplicates are generated by splitting the RC sample twice to create two samples from the same interval. Field duplicates have historically been submitted at a nominal rate of 1-in-40 samples. In 2015, the field duplicate insertion rate was increased to 1-in-20. Since 2017 for DD samples, duplicate samples are being collected as laboratory duplicates instead of quarter core field duplicate samples.

8.6 QP Opinion

In the opinion of the QP the sample preparation, security, and analytical procedures meets industry standards for data quality and integrity. There are no factors related to sampling or sample preparation that would materially impact the accuracy or reliability of the samples or the assay results. The outcomes of the QA/QC procedures indicate that the assay results are within acceptable levels of accuracy and precision and the resulting database is sufficient to support the estimation of Mineral Resources.

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9 DATA VERIFICATION

This section has not been changed from the 2021MR and remains the most current study work available.

Independent detailed quality assurance and quality control (QA/QC) analysis is undertaken routinely on data from the Çöpler project.

This work was discussed in detail in the Çöpler District Master Plan 2020 (CDMP20TR). The reader is referred to that report for all QA/QC of data used to develop resource models prior to November 2020.

The QA/QC pertaining to data used in the updated Ardich resource model is described here.

9.1 Çöpler Deposit Data Verification

The independent quality assurance and quality control (QA/QC) review presented in the CDMP20TR confirms that the Çöpler drillhole data sampling and assaying is of a good standard and suitable for the purpose of mineral resource estimation and the reporting of exploration results. This is especially true for gold, which is the primary metal of economic interest. The confidence in the silver, copper, sulfur, and carbon analyses is at a level that at minimum supports modelling for geometallurgical and by-product metal characterisation.

In 2014, an independent database audit and review of available QA/QC data was undertaken to ensure the data are of sufficient quality to support resource estimation (the 2014 audit). The database audit covered data collected from 2000 to December 2013.

A further independent audit of the Çöpler deposit database as of 15 July 2015 was completed that year to verify the data are of sufficient quality to support Mineral Resource estimation of gold, copper, and silver for the Çöpler deposit (the 2015 audit). The 2015 audit focused on the 121 drillholes (12,959.8 m) completed since the 2014 audit. Available QA/QC data were evaluated to ensure the assay data are suitable to support resource estimation.

A data audit covering new data obtained from 2015 through 2020 was completed in June 2020 (the 2020 audit):

•Yetkin, E., 2020 (2020a). Çöpler Project Drill Data Validation, Verification & QA/QC Review. 30 June 2020.

The 2020 audit discusses some minor inconsistencies and outliers but overall confirms the previous findings that the Çöpler drillhole data sampling and assaying is of a good standard and suitable for the purpose of Mineral Resource estimation and the reporting of exploration results.

9.2 Çakmaktepe Deposit Data Verification

The independent QA/QC review presented in the CDMP20TR confirms that the Çakmaktepe drillhole data sampling and assaying is of a good standard and suitable for the purpose of mineral resource estimation and the reporting of exploration results. This is especially true for gold, which is the primary metal of economic interest. The confidence in the silver, copper, sulfur, and carbon analyses is at a level that at minimum supports modelling for geometallurgical and by-product metal characterisation.

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9.3 Ardich Deposit Data Verification

9.3.1 Data Verification – Ardich

Independent data verification was conducted on the Ardich drilling databases and available QA/QC sample data for drilling completed from the first Ardich hole drilled on 1 August 2017 to the established data cut-off date for the Mineral Resource modelling of 29 May 2021.

This verification was completed in stages as drill programmes progressed, and is reported in nine reports:

•Mineral Consultancy, 2018. Ardich Project Drill Data QA/QC Review. 28 February 2018.

•Yetkin, E., 2018 (2018a). Ardich Project Drill Data QA/QC Review. 29 July 2018.

•Yetkin, E., 2018 (2018b). Ardich Project Drill Data QA/QC Review. 29 October 2018.

•Yetkin, E., 2019 (2019a). Ardich Project Drill Data Validation, Verification & QA/QC Review. 8 March 2019.

•Yetkin, E., 2019 (2019b). Ardich Project Drill Data Validation, Verification & QA/QC Review. 31 October 2019.

•Yetkin, E., 2020 (2020c). Ardich Project Drill Data Validation, Verification & QA/QC Review. 30 March 2020.

•Yetkin, E., 2020 (2020d). Ardich Project Drill Data Validation, Verification & QA/QC Review. 30 November 2020.

•Yetkin, E., 2020 (2020e). Ardich Project Drill Data Validation, Verification & QA/QC Review. 30 January 2021.

•Yetkin, E., 2020 (2020f). Ardich Project Drill Data Validation, Verification & QA/QC Review. 30 July 2021.

It is concluded that the Ardich drillhole data sampling and assaying is of a high standard and suitable for the purpose of Mineral Resource estimation and the reporting of exploration results.

9.3.2 Collar Location – Ardich

Collar positions were verified against the pre-mine topographic surface DTM to check for inconsistencies in elevation. The threshold difference between the DTM and the drillhole collar elevation used for validation was a ±4 m difference in data up to 2020, at which time the tolerance was decreased to ±3 m.

One hole was found to have a difference outside the tolerance limits – AR214, with 7.12 m difference. All other differences were <3 m.

As Ardich has not been mined to date, this discrepancy can be resolved by re-surveying the collar location.

9.3.3 Down-hole Surveys – Ardich

All the Ardich drillholes were downhole surveyed using a multi-shot (Devico or Reflex) with readings spaced at 10 m on average (range of 4–110 m).

Seven Ardich holes were found to have no down-hole survey data and three successive survey intervals were found to have large gaps between readings (95–170 m).

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A comparison of successive down-hole survey readings for a given drillhole was undertaken using a maximum 5° variation over 30 m (0.17°/m) in either inclination or azimuth to flag records with excessive deviations. A total of three spurious readings were deemed to be out of acceptable limits, and those data were removed from the resource database.

A recommended magnetic declination correction discussed in the CDMP20TR has been implemented for Ardich data.

9.3.4 Geology, Density, and Geotechnical Logs – Ardich

The drillhole database lithology table was checked for alphanumeric categorical code validity and interval reporting consistency with the log key sheets. No mismatches were identified, and all entries were found to be identical to the codes provided in log key sheets. One lithology interval was shown to have an overlapping FROM–TO and there were three intervals that were missing lithology records.

Some minor discrepancies were identified in other coding in the database, such as lower-case ‘fault’ codes used instead of upper-case, or a lithology of ‘CLASTICS’ rather than ‘CLASTIC’, in both cases causing two different unique categories to be created, and some new codes created in the ‘redox’ and ‘alteration’ tables that do not appear in the log key sheet.

Several logged intervals were re-logged following the identification of incompatible geochemistry, and all of these were updated in the lithology logs used in the 2021 resource modelling dataset.

Density data were reviewed during six of the nine Ardich verification campaigns. Density measurements are collected using the same process described in Section 12.1.4 for the Çöpler deposit. A systematic truncation from four decimal places to three decimal places was observed, and several transcription errors in FROM–TO records were identified. Manually calculated spot check values were within ~2% of the density reading supplied in the resource database. The density samples are representative in a spatial and geological context. On a total project basis, there are no obvious density outliers.

9.3.5 Assays – Ardich

There were two different independent laboratories used for assays and geochemical analyses for the entire Ardich database to date, these were:

•ALS Global

•BV (ACME)

The variety of laboratories resulted in a variety of method codes for fire assay, four-acid digestion, multi-element, and Leco analyses.

In consistency checks on the ‘tblVWDHAssays_ALL’ assay table, four samples were found to have missing assay entries. The highest 1% of assays were checked for transcription errors, with no major errors identified.

9.3.6 Witness Samples – Ardich

No witness samples are known of for Ardich.

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9.3.7 Quality Assurance / Quality Control (QA/QC) Results – Ardich

Ardich QA/QC data was independently reviewed on a campaign basis at milestone times in the evolution of the exploration programme. There are currently nine individual reports describing the results. The collective results are reported in this section.

The Ardich QA/QC programme follows suggested guidelines for QC sample insertion rates:

•3%–5% CRMs and blanks

•5%–10% field duplicates

•3%–5% pulp duplicates

•5% of coarse rejects/pulps to a third-party external laboratory

9.3.7.1 Screen Analyses – Ardich

No screen analysis has been undertaken to date on Ardich material.

9.3.7.2 Certified Reference Material (CRM) Samples – Ardich

The principal assay laboratory for drill samples was ALS İzmir, with umpire samples principally submitted to the BV (ACME) laboratory.

Au CRMs were submitted across the entire Ardich database, plus S (Leco) and C (Leco) CRMs in the later programmes. The average insertion rate was of the order of 3.5%, which meets the guideline.

The performance of the CRM sample data was assessed by plotting the laboratory assay values for Au (FA and CL), Ag (4A), Cu (4A), S (4A and Leco), and C (Leco) of the CRMs against time on control charts.

A review of the CRM results from the samples submitted indicated that both ALS İzmir and BV (ACME) had acceptable overall performances for the listed Au CRMs used during the programme, although ALS consistently had issues with isolated ±2SDs as well as failed cases of ±3SDs. In general, the ALS shows high bias in almost all Au CRMs at varying levels, being more evident in low Au and cut-off Au grades, which are also responsible for the most of the +2SD and +3SD occurrences. Few of these failed cases appeared to be as a result of mislabelling. No unexplained extreme outliers were identified. Several CRMs had insufficient data to identify any change in performance over time.

The performance of Ag, Cu, S, and C CRMs was also reviewed, showing ALS had an acceptable overall performance with isolated cases to be followed up for Ag, Cu and S.

ALS and BV (ACME) performance both for S (4A) and S (Leco) are generally acceptable other than calibration-related bias noted for low-grade CRMs. C (Leco) performance of OREAS20A (ALS and BV (ACME)) and OREAS25A (ALS and BV (ACME)) returned acceptable results both for low-grade and cut-off grade.

Timely monitoring of the CRM performance will ensure that the replicate assays stay within range, that systematic analytical drift is promptly corrected, and that mislabelled samples are promptly identified. The extreme-outlier cases need to be investigated and if these are found to be mislabelling then the organisational procedures should be reviewed and updated. If it transpires that these are not mis-labelled samples and the errors are found to be laboratory-related, then re-assay procedures needed to confirm the assays for the relevant batches.

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9.3.7.3 Blank Samples – Ardich

Blanks were inserted into the sample stream as a check for cross-contamination during sample preparation. The insertion rate was of the order of 3%, which meets the guideline.

For ALS İzmir, Au assays for blanks were assessed by charting the laboratory assay values and assessing performance versus the maximum accepted threshold value of 0.05 g/t Au, which is 10 times the lower DL. All blank assays were below 3DL except for one sample, however it was noted that there were several occurrences where consecutive blanks assayed close to the threshold.

For Ag, all blank assays were below the maximum threshold value of 0.5 g/t Ag. For Cu, the threshold level is 10 ppm Cu and there were several samples that assayed slightly above, at, or close to the threshold value. For sulfur (both for 4A and Leco) the threshold value is 0.1% S and all blank sample assays were below this value. For carbon (Leco) the threshold value is 0.1% C and there are 92 assays above the threshold. These results show that the blank material used to monitor Au and other elements may not be suitable for C analysis, or that the samples are contaminated during the sample preparation. Other than these no obvious contamination issues are apparent within the assay database.

For BV (ACME) blanks returned all below threshold values for Au, Ag, Cu, S (4A), S (Leco), and C (Leco) analyses. Only nine blanks were submitted to BV (ACME), which does not meet the 3%–5% insertion rate guideline.

9.3.7.4 Duplicate Samples – Ardich

Duplicate sample data was analysed to determine the reproducibility of assays according to the combination of geological, sampling, and analytic variances. The insertion rate was of the order of 5%, which meets the guideline.

The duplicates in each of the six QA/QC reviews have an average absolute relative difference of between 0.029–0.140 for Au, with a sample-weighted average of approximately 0.053, which falls within or below the rule-of-thumb of 0.10–0.20 absolute relative difference range for acceptable laboratory duplicate samples for each campaign.

The Au assay variance in each campaign, given by the average percent difference, is within the range of –2.6%–3.8%, with an average of approximately –0.25%, which falls within the rule-of-thumb of ±10% precision window.

The absolute relative difference and average percent difference results were equally encouraging for S (4A) where data was obtained (from drillhole AR56 onwards).

The high precision of the duplicates reflects the inherent sample homogeneity of laboratory-prepared duplicate samples from coarse rejects, which allows more representative sampling of the grade population.

9.3.7.5 Check Assays – Ardich

All six QA/QC campaigns report the results of umpire assays with pulp duplicates submitted to BV (ACME) for independent analysis.

The rate of check assay was lower for the earlier campaigns, as low as 2% in the first campaign, but the overall average is 4.7%, which is approaching the guideline.

Generally, the results show low-level artefacts due to differing DLs between the two laboratories, and the occasional outlier result, but overall, the scatter plots demonstrate strong linear correlation.

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The check assays in each of the six QA/QC reviews have an average absolute relative difference of between 0.042–0.078 for Au, with a sample-weighted average of approximately 0.064, which falls below the rule-of-thumb of 0.10–0.20 absolute relative difference range for acceptable laboratory duplicate samples in every campaign.

Two of the earlier campaigns showed questionable performance for Ag and S; a result that is considered to be moderated by the small number of samples submitted in these early campaigns.

9.3.8 Discussion – Ardich

The independent QA/QC reviews confirm that the Ardich drillhole data sampling and assaying is of a high standard and suitable for the purpose of mineral resource estimation and the reporting of exploration results. This is especially true for gold, which is the primary metal of economic interest. The confidence in the silver, copper, sulfur, and carbon analyses is at a level that at minimum supports modelling for geometallurgical and by-product metal characterisation.

9.4 Bayramdere Deposit Data Verification

The Bayramdere sampling project was part of the near-mine programme that also included the Yakuplu East and Yakuplu South-east areas.

Independent data verification was conducted during and immediately following the 2015 drilling programme on the Project, and a data audit for Bayramdere drilling was completed in January 2016 (Cube Consulting, 2016b).

The independent data verification concluded that the sample data is considered to be of an acceptable standard and appropriate for the purpose of Mineral Resource estimation and the reporting of exploration results.

9.5 QP Opinion

In the opinion of the QP the data is adequate for the purposes used in the CDMP21TRS.

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10 MINERAL PROCESSING AND METALLURGICAL TESTING

This section has not been changed from the 2021MR and remains the most current study work available.

10.1 Oxide Ore for Heap Leaching

10.1.1 Testwork – Çöpler Oxide

Metallurgical testwork for Çöpler oxide ore for heap leaching commenced in September 2004. Much of this testing was carried out by Resource Development Inc. (RDi) of Wheat Ridge Colorado, with oversight from Ausenco Limited of Brisbane, Australia, and Pennstrom Consulting of Highlands Ranch, Colorado. Additional follow-up metallurgical testwork was conducted by AMMTEC Limited (AMMTEC) of Perth, Australia in 2009 and by McClelland Laboratories, and supervised by Metallurgium more recently.

The heap leaching facilities were commissioned at the Çöpler project in late-2010 and have operated continuously since that time. Operations are currently ongoing.

10.1.2 Testwork – Çakmaktepe Oxide

Metallurgical testwork on Çakmaktepe oxide ore for heap leaching was undertaken at the on-site metallurgical laboratory, initially under the supervision of Kappes, Cassiday & Associates. The initial testwork in 2015 undertook bottle roll and column leach tests. The results compare to the Çöpler oxide ore, with similar behaviour and leach kinetics. Subsequently, Çakmaktepe oxide ore was heap leached together with Çöpler oxide ore.

10.1.3 Testwork – Ardich Oxide

Metallurgical testwork on Ardich oxide for heap leaching has been undertaken at McClelland laboratories and supervised by Metallurgium. An initial testwork programme including bottle roll and column leach was carried out in 2019. This initial programme identified two distinct domains with respect to gold recovery based on sulfide sulfur (SS) content of <1% and 1%–2%. The column test results indicated that the listwanite, dolomite, and jasperoid lithologies have physical properties amenable to heap leaching. The column tests were undertaken at a crush size of P80 of 12.5 mm.

This initial test programme has been followed up in 2020 and 2021 with further testwork, with final results yet to be released.

10.1.3.1 Ardich Crushing Testwork

Crushing testwork on six Ardich composite samples was performed as part of the 2019 McClelland testwork programme, Crushing Work index (CWi), and Abrasion index (Ai). The CWi values ranged from 4.0–6.9 kWh/t, indicating that the material was very soft. The jasperoid was the hardest material, with a CWi of 6.9 kWh/t. The Ai values ranged from 0.12-0.90. The jasperoid was the most abrasive (0.90, Very Abrasive), whereas all other lithology types ranged from 0.12–0.26 (Abrasive to Moderately Abrasive).

10.1.4 Testwork – Bayramdere Oxide

Metallurgical testwork has been completed to characterise the Bayramdere oxide mineralisation and determine its suitability for potential heap leaching. In total, five PQ (85 mm diameter) DD (diamond core drilling) holes were completed in 2014 for this purpose and 91 m of half-core have been provided for intermittent bottle roll leach (IBRL) test and column leach testing.

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In the IBRL tests, the gold extraction ranges from 54% to 97% at the end of 11 tests with the consumption of 0.85 kg/t NaCN.

In the column test, gold extraction is 84% in the two duplicate columns.

Final gold extraction in column testing is approximately 84% with reasonable leach kinetics. The extracted gold quantity will be economic for heap leach processing if haul costs are not excessive.

10.1.5 Heap Leach Gold Recovery

The heap leaching process gold recovery assumptions have been updated to reflect actual performance of the operation. The gold recovery assumptions are summarised for Çöpler oxide in Table 10.1, Çakmaktepe oxide in Table 10.2 (including Bayramdere), and Ardich oxide in Table 10.3.

Table 10.1 Çöpler Gold Recovery Assumptions for Heap Leaching of Oxide

Oxide Ore Type		Çöpler Zone
	Manganese		Marble		Main		Main East		Main West		West
Diorite		71.2		62.3		71.2		71.2		62.3		62.3
Metasediment		66.8		66.8		66.8		66.8		66.8		66.8
Limestone / Marble		78.4		75.7		68.6		78.4		75.7		75.7
Gossan		71.2		65.1		71.2		71.2		65.1		65.1
Manganese Diorite		71.2		62.3		71.2		71.2		62.3		62.3

Table 10.2 Çakmaktepe Gold Recovery Assumptions for Heap Leaching of Oxide (incl. Bayramdere)

Oxide Ore Type		Çakmaktepe Zone
	Central		North		East		South-east		Bayramdere
Limestone / Marble		70.0		59.0		67.0		–		75.0
Metasediment		80.0		14.0		–		–		–
Gossan		–		59.0		67.0		75.0		75.0
Jasperoid		73.0		59.0		–		–		–
Diorite		61.0		38.0		–		–		–
Ophiolite		70.0		63.0		67.0		75.0		75.0

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Table 10.3 Ardich Gold Recovery Assumptions for Heap Leaching of Oxide

Ore Type		Ardich Zone
	Main		East
Sulfur <1%
Jasperoid		50.0		50.0
Listwanite		73.0		55.0
Dolomite		73.0		55.0
Sulfur 1%–2%
Jasperoid		40.0		40.0
Listwanite		58.0		45.0
Dolomite		58.0		45.0

The original gold recovery assumptions for Çöpler ores were developed in 2008, based on the results of column leach and bottle roll testing performed by RDi between 2005–2008. These recovery assumptions are reviewed and updated annually based on the following information:

•An analysis of the results of additional column leach and bottle roll tests performed on monthly composite samples of heap leach feed material conducted at the Çöpler project from July 2011 through December 2019.

•Use of a MS Excel-based heap leach production model that is calibrated against actual gold production data at the Çöpler mine from start-up.

The recovery values listed in Table 10.1, Table 10.2, and Table 10.3 consider heap leaching of ore crushed to 80% passing 12.5 mm, agglomerated, and placed on a lined heap leach pad for treatment.

10.2 Sulfide Ores

Sulfide material (i.e., material with >2% sulfur content) is not suitable for treatment by the heap leaching process.

10.2.1 Historical Testwork – Çöpler Sulfide

Historical testing was conducted on samples from the sulfide material in several phases. RDi performed several sulfide processing scoping-level investigations from 2006–2009. A two-phase programme on sulfide samples was conducted at SGS laboratory in Ankara, Turkey (SGS) in 2009 and 2010 to support a pre-feasibility study (PFS) completed in 2011, (Samuel, 2011). A QEMSCAN (quantitative evaluation of minerals by scanning electron microscopy) mineralogy study on three sulfide (and six oxide) samples was performed by AMMTEC in December 2008.

The historical work completed at both RDi and SGS concentrated on evaluating sulfide processing options, including direct cyanidation, flotation, cyanidation of flotation concentrates, pressure oxidation (POX) coupled with cyanidation, and roasting coupled with cyanidation. The evaluation of the historical data in the PFS resulted in the selection of POX coupled with cyanidation as the process to further evaluate with testing and a FS.

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Initial metallurgical testwork carried out by RDi indicated that 11%–30% of the gold content in the Çöpler sulfide material may be amenable to whole-ore cyanidation, as demonstrated by diagnostic leaching. Between 60%–80% of the gold content was found to be associated with sulfide minerals and would require some type of oxidation step to liberate the gold for cyanidation.

The RDi scoping studies indicated that pre-treatment using POX was the most effective treatment and displayed the potential to achieve greater than 90% gold extractions. Flotation tests indicated that gold could be recovered by flotation, but the concentrates were low-grade with relatively high mass pulls, and relatively low-gold recovery. Testwork indicated that flotation concentrate, and tailings did not leach well using cyanide, even after being finely ground.

10.2.2 Sulfide Mineralogy

In December 2008, Anagold commissioned AMMTEC to complete a QEMSCAN precious metals search (PMS), trace mineral search (TMS), and energy dispersive spectra signal (EDS) mineralogy analyses performed on three sulfide mineralisation samples. Analyses were performed on samples of diorite, metasediment, and massive pyrite rock types.

The findings from the 2008 QEMSCAN analyses indicated that the gangue mineralisation in the sulfide mineralisation is composed mainly of quartz, micas / clays, and feldspars, (displaying relative abundances of approximately 31%, 27%, and 21%, respectively). The sulfide mineralisation consists of pyrite, arsenopyrite, chalcopyrite, and sphalerite.

A gold deportment study was performed by AMTEL Ltd. (AMTEL) on samples of MC4 composite after flotation separation. Although flotation was not part of the flow sheet, it is a useful method of concentrating the sulfides (the main gold carriers) to improve analysis statistics.

The combined concentrate represented 18.5% of the feed mass and assayed 9.8 g/t Au and 23% SS. Recoveries of gold and sulfur to concentrate were 72.7% and 90% respectively. Flotation tailings assayed 0.68 g/t Au and 0.48% SS.

The detailed mineralogical analysis confirms that the gold is primarily carried by sulfide minerals. In the calculated head, 83% of all gold is in sulfides (free or locked) and only 2.4% was held in rock. The remainder of the gold (14%) was present as free gold, and this correlates well with a direct cyanidation recovery of only 17% when the ore was ground to a P80 of 90 µm.

Of the gold that is in sulfides, the majority (78%) is in sub-microscopic form. This confirms the refractory nature of the ore and explains why oxidation of the sulfides is necessary to make the gold available for leaching. Arsenopyrite was the sulfide mineral found to have the highest contained gold, averaging 123 g/t Au by one measure and 182 g/t Au by a second. Gold in pyrite was more than an order of magnitude lower than arsenopyrite and averaged 7.0 g/t Au. Marcasite, a mineral chemically similar to pyrite, carried an average of 17.8 g/t Au. Of the gold contained in sulfides, 50% was found to be in arsenopyrite, 25% in pyrite, and 20% in marcasite.

In summary, the AMTEL gold department study is consistent with previous mineralogy studies and confirms that a large portion of the gold is present as sub-microscopic particles, primarily in sulfides, largely arsenopyrite. The study also concluded that whole-ore oxidation would be required as a pre-treatment to cyanidation to liberate the majority of the gold contained in the sulfide materials.

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10.2.3 Direct Cyanidation

Hazen performed direct cyanidation carbon-in-leach (CIL) tests at various grind sizes with no pre-treatment on the individual sulfide rock type composites to establish baseline gold extractions. The goal of these tests was to examine gold extraction variability with grind size. These samples were subsequently used to prepare feed composites used in the Hazen pilot plant programme.

The testwork demonstrated that the bulk of the Çöpler sulfide samples are refractory to direct cyanidation, and that extractions do not improve significantly with finer grinding.

Table 10.4 Gold Deportment in Flotation Separated Streams

Form and Carrier of Gold	Concentrate<br>(g/t)	Tails<br>(g/t)
Assayed Grade	10.187 ± 0.167	0.837 ± 0.028
Free / Liberated Gold Grains
>40 µm	0.106	0.004 *
5–4 µm	0.346	0.003
<5 µm	0.871	0.146
Exposed Associated Gold Grains
Free Sulfides +5 µm	0.350	0.018
–5 µm	–	–
Rock-Sulfide Composites	0.125	0.052
Rock Particles	0.021	0.035
Enclosed Associated Gold Grains
Free Sulfides +5 µm	0.977	0.007
–5 µm	0.292	0.029
Rock-Sulfide Composites	0.338	0.023
Rock Particles	0.014	0.031
Sub-microscopic Gold
Free Sulfides +5 µm	4.156	0.020
–5 µm	1.244	0.157
Associated Sulfides	1.605	0.304
Total (mineralogically counted)	10.444(102.5%)	0.829(99.0%)

* From a very small number of grains (1 free grain, from ~2 kg of material)

10.2.4 Flotation Testwork

Flotation testwork has been undertaken on Çöpler sulfide samples since before 2006 with a series of testwork programmes and studies undertaken by RDi, FL Smidth, and the on-site metallurgical laboratory.

Initially the testwork was focused on development of a viable flowsheet to recover gold to enable subsequent recovery as doré. This work was unsuccessful due to a generally poor flotation response, resulting in the adoption of the current POX and CIP gold recovery flowsheet.

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In 2019, flotation was again considered for incorporation into the POX / CIP circuit to improve both sulfur and gold recovery and enable the POX circuit to operate at optimum conditions.

Testwork was conducted on fresh material from the existing sulfide circuit. A total of 20 tests were conducted as part of this programme.

The key variables considered in determining throughput for flotation are SS flotation recovery and flotation mass pull. Gold recovery to concentrate and gold recovery of the flotation tails are also determined. Of the 20 tests undertaken, a total of eight flotation testwork tests are considered representative due to their relative commonality of flotation conditions, and the SS feed grade is within the range that the flotation plant is expected to operate. The results ranged from 65% to 81% SS recovery, and 43% to 55% Au recovery to concentrate.

The mass pull for sulfide flotation is typically related to SS grade. Figure 10.1, shows the relationship of mass pull to SS feed grade.

Figure 10.1 Feed SS% – Mass Pull Relationship

Anagold, 2020

Float Concentrate Mass Pull = 277.09 x Feed SS%2 – 15.165 x Feed SS% + 0.3298

10.2.5 Testwork – Comminution

The comminution properties for the three major ore domains (metasediment, diorite, and manganese diorite) have been measured during all testwork stages. Rock competence drives semi-autogenous grind (SAG) mill selection, Bond Work index (BWi) drives ball mill selection, and Ai is used to estimate media and mill liner consumption rates. The major domains exhibit moderate comminution characteristics.

As part of the flotation circuit sizing, the throughput capacity of the installed crushing and grinding circuit was determined on review of testwork and plant actual performance.

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The review of the grinding circuit determined that the throughput has exceeded design expectations since commissioning due to the processing of ore that is softer than the design comminution testwork identified. The design maximum feed rate of 306 t/h was achievable with close to full milling power being consumed. However, an average throughput rate of 370 t/h was achieved in the period late-2019 through early-2020 with the SAG and ball mills drawing approximately half of their design power.

A simulation model of the comminution circuit was prepared (in JKSimMet) and calibrated to this actual plant performance. This calibrated simulation was then used to estimate plant performance with future harder ores, having properties approximating design expectations. Wood’s simulation showed that the plant is expected to be able to process the target rate of 400 t/h of design-hardness ore with the mills at full design operating power.

10.2.6 Testwork – POX

Three continuous pilot plant programmes have been conducted for the POX sulfide plant: the first two programmes at Hazen Research, Inc. (Hazen) comprising a total of four test campaigns, and the third programme at SGS Lakefield Oretest, Perth (SGS Perth). Three campaigns were completed during the first pilot plant programme, with the first campaign commencing in February 2012. The second pilot programme incorporating one campaign, was conducted in December 2012. The third pilot programme, conducted in August 2015, included a single campaign that tested multiple lithologies at high and low-acidulation extents.

The pilot plant facility for the first pilot programme included the following continuous circuits: acidulation, POX autoclave, hot cure (HC), primary neutralisation (PN), six-stage counter current decantation (CCD), and mixed sulfide precipitation (MSP). Ore preparation (grinding), cyanidation, activated carbon gold recovery, cyanide destruction, tailings neutralisation, and final tailings production were all completed on a batch basis.

In 2015, Anagold performed confirmatory pilot testing on a range of ore-types and composite blends treated at ‘high’ and ‘low’ acidulation conditions. This programme comprised a single pilot plant campaign, Campaign 5, which was conducted at SGS Perth during August and September. Apart from testing the impact of acidulation chemistry, one of the key purposes of the campaign was to produce samples for repeat thickener vendor testing. This was prompted by the inconsistent vendor data generated during Campaigns 1–4.

10.2.7 Testwork: Pyrite Recovery from Copper-Rich Ores

Preliminary metallurgical testwork has been undertaken to investigate the potential to produce a copper–gold concentrate for sale and a pyrite concentrate to supplement POX operations utilising a copper-rich portion of the Çöpler resource.

The testwork was conducted at ALS using drillhole samples to produce a master composite and eight individual composites representing copper-bearing zones of the Çöpler mine. The composites copper grades were between 0.05%–0.43% Cu, and 0.16–1.54 g/t Au. Silver grades were between 1.0–3.0 g/t. An elevated arsenic content was measured for Composite 5 at approximately 0.16% As. Sulfide sulphur to copper ratios for many of the composites indicated potentially high pyrite to copper sulfide ratios, which would require chemical conditions to control pyrite flotation.

Results of the mineralogical analysis indicated that chalcopyrite was the predominant copper sulfide mineral. Approximately 4% of the copper was measured as bornite, and 1% as secondary copper sulfide minerals and arsenic sulphosalts tennantite and enargite.

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A flotation flowsheet was developed that included a copper and pyrite circuit with primary grinding to a nominal 150μm K80. To control pyrite recovery, lime was used to elevate both the copper roughers and copper cleaner pH to 10, and a dithiophosphate collector 3477 was used as the copper collector. The copper rougher concentrate was reground to approximately 25 μm K80 to produce high grade concentrates from three cleaner stages for most of the composites. The copper circuit tailings fed a pyrite circuit where potassium amyl xanthate (PAX) was used as the pyrite collector. In a locked-cycle test, approximately 83% copper and 50% gold was recovered to a copper concentrate, which measured approximately 27% Cu, 46 g/t Au and 0.3% As. Approximately 17% gold was recovered to the pyrite concentrate, which measured 4 g/t Au.

Gold recoveries did not trend with sulfur recoveries, therefore a strong association of gold with pyrite does not appear to be evident. However, for Composite 5, which measured the highest arsenic content of the individual composites, gold recoveries trended closely with arsenic recoveries to the product streams, thereby potentially indicating a close association of gold with arsenopyrite for this particular feed type.

Comminution testing was completed with unique comminution composites, representing the same feed material as that used for the flotation test but using drill core and crushed rock samples. The composites were characterised as soft-to-medium hardness with respect to ball milling, and Bond ball mill work indices ranged between 11.0–15.7 kWh/t when using a closing screen of 150μm. Axb values derived from SMC tests ranged between 56 and 124. The Bond Crusher work index for six composites tested ranged between 3–7 kWh/t, which indicated very soft material in terms of crushing.

Further flotation testing is suggested to determine whether improvements to the copper and gold performance in the copper circuit is possible with changes to collector type and dosage, and copper regrind discharge sizing. Additionally, once a flowsheet is optimised, variability testwork is recommended

10.2.8 Overall Circuit Performance

The recovery of gold across a laboratory carbon-in-pulp (CIP) circuit was measured for a number of variability samples representing each of the three major ore types.

In addition to the testwork, the commercial sulfide POX plant commenced commissioning in December 2018, with actual results reviewed to validate the recovery.

10.2.8.1 POX Gold Recovery

The gold recovery results of the acceptable tests are plotted in Figure 10.2, Figure 10.3, and Figure 10.4, together with an appropriate recovery model curve in each instance.

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Figure 10.2 Metasediment Gold Recovery Results and Model

Anagold, 2016

The results are plotted in terms of feed grade so that predictions of recovery during operations can be made by knowing the feed grade.

Figure 10.3 Diorite Gold Recovery and Model

Anagold, 2016

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Note that Figure 10.2 and Figure 10.3 show a number of results that tend to form a regular curve at the top of the datasets. In each instance, where the results are on this curve the solid tails Au grade was below the limit of detection and an assigned tails grade, equal to half the limit of detection, was set for calculation purposes.

Figure 10.4 Manganese Diorite Gold Recovery and Model

Anagold, 2016

The recovery model is represented by the equation:

Gold Recovery (%) = a x (1 – exp (–b x (Au head grade in g/t – c))) + d

Parameter ‘a’ is the only one of the four that has a direct process meaning, representing the maximum recovery the equation can generate. The parameter ‘d’ represents circuit losses in a commercial operation.

The parameters used to generate the curves in Figure 10.2, Figure 10.3, and Figure 10.4 are shown in Table 10.5, and include an allowance for operational losses of 1%.

Table 10.5 Gold POX Recovery Model Parameters

Material Type	a	b	c	d
Metasediment	97.7	1.4	–1.4	–1.0
Diorite	98.3	1.4	–1.5	–1.0
Manganese Diorite	96.7	1.2	–1.4	–1.0

The POX commissioning and ramp-up allowances in Table 10.6 have been made on top of the base recoveries.

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Table 10.6 Commissioning and Ramp-up Allowances

Recovery Corrections	Gold Recovery Deduction<br>(%)
Commissioning to June 2019	–3.30
Ramp-up July 2019 to June 2020	–2.30
Flotation Commissioning	–0.75

10.2.8.2 POX Silver Recovery

The silver recovery pattern is much less clear than gold because silver is not released by the oxidation process. Silver recovery is determined from actual plant recovery over the period January 2019 through February 2020.

The silver recovery calculates to 3.0%.

10.2.8.3 Flotation Gold Recovery

From the testwork, it is estimated that the flotation concentrate reporting to the POX circuit will achieve the same overall recovery as the ore directly reporting to POX. Gold recovery to the flotation concentrate is estimated to be 55%.

The flotation tails reporting directly to the leach circuit is estimated to have a gold recovery of 43%, based on testwork.

An allowance of 0.75% reduced gold recovery during commissioning and ramp-up of the flotation circuit (Year 1 of flotation operation) has been included.

10.3 Mineral Processing and Metallurgical Discussion

A large amount of POX testwork has been performed on Çöpler sulfide ore across several pilot plant campaigns. The processes used have been shown to be robust, as demonstrated through operational performance during commissioning, ramp-up, and operations.

The addition of a flotation circuit to the sulfide plant is estimated to provide stability and flexibility to the POX circuit operation to maximise throughput and oxygen utilisation by maintaining optimum sulfur grade to the autoclaves.

The final construction and commencement of commissioning of the flotation circuit in January 2022 is expected to confirm the assumptions developed in the design.

Ongoing testwork and analysis is also recommended on POX oxidation and leach recovery to improve and optimise circuit performance. This should include detailed assessment of gold deportment in final tailings.

Further metallurgical testing of Ardich material types, both oxide and sulfide, is recommended to optimise the feeds to the heap leach and POX and flotation circuits, respectively.

Further flotation testing on the copper–pyrite flowsheet is suggested to determine whether improvements to the copper and gold performance in the copper circuit is possible with changes to collector type and dosage, and copper regrind discharge sizing. Additionally, once a flowsheet is optimised, variability testwork is recommended.

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10.4 QP Opinion

In the opinion of the QP, the data is adequate for the purposes used in the CDMP21TRS and the analytical procedures used in the analysis are of conventional industry practice.

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11 MINERAL RESOURCES ESTIMATES

This section has not been changed from the 2021MR and remains the most current study work available. The CDMP21TRS QPs have reviewed and accepted this information for use in the CDMP21TRS.

Mineral Resources for the Project have been estimated using industry best practices and conform to the requirements of S-K 1300.

The resource model for Ardich has been updated in 2021 and is reported in detail here.

11.1 Çöpler Deposit

At Çöpler, a resource model was constructed to define the geometry of the gold mineralisation. Grades were estimated using exploration drilling data and then calibrated against the production grade control data. Steps for the gold modelling process included:

•Creation of wireframes that constrain gold mineralisation.

This step incorporated structural trends to guide the shape of the wireframes along known geological features within the deposit. Mineralised trends commonly followed lithological contacts, such as the diorite / marble contact, and structural features identified by surface mapping.

•Gold mineralisation was then estimated using a method termed probability assigned constrained kriging (PACK) and then trimmed using the gold mineralisation shell.

PACK first uses a probabilistic model or envelope (indicator envelope) to define the limits of the potentially economic mineralisation. The model cells and drillhole composites within these indicator envelopes were then used for grade estimations. The PACK process was designed to prevent economic grades inside the indicator envelope from being smeared into the waste and restricts low-grade material outside the indicator envelopes from diluting the mineralised material inside the envelope.

•The parameters used to construct the indicator envelopes were calibrated such that the estimated tonnes and grades approximated the historical production data.

Au, Ag, and Cu were interpolated into the parent cells using ordinary kriging (OK), while As, Mn, Fe, and Zn were interpolated using inverse distance method, weighted to the power of two (ID2).

11.1.1 Çöpler Mineral Resource Estimate – Key Assumptions

The estimation methods at Çöpler were designed to address the variable nature of the epithermal, structural, and disseminated styles of gold mineralisation, while honouring the bi-modal distribution of the sulfur mineralisation and the oxide / sulfide boundary.

No obvious correlations were observed between Au and total sulfur; they were therefore domained and estimated separately. Au also showed little correlation with lithology and was therefore domained simply according to model zone (Manganese, Main, Marble, and West), to reflect the different trends of the mineralisation that commonly follow structures and lithological contacts (see Figure 11.1).

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Figure 11.1 Çöpler Model Zones

Anagold, 2016

The percentage of total sulfur is the main criterion used to delineate between ‘oxide’ and ‘sulfide’ material types:

•Oxide material (S <2%) is processed using a heap leach method and has a cut-off grade of approximately 0.3 g/t Au.

•Sulfide material (S ≥ 2%) is processed in the sulfide plant and has a cut-off grade of approximately 1.5 g/t Au.

Total sulfur assay data exhibits a bi-modal distribution with a distinct inflection point at 2% S, and also shows a good correlation with logged lithology. The 2% S inflection point also agrees well with a 1% pyrite break point in the drillhole logs.

As a result, sulfur was modelled using oxide and sulfide sub-domains within each lithology, and gold PACK models were constructed separately for oxide and sulfide within each lithology using the respective Au cut-offs.

The gold models were then reconciled to historical production data and the resource modelling parameters were adjusted to best match the historical data. Mineral Resource categories were applied to each model cell based on a combination of parameters including drillhole density and data quality.

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11.1.2 Çöpler Base Indicator Model

A parent cell size of 10 m x 10 m x 5 m was selected, with the 10 m easting and northing dimensions representing approximately one half the average drillhole spacing, and the 5 m height of the cells representing the mining bench height. Cell model prototype parameters are provided in Table 11.1. The Mineral Resource model has an implicit selective mining unit (SMU) size of 5 m x 10 m x 5 m. The cell model is not rotated.

Table 11.1 Çöpler Block Model Parameters

Direction	Minimum<br>(m)	Maximum<br>(m)	Range<br>(m)	Cell Size<br>(m)	No. of Cells
East	457,100	461,100	4,000	10	400
North	4,362,500	4,365,100	2,600	10	260
RL	400	1,750	1,350	5	270

Drillhole intervals were composited to 10 m down-hole lengths and then assigned Au indicator values based on their composited Au grade. The sulfur indicator values were assigned to 5 m composites. Composites below the threshold were assigned ‘0’ and composites at or above the threshold were assigned ‘1’.

Gold and sulfur indicator values were then interpolated into the parent cell model. The interpolated indicators represent a distance-weighted average of the composite indicators that occur within the search neighbourhood and therefore have values anywhere in the range 0–1. The interpolated indicator was used to create an envelope encapsulating the mineralisation above 0.3 g/t Au (the indicator envelope).

Exploratory data analyses (EDA) and capping studies were performed on samples within the indicator envelope.

11.1.3 Çöpler Domains

The model cells within the indicator envelope were assigned into four zones that represent the four geologically distinct zones (Manganese, Main, Marble, and West) using wireframe solids.

The position of the boundary between the Manganese Zone and the Main Zone was selected between discrete diorite intrusive events. The boundaries for the Marble Zone were selected along one limb of a diorite intrusion associated with a region of higher grade gold mineralisation. The boundary direction then follows the north-easterly trend of the mineralisation. The extension of this boundary includes a larger discrete diorite intrusion that carries minor gold mineralisation along its contact with the metasediment.

The tops of the model zone boundaries wireframe solids were trimmed to the original (pre-mining) topography.

11.1.4 Çöpler Geological Model

Exploration drillhole data and surface mapping were used to create 3D solid interpretation wireframes for the four main geological units: marble, diorite, metasediment, and manganese diorite. Surface mapping was used to provide indicative contact locations in areas of sparse drilling. In areas where the two datasets did not match, priority was given to the drillhole data. Blasthole data were not used to generate the lithology interpretations but were referenced to provide guidance in zones of wide-spaced drilling and in areas with missing drillhole data. The interpretation was adjusted in the Manganese Zone after referencing the blasthole data.

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A typical cross-section illustrating the lithology interpretation at Çöpler is shown in Figure 11.2.

Figure 11.2 Çöpler Lithology – Marble Zone, Cross-Section 459,700 mE (looking west)

Anagold, 2016

11.1.5 Çöpler Data Summary

The cut-off date for the export of the drillholes from the database to be used in the resource modelling was 15 July 2015. The extract contained 1,957 drillholes with a total of 297,798.2 m of drilling. Of this, a total of 1,880 drillholes have collar coordinates within the extents used to construct the resource model. In general, the drillhole spacing ranged from 5–60 m, averaging approximately 20 m. Most drillholes are either vertical or inclined at 60°. Approximately 2% of the drillholes had missing assays; these were set to a null value and not used in the statistics or mineral resource estimation.

11.1.6 Çöpler Exploratory Data Analysis

Detailed exploratory data analysis (EDA) was conducted on the Çöpler resource modelling dataset. This is discussed in detail in the CDMP20TR. A summary of findings of the statistical analyses follow.

11.1.6.1 Çöpler Statistical Summary

Detailed statistical analyses were undertaken to assist with the understanding of the mineralisation distribution in the various domains. The statistical review included typical univariate statistics (tabulations, histograms, box plots) and bivariate statistics (scatter plots, correlations).

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A summary of key findings follows:

•A histogram of sulfur grade in the 1 m composites shows a bi-modal distribution, with the lowest mode at or near trace S (27% of the dataset), and the second mode at approximately 3.25% S (7% of the dataset).

•Mean Au grade statistics are similar for diorite, metasediment, and marble but higher in the manganese diorite. When reviewing the data spatially, however, the higher grade Au mineralisation commonly occurs along the lithological contacts.

•Mean Ag grades are similar for diorite and metasediment, but lower in marble and higher in manganese diorite.

•Mean Cu grades varied between lithologies, but in general are higher in the diorite and metasediment.

•Mean Au grades in diorite, metasediment, and the marble are higher within the sulfide material. Manganese diorite carries a higher mean Au grade within the oxide material relative to the sulfide material.

•Distinctively different sulfur populations were observed for each lithology (although each lithology hosts both low and high-sulfur mineralisation) suggesting that sulfur should be domained by lithology for estimation. This approach was taken on the current model.

•The diorite, metasediment, and manganese diorite showed similar As grades, but the marble As was lower.

•There is moderate correlation between:

•Au and As

•Cu and Fe

•Minor correlations occur between:

•Au and Ag

•Ag and As

•Ag and Mn

•While correlation probably exists between gold and sulfur on a mineralogical level, as suggested by the correlation between gold and arsenic, and the observed presence of arsenopyrite (FeAsS), this correlation is probably masked by the much larger episode of non-auriferous sulfide mineralisation. This suggests that it is reasonable to model silver, copper, zinc, arsenic, and manganese using the gold statistical model.

•Regarding core recovery:

•No correlation was identified between any of the elemental grades and core recovery, and

•There is no obvious increase or decrease in Au grade with lower core recovery.

•In a twinned hole analysis, RC and DD showed good agreement:

•the average RC Au grade was slightly higher than the average DD hole grade.

•No significant changes in grades were noted for the RC holes above or below the water table.

•For sulfur, little difference in grade was noted between DD holes and RC holes.

•For Cu, little difference was noted between the DD holes and RC holes, but the grades were very low.

•Contact plots were constructed for the different combinations of lithological contacts and categorised by material located within the oxide or sulfide portion of the deposit. In general, no hard contacts were observed for Au. The higher grade Au mineralisation commonly occurs along the lithological contacts, which indicated that the gold mineralisation should not be modelled separately for each of the lithological domains.

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11.1.7 Çöpler Top Cutting

In mineral deposits with skewed distributions, it is not uncommon for a small number of the highest assays to account for a significant and disproportionate quantity of the total metal content in the model estimates. Although these assays are real and reproducible, they commonly show little continuity, and can add a significant amount of uncertainty to a mineral resource estimate.

One method of constraining the influence of these samples is to apply a top cut to the assays before compositing and grade estimation. Top cutting was performed on the 1 m composites prior to compositing into the 5 m composites used for the grade estimations. Au was studied and capped by domain and low and high-sulfur category. Top cut thresholds for Ag, Cu, S, As, Fe, Mn, and Zn were applied globally. The top cut thresholds applied before compositing are summarised in Table 11.2 and Table 11.3.

Table 11.2 Çöpler Top Cuts for Au

Domain	Top Cut Au<br>(g/t)
Oxide (S <2%)
Manganese Zone	18
Main Zone	16
Marble Zone	30
West Zone	16
Sulfide (S ≥2%)
Manganese Zone	18
Main Zone	14
Marble Zone	25
West Zone	14

Table 11.3 Çöpler Top Cuts for Non-Au Elements, Applied Globally

Element	Unit	Top Cut
Ag	g/t	300
Cu	%	5
S	%	20
C	%	13
As	ppm	30,000
Fe	%	50
Mn	ppm	100,000
Zn	ppm	60,000

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11.1.8 Çöpler Drillhole Compositing

Samples used for grade estimation were prepared by first compositing the raw sample lengths to 1 m down-hole intervals. Au composites were capped globally at 40 g/t Au for the EDA. The 1 m composites were subsequently top cut at the relevant threshold according to the statistics of each model zone and oxide / sulfide domain. These 1 m composites were then composited into 5 m down-hole for additional statistical analysis and grade estimation.

The 5 m composite interval for grade estimation was selected as it was considered to notionally match the mining bench height. The 5 m composites were not truncated at lithological contacts, nor domain boundaries.

11.1.9 Çöpler Variography

The EDA showed that the trends of the Au mineralisation followed lithological contacts and structures that vary by domain. As a result, variograms (correlograms) were calculated for Au, Ag, and Cu composites for each domain categorised by oxide (S <2%) and sulfide (S ≥2%).

The directions of the anisotropy axes were determined by creating multi-directional variograms, variogram models, and visual observation of the tabular shaped trends of the mineralisation. After the anisotropy had been determined, three directional variograms were calculated and modelled in each of the three primary directions of anisotropy. Given the low and high-sulfur domain variograms showed similar structures, albeit with the low-sulfur domain variogram structures better defined, the low-sulfur domain variograms were used for the grade estimation.

11.1.10 Çöpler Resource Model Estimation

11.1.11 Çöpler Sulfur Model

The total sulfur model was designed to emulate the hard 2% S threshold used during ore control to delineate material to be processed on the heap leach pad or sent to the pressure oxidation (POX) plant.

EDA showed that sulfur should be modelled separately in each of the four main lithological units (diorite, metasediment, marble, and manganese diorite). The sulfur estimate proved to be very sensitive. Minor changes in the estimation parameters causes the reclassification of material from high to low-sulfur and vice versa. The change in the sulfur categorisation has an impact on what cut-off grade is used and what mining and processing cost is applied.

To match the proportion of material greater than and less than 2% sulfur in each lithological domain, a sulfur indicator was generated using a discriminator of 2% sulfur. To accomplish this, a sulfur indicator field was created in the drillhole data, and populated as follows:

•S Indicator = 0 where S < 2%

•S Indicator = 1 where S ≥ 2%

The S indicator was then interpolated into the cell model using nearest neighbour (NN) and inverse distance method, weighted to the power of two (ID2) methods. The ID2 interpolated indicators represent a distance-weighted average of the composite indicators and therefore have values anywhere in the range 0–1. In contrast, the NN interpolated indicators represent only the closest composite indicator and therefore can only have the value ‘0’ or ‘1’.

The number of cells above and below 2% sulfur was initially defined using the NN result (Indicator 0 = S <2% and Indicator 1 = S ≥2%). The ID2 indicator estimate was calibrated against the NN model to make the proportion of low and high-sulfur material honour the NN proportions.

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Sulfur indicator ID2 estimate thresholds that honoured the results of the NN estimation for low and high-sulfur proportions were:

•Diorite = 0.50

•Metasediments = 0.51

•Marble = 0.26

•Manganese diorite = 0.36

A soft boundary approach was achieved at lithological contacts by slightly raising the maximum indicator estimate for the oxide estimate and lowering the minimum indicator estimate for the sulfide estimate.

The sulfur model was not constrained by the mineralisation envelope. This means sulfur was also estimated into the waste rock cells; this was for the purpose of waste rock characterisation.

11.1.12 Çöpler Gold and Other Metal Models

A total of nine elements, Au, Ag, Cu, S, C, Zn, Fe, As, and Mn were estimated. Au, Cu, and Ag were estimated using OK and the remaining elements were interpolated using the ID2 method. Zn, Fe, As, and Mn, which are only used for material type classification, were restricted to within the mineralisation envelope. All cells were estimated using a discretisation matrix of 3 x 3 x 1.

The volume of the mineralisation envelope was calibrated to past production by:

1. Creating a production cell model:

•Constructing a 3 m x 3 m x 5 m cell model in the areas that had already been mined.

•Populating the 3 m x 3 m x 5 m cells with the ore control tonnes and grades estimated from blasthole assays.

•Tabulating ore control tonnes and grade from January 2014 through October 2015.

2. Building an indicator model and estimation of gold grade:

•The low-grade estimates were achieved using an indicator approach defined by an 0.3 g/t Au discriminator. First a low-grade Au indicator field was established in the drillhole 5 m composite file: if the composite grade was <0.3 g/t Au, the low-grade indicator field was set to zero (IND1=0); if the composite grade was ≥0.3 g/t Au, the low-grade indicator was set to one (IND1=1). The low-grade indicator was then interpolated into all cells using ID2, and those cells with an estimated low-grade indicator of greater than 0.3 (i.e., IND1 > 0.3) were selected to define the indicator envelope. Only composites within the indicator envelope were used to estimate the Au grade.

•Similarly, a high-grade gold estimate was developed using a high-sulfur indicator model with a discriminator of 1.5 g/t Au to reflect the higher cut-off required for processing the material through the POX plant. The high-grade gold estimate uses the same indicator estimate threshold of 0.3 (i.e., IND2 > 0.3) to define the boundary limits.

•The low-grade gold estimates were applied to those cells with estimated sulfur grades <2%, and the high-grade gold estimates were applied to those cells with estimated sulfur grades ≥2% S.

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3. Calibrating the PACK model:

•The PACK model parameters were then adjusted so that the gold ounces in the PACK model approximates the gold ounces reported from the ore control model.

•After the gold ounces were calibrated by zone and material type, cells with estimated Au grades below the selected indicator threshold were set to waste.

11.1.13 Çöpler Density Model

Density measurements were performed on representative DD core by the site exploration geologists. Measurements were taken using the wax-coated water displacement method (Archimedes method). Density data were reviewed spatially and statistically. Density values that fell outside expected upper and lower density limits (shown in Table 11.4) were considered to be outliers and removed.

Table 11.4 Çöpler Upper and Lower Density Limits by Lithology

Lithology	Density Lower Limit<br>(t/m3)	Density Upper Limit<br>(t/m3)
Diorite	1.7	3.5
Metasediment	1.7	3.5
Marble	1.7	3.5
Manganese Diorite	–	–

Density values were assigned to the cell model based on rock type and depth below the surface. The density samples were first flagged by lithological code. Since lithological codes were not available for many of the density samples, Lithology was assigned using the lithological wireframes for all density values.

Densities used in the resource model are summarised in Table 11.5.

Table 11.5 Density Values Assigned to the Çöpler Cell Model by Lithology and Vertical Depth Below Surface

Lithology		Depth<br>(m)		No. Density Data		Assigned Density<br>(t/m3)
Diorite		0–20		111		2.22
	20–40		173		2.42
	40–60		155		2.44
	60+		1,653		2.50
Metasediment		0–20		86		2.38
	20–40		209		2.51
	40–60		219		2.54
	60+		1,769		2.63
Marble		all		1,099		2.57
Manganese Diorite		all		23		2.63

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11.1.14 Çöpler Oxidation Model

The oxidation model reflects oxidation due to surficial weathering and/or oxidation resulting from the manganese alteration. Oxide (low-sulfur material (S <2%)) can be processed by heap leaching while sulfide (high-sulfur material (S ≥2%)) is processed through the POX plant.

The low and high sulfur criteria were further finessed using the logged colour codes and pyrite percentages recorded in the drillhole logs. Review of the logs showed a generally relatively sharp colour change from orange–brown tones to grey–black tones (Figure 11.3). A wireframe was constructed to represent this logged colour change. The wireframe was further refined using the logged visual estimates of pyrite. Near-surface material is highly oxidised and usually does not include visually identifiable sulfides, while visual sulfide percentage increases with depth to a point (pyrite ≥1%) where the percent pyrite can be estimated and recorded in the drill logs. In general, the 1% visual pyrite boundary matched the red–grey colour boundary within approximately 5 m, but locally deviated up to 10 m. The 5 m variance is considered to be within the accuracy of the data, as it reflects the composite sample length and the mining bench height.

The resulting oxide / sulfide wireframe boundary was compared to the sulfur estimates model. This comparison showed that the S <2% and S ≥2% domains matched the oxide / sulfide boundary reasonably well, although there are local areas of material with S <2% below the oxide / sulfide surface which are due in part to deeper weathering along structures. As a result, the oxide boundary surface is considered to be somewhat conservative locally in estimating the amount of oxide material.

Blasthole data from Main Zone that contains both Au fire assays (AuFA) and cyanide leach assays (AuCN) show that the gold recovery significantly decreases below the oxide / sulfide boundary. This implies there is low-sulfur material below the oxide / sulfide boundary that has not oxidised, and hence lower recoveries are obtained by cyanide leaching. As a result, the oxide / sulfide boundary is used in the Main Zone to delineate material types. In the Manganese Zone and Marble Zone, however, the estimated sulfur content is used to delineate material.

In the eastern portion of the Çöpler deposit, the oxidation profile is better-developed and follows the diorite intrusion. This contrasts with the much shallower oxidation profile in the western portion of the mining operation.

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Figure 11.3 Çöpler Drill Core Showing Colour Change from Oxide to Sulfide

Anagold, 2016

11.1.15 Çöpler Model Validation

Model validation was approached in several ways:

•The estimated Au grades in the model were compared to the composite grades by visual inspection in plan views, north–south cross-sections, and east–west cross-sections. In general, the model and composite grades compared well visually.

•The cell model was checked for global bias by comparing the mean Au, Ag, Cu, and S grades (with no cut-off) from the model (OK/ID2 grades) with means from NN estimates for cells of Indicated classification. The NN estimator produces a theoretically unbiased (de clustered) estimate of the mean value when no cut-off grade is imposed and provides a reasonable basis for checking the performance of different estimation methods. In general, an estimate is considered acceptable if the bias is at or below 5% (relative difference).

•Local trends in the grade estimates (also known as drift analysis) were assessed by plotting the mean values from the NN estimate versus the kriged results for Indicated model cells in east–west, north–south and vertical directions (swath plots). The global comparisons agree well, however the swath plots do illustrate the existence of slight local differences between the NN and kriged model grades.

11.1.16 Çöpler Mineral Resource Classification

Grade estimates were classified using the following Anagold guidelines:

•Indicated Mineral Resource should be quantified within relative ±15% with 90% confidence on an annual basis, and

•Measured Mineral Resources should be known within ±15% with 90% confidence on a quarterly basis.

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Based on these guidelines, the drilling is generally sufficiently close-spaced enough to permit confirmation of or assumption of continuity (Measured vs. Indicated, respectively) between data points. For the Çöpler model, a drillhole spacing study was performed to determine the nominal drillhole spacing required to classify material as Indicated.

Confidence limits were calculated on a single block that represents one month of POX production (based on 1.9 Mtpa). The confidence limits, a review of continuity on sections and plans, and an assessment of data quality were used to determine minimum drillhole spacing by domain. A spacing of 40 m x 40 m in the Marble Zone, 50 m x 50 m in the Manganese Zone and West Zones, and 60 m x 60 m in the Main Zone was required to meet the requirements for Indicated. An 80 m x 80 m spacing was required for Inferred in all domains. Model Cells with a drillhole spacing that was greater than 80 m were not classified as Mineral Resource.

The resultant classification was then ‘smoothed’ to remove the isolated cells that are not of the same classification tenor as the proximal surrounding cells.

The resulting classification shows that much of the deposit can be classified as Indicated, with Inferred cells forming a halo around the Indicated mineralisation Figure 11.4. A small quantity of cells classified as Measured.

Figure 11.4 Projected Plan View of Çöpler Resource Classification

OreWin, 2020

Only model cells with Au >0.3 g/t shown

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11.1.17 Çöpler Model Validation

Model validation was approached in several ways:

•Local trends in the grade estimates (also known as drift analysis) were assessed by plotting the mean values from the NN estimate versus the kriged results for Indicated model cells in east–west, north–south and vertical directions (swath plots). The global comparisons agree well, however swath plots illustrate the existence of slight local differences between the NN and kriged model grades.

11.1.18 Çöpler Assessment of Reasonable Prospects of Eventual Economic Extraction

Refer to Section 11.5 for the assessments for Mineral Resource estimates meeting reasonable prospects for eventual economic extraction.

11.1.19 Çöpler Deposit Mineral Resource Tabulation

Mineral Resources are reported exclusive of Mineral Reserves in Table 11.57 according to Mineral Resource classification and material type. Mineral Resources are presented showing only the SSR attributable proportion. Mineral Resources that are not Mineral Reserves have not demonstrated economic viability. The pit shell used to constrain the resource has been updated to reflect the increase in gold price. Depletion from mining has been included.

11.2 Çakmaktepe

The Çöpler district hosts various styles of mineralisation, mainly epithermal, skarn and contact style gold and gold–copper mineralisation. The Çakmaktepe North zone of the Çakmaktepe deposit is a strongly sheared zone with strong epithermal characteristics and grade associations with intrusive diorite dykes. As with the other prospects the mineral association is dominantly Au–Cu–Ag. Other mineralised zones belonging to the Çakmaktepe deposit are generally contact styles of mineralisation where Au–Cu–Ag have been emplaced along thrust surfaces next to ophiolite, limestone, and metasediment. Epithermal veining and replacement alteration textures are prevalent.

Oxide mining began in the Çakmaktepe Central and East pits in November 2018. Mining continued through September 2019 within the same two pits. Oxide ore material was transported to the Çöpler oxide processing facility for inclusion on the heap leach pad.

A geological model was constructed along with a cell model estimating grades for Au, Cu, Ag, S, and C. Estimated grades were constrained by mineralised envelopes.

11.2.1 Çakmaktepe Domains

At Çakmaktepe, mineralisation follows structural controls and designated lithological contact orientations. Grades trends and element associations were investigated, and several separate domains were identified and are shown in Figure 11.5.

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Mineralisation at Çakmaktepe often overlaps multiple lithological units along its boundary, rather than being hosted within a single rock type. For this reason, grade shells were constructed for gold and copper to constrain estimates within mineralised zones. The mineralised shapes for gold and copper are lenticular with thicknesses ranging from 5–40 m, the average thicknesses being approximately 6 m.

Grade shells were also developed for silver. However, because silver mineralisation tends to be more dispersed and more difficult to follow across the deposit than gold, different methods were used for silver grade shells depending on which area was being modelled.

Sulfur grades follow lithological units. Higher S values are seen in diorite and metasediment, with decreased S in gossan, jasperoid, ophiolite, and marble.

The key points in relation to Çakmaktepe mineralisation domains are:

•Çakmaktepe North is located on a vertical shear structure with elevated metal grades within jasperoid unit. Several low-angle structures dipping to the north-east carry grades along the marble to metasediment contact. Intrusive diorite/s, orientated vertically, crosscut all other lithological units. Mineralisation within/around the diorite is limited in Çakmaktepe North.

•Çakmaktepe Central mineralisation follows the marble contact, which dips gradually to the north-east. The marble unit is approximately 15 m thick and located between the ophiolite and metasediment units.

•Mineralisation in the Çakmaktepe East area is near-surface and within the gossan unit, which is relatively flat lying and localised.

•The South-east area seems to be controlled by a massive diorite body with gossan at the surface. Mineralisation is weak and near-surface.

Figure 11.5 Çakmaktepe Model Domains (oblique view)

Anagold, 2020

Contacts for lithological shapes used the raw logged interval depth in 3D space. Surfaces were generated through implicit modelling of contact locations in the drillholes.

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Construction of the lithological shapes assumed the following:

•Diorites are intrusive units that can exist as large bodies or thin sills cross-cutting other units.

•Jasperoid is an alteration product but treated here as a lithological unit. Jasperoids occur along shear zones and are high in pyrite. Jasperoid can exist in pods and can be discordant to surrounding stratigraphy.

•Gossan is primarily the result of surficial oxidation, with the shape influenced by the local topographic elevation.

•In most areas, marble overlays metasediment, with ophiolite above marble.

•Offsets in lithological units help to define fault locations and structural boundaries.

A series of fault surface wireframes were developed to represent the structural knowledge at Çakmaktepe. These structures extended beyond the Çakmaktepe model area to take into consideration the spatial relationships between Çakmaktepe and Ardich. The incorporation of modelled 3D faults into the geological model highlighted a discrepancy between the Ardich lithological concept and the Çakmaktepe geological units. Given the correlation of the two deposits was not clearly defined at the time of this model, interpreted faults were excluded from the Çakmaktepe geological model.

11.2.2 Çakmaktepe Data Summary

The cut-off date for the export of the drillholes from the database to be used in the resource modelling was 31 October 2019. The extract contained 1,109 drillholes with a drilling date range of September 2007–October 2019, totalling of 119,001 m of drilling.

11.2.2.1 Çakmaktepe Drillhole Compositing

The original sample lengths in the Çakmaktepe dataset are predominately 1 m, with some 2 m sampling through zones presumed at the time of drilling to be waste. The average sample length is 1.02 m. The shortest interval was 0.1 m, and the maximum length was 3.1 m.

Samples were composited to 5 m lengths for use in statistical analysis and construction of mineralisation boundaries. Often, composites along lithological boundaries were selected to match geological control with mineralisation.

Composites were then flagged within the mineralisation shapes. Lithology is also coded into the composite file based on the interpreted shapes.

11.2.3 Çakmaktepe Exploratory Data Analysis

Detailed exploratory data analysis (EDA) was conducted on the Çakmaktepe resource modelling dataset. This is discussed in detail in the CDMP20TR. A summary of findings of the statistical analyses follows.

11.2.3.1 Çakmaktepe Statistical Summary

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A summary of key findings follows:

•Box plots confirm observations made from histograms and probability plots that gossan and jasperoid contain significantly higher Au grades and the remaining units (diorite, metasediment, ophiolite, and marble) have lower Au grades.

•Box plots of sulfur show higher sulfur content in diorite and metasediment with moderate sulfur grades in gossan and jasperoid. Low sulfur is consistently seen in ophiolite and marble. For this reason, the sulfur estimate uses lithologic contacts as domain boundaries.

•Mineralisation tends to spatially follow lithological contacts.

•For Çakmaktepe Central, the probability plot is relatively straight, indicating only one population is present in the distribution.

•Core recoveries are between 80%–90%, reflecting strongly sheared, brecciated, altered and in areas of limestone, karstic ground (cavities) being drilled at Çakmaktepe.

•Contact plots were created to show grades change across geological boundaries. Jasperoid and gossan are favourable mineralisation hosts and show abrupt grade changes when compared to the other lithologies (marble, metasediment, diorite, ophiolite).

Grade shell boundaries were constructed to follow lithological contacts and were used as hard domains in the grade estimation process.

11.2.4 Çakmaktepe Top Cutting

Top cuts were selected based on the log probability plot, supported by the Projection of the data trend to the expected upper grade (Y-axis value) using the top sample value curve (Table 11.6). Top cutting occurred after compositing to 5 m. A spatial review of top values by domain shows randomly spaced samples rather than a localised body of higher grades.

High-yield limits were included outside of the grade shells to restrict the extrapolation of higher grades within the applied search distance. For Au, sample values above 4 g/t Au were restricted to a distance of 10 m x 10 m x 5 m in the East and South-east areas. For Central, a high-yield limit of 8 g/t Au was used. The high-yield limit was increased to 12 g/t Au in the Çakmaktepe North area. For copper, samples above 2% Cu were restricted to 10 m x 10 m x 5 m in Central and 3% Cu in North and East.

Table 11.6 Çakmaktepe Top Cuts for Au, Cu, and Ag

Element		Çakmaktepe Area		Top Cut Grade		No. Samples Cut
Au<br>(g/t)		North		15.0		2
	Central		9.0		7
	East		5.5		1
	South-east		5.0		4
Cu<br>(%)		North		4.0		2
	Central		3.0		2
	East		4.0		2
	South-east		1.0		2
Ag<br>(g/t)		North		180		2
	Central		130		3
	East		150		5
	South-east		60		5

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11.2.5 Çakmaktepe Resource Model Estimation

11.2.5.1 Çakmaktepe Cell Model

A cell model was constructed by first coding the interpreted lithology shapes into the cells. These were then flagged by each of the grade shells and model domains. A project-wide solid was used to trim out distant cells at model edges.

The cell model limits are shown in Table 11.7.

The model was not rotated, and no sub-celling was used.

Table 11.7 Çakmaktepe Cell Model Parameters

Direction	Minimum<br>(m)	Maximum<br>(m)	Range<br>(m)	Cell Size<br>(m)	No. of Cells
East	463,400	465,700	2,300	5	460
North	4,364,800	4,366,700	1,900	5	380
RL	1,050	1,850	800	5	160

11.2.5.2 Çakmaktepe Estimation Method

Au, Ag, Cu, S, and C were interpolated using ID3 and NN methods. Au, Cu, and Ag were estimated according to grade shell constraints. S and C were estimated by modelled lithological units. All grade shell boundaries were treated as hard. Mineralisation domains were treated as soft boundaries allowing the selection of samples from nearby domains.

A single search distance was used within the Au, Cu, and Ag grade shells. A two pass method was used to estimate cells outside of the grade shells. Search ranges and sample requirements varied by estimation pass. Search orientations were selected to match the mineralised dip and dip-direction.

Au was interpolated within each gold grade shell using only composite samples inside the shell. Au grade was then interpolated into cells outside the grade shell using domain-specific parameters. Cu was estimated using the same method as Au, by first interpolating grade within the copper grade shells and then interpolating outside the grade shells in two passes. Ag estimation followed the same technique as Au and Cu by interpolating within the silver grade shell and then interpolating outside the grade shell by domain.

Sulfur and carbon content is linked to lithology. Lithological shapes were used as hard boundaries to interpolate S and C grades. No preferred orientation of S or C grades was observed; therefore, a spherical search was used.

A NN estimate was completed for all variables using the same composites, same domains, same search ranges and same top cut values as the ID3 estimates. The resulting NN model was used for estimation validation to detect potential estimation bias by domain.

11.2.6 Çakmaktepe Density Model

Density measurements were collected on DD core samples spaced nominally 3 m apart down-hole. Density values were statistically analysed by lithology with outliers and non-representative values excluded from the analysis.

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A review of histograms of density within each rock type aided in the selection of bottom and top cut values.

Selected lower and upper cut density values by lithology are shown in Table 11.8.

Table 11.8 Çakmaktepe Upper and Lower Density Limits by Lithology

Lithology	No. of Density Data	Bottom Cut	Top Cut
Cataclasite	33	2.60	2.62
Diorite	1,496	2.00	3.00
Gossan	407	2.00	2.90
Jasperoid	1,972	2.00	3.20
Listwanite	29	2.28	2.80
Marble	4,041	1.91	3.50
Metasediment	4,114	2.00	3.30
Ophiolite	3,400	2.00	3.00

Densities used in the resource model are summarised in Table 11.9.

Table 11.9 Density Values Assigned to the Çakmaktepe Cell Model by Lithology

Lithology	No. Density Data	Assigned Density<br>(t/m3)
Gossan	389	2.48
Jasperoid	1,755	2.60
Diorite	1,186	2.54
Ophiolite	2,870	2.41
Metasediment	3,666	2.65
Marble	3,533	2.64

11.2.7 Çakmaktepe Model Validation

Validation of the 5 m x 5 m x 5 m model estimates included visual inspection of grade estimates, comparisons of cell grades to drillhole data, checks for global bias, check of local bias (swath plot), metal reduction calculation, and comparison of estimates within the Central and East areas using grade tonnage curves.

Visual inspection of plans and sections and 3D visualisation confirmed that the cell model estimates honour the drillhole data and grade shell boundaries.

11.2.8 Çakmaktepe Comparison to Production Data

Mining occurred in the Çakmaktepe Central and East pits, primarily during 2019. Blasthole data from these two pits were used to construct an Au grade estimate for comparison of the production model to the Mineral Resource model.

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The production model, using blasthole assay data, was set up to follow the same parent cell size used in the resource model of 5 m x 5 m x 5 m. This generates cell centroids with the same centroid coordinates as the resource model for relational comparisons by cell.

Plotting the grade / tonnage curve for Au shows the number of tonnes to be similar in both models, with a crossover of the resource to production model tonnes occurring between the 0.8–2.8 g/t Au cut-offs. A large variance is seen when comparing Au grades between the two models. The increased grade in the production model results in more gold ounces. The largest positive and negative variances between the two models were investigated. The following observations were made:

•Estimate variances exist throughout the two cell models. An overall bias towards higher grade blastholes results in higher cell grades in the production model.

•Comparison of cut-off grades shows a larger variance in gold ounces between the two models as the cut-off grade is increased. Variances were plotted on a grade / tonnage curve by pit for a comparison of gold ounces by area.

•Variances were not limited to specific locations. Positive and negative variances were mixed throughout the Central and East pit. This suggests the selected modelling method for the resource grade estimation is not bias high or low, but likely producing a gold model more generalised than the variability seen within the deposit.

•When using the ID3 interpolation method, cell grades closely match drillhole composite values. Investigation of areas where exploration drilling crosses cells shows lower estimated grades in the resource model and higher estimated grades in the production model. This illustrates the variance in the two drillhole datasets – exploration to blasthole data.

These observations indicate that the variances between the two datasets are likely greater than the software tools available to match the deposit grade distribution and short range variability to the resource model. To compensate for the model variances, increasing the exploration drill density to the deposit variability is preferred. However, increasing the exploration drill density is probably not feasible due to the high inherent variance seen in the deposit. This presents a risk that mining may not match the predictive abilities of the resource model using the available exploration data.

11.2.9 Çakmaktepe Mineral Resource Classification

In summary, assignment of model classification followed these steps:

•Sample spacing was calculated based on samples from drillholes containing assay values. The calculation of sample spacing did not use limiting boundaries such as domains or lithological shapes.

•Inferred and indicated classification was assigned based on drill sample distances (20 m and 35 m).

•Indicated classification was then restricted to those cells within the modelled mineral grade shells for gold, copper, and silver.

•South-east estimates were downgraded to Inferred.

Grade estimates within the grade shells were visually confirmed by comparing the grade of the cell with the grade shell boundary. Higher grades exist inside the grade shell with a drop in grade tenor evident when crossing the grade shell boundary. Grade shells follow geological features such as lithological contacts and the Çakmaktepe North shear structure. Estimates outside of the grade shells were set to generalised orientations honouring the trends of the low-grade mineralisation and orientations of the major lithological units.

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Hard grade boundaries were used for gold, silver, and copper. The sharp changes in grade are expected, rather than being an artefact of the estimate, due to the close relationship between mineralisation and structural features. This relationship is supported by close-spaced drilling throughout Çakmaktepe and crossing holes in areas such as the shear zone in Çakmaktepe North.

11.2.10 Çakmaktepe Assessment of Reasonable Prospects of Eventual Economic Extraction

Refer to Section 11.5 for the assessments for Mineral Resource estimates meeting reasonable prospects for eventual economic extraction.

11.2.11 Çakmaktepe Mineral Resource Tabulation

Çakmaktepe Mineral Resources are reported exclusive of Mineral Reserves and have been tabulated by Mineral Resource classification and oxidation state in Table 11.57. Mineral Resources are presented showing only the SSR attributable proportion. Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

11.3 Ardich

The latest mineral asset to be intensively studied in the Çöpler district suite of mineralised zones is Ardich, which is located approximately 6 km east of the current Çöpler pit and 1 km north of the Çakmaktepe pits. The Ardich deposit is accessed by the İliç-Yakuplu village road, which is open throughout the year.

Ardich mineralisation was discovered in August 2017. Ardich does not appear to have hosted historical mining or trenching in the way that Çöpler and Çakmaktepe have.

The local geology at Ardich is dominated by ophiolites, listwanites, dolomites, cataclasites and limestones, with lesser amounts of jasperoid and diorite Figure 11.6. This lithology assemblage occurs within a complex north-west trending structural zone that is cut by multiple high-angle faults, which together result in multiple rotated fault blocks and mineralised zones.

The jasperoids and diorites form a volumetrically minor part of the lithology sequence, however, they appear to be the most important controls of the mineralisation at Ardich.

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Figure 11.6 Ardich Geology Schematic (long-section)

Anagold, 2021

11.3.1 Ardich Geological Model

The 2021 Ardich model update is underpinned by the structural and lithological interpretations developed for the previous geological modelling at Ardich.

Overall, the lithological and structural framework for the geological model remains conceptually similar to the previous interpretations. However, the July 2021 update includes six additional faults, which dissect the deposit into more structural domains than in previous modelling.

Standalone lithological interpretations were developed for each structural domain. This was necessary to enable the offsets / discontinuities and changes in behaviour of the lithological units in each structural domain to be represented appropriately.

11.3.2 Ardich Structural Interpretation

There are two dominant structural trends interpreted at Ardich:

1. North-west / south-east trending faults form the primary structural features.

2. A series of smaller, less pervasive secondary faults crosscut the primary structures in a north-east / south-west trend.

The combination of these faults creates unique structural blocks that have moved vertically relative to each other as well as pivoted / rotated. The nomenclature for, and characteristics of, each of the interpreted faults is listed in Table 11.10.

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The location and attitude of the faults was largely determined from incongruous lithological occurrences observed down drillholes, further enhanced at times by logged commentary regarding evidence for faulting in the lithological dataset.

Table 11.10 Interpreted Fault Trends

Fault Number	Trending Direction	Fault Dominance
1, 5, 8, 9	North-west / South-east	Primary
7, 10, 11, 16	North-west / South-east	Secondary
2	East / West	Primary
3, 6	North-east / South-west	Primary
4, 12, 13, 14, 15	North-east / South-west	Secondary

The delineation of the deposit by the fault wireframes allowed for the creation of a domain field named ‘FAULTZON’ in each drillhole sample and model cell to identify its location within the structural framework. The interactions of the faults resulted in 25 unique FAULTZON domains. The location and inter-relationships of the faults and the FAULTZONs are shown in plan view in Figure 11.7.

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Figure 11.7 Ardich Structural Framework: Interpreted Faults and Resultant FAULTZONs

OreWin, 2021

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11.3.3 Ardich Lithological Interpretation

The typical stratigraphic sequence encountered at Ardich is (from surface down): ophiolite, listwanite, dolomite, cataclasite, and limestone, with jasperoid and diorite variably present within any of these strata. An example cross section is shown in Figure 11.8.

In addition to this typical sequence, secondary repeat occurrences of ophiolite, listwanite, dolomite, and cataclasite can be present in the stratigraphic sequence in some FAULTZONs. A silica cap unit is present in FAULTZON 6 (SE).

The jasperoid and diorite lithologies have a relatively small volumetric presence and were formed by different geological mechanisms compared to the more extensive (meta)sedimentary stratigraphic host lithologies (ophiolite, dolomite, cataclasite, hornfels, and limestone).

Wireframe surfaces and solids were created to represent the interpreted nature of the main lithologies. The field ‘MODLITH’ was created in the sample and model files to identify the lithology domain that the sample or cell represents.

An additional field ‘LITHNUM’ was also created in the sample and model files to provide a numerical identifier for each lithology to differentiate between the repeat occurrences of the same lithological units within a FAULTZON.

Figure 11.8 Schematic Example Cross-Section of Ardich Lithology Model (looking north-west)

OreWin, 2021

11.3.4 Ardich Mineralisation

The mineralisation at Ardich is related to crystalline and chalcedonic quartz veins within the brecciated and silicified jasperoid, listwanite, and dolomite zones. The mineralisation is predominantly in the form of oxide, with sulfide mineralisation confined to limited pyrite-rich jasperoid zones.

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The mineralisation is considered to be related to fluids associated with diorite intrusions. This manifests as either direct contact mineralisation surrounding the diorites or as diorite-derived mineralising fluids that have migrated along lithological or fault contacts.

Gold grades increase at jasperoid / dolomite / listwanite contacts and within the silica-rich listwanites, which act as horizontal traps for higher grade gold-bearing mineralisation. Increases in gold grade often occur along the lithological contacts. A rapid down-hole change in gold grade tenor, notionally from mineralised to unmineralised material, can be seen in many drillholes, indicating that the mineralisation is tightly constrained within the controlling features rather than generally disseminated across the deposit.

Four distinct styles of mineralisation have been observed:

•Jasperoid-related mineralisation

In this most-prevalent style of mineralisation at Ardich, gold mineralisation can either occur throughout the entirety of a jasperoid unit or display a concentration towards the middle-to-lower part of the unit. The formation of the jasperoid, and the gold mineralisation more broadly, is interpreted to be related to the fluids associated with diorite intrusions.

The diorites from which jasperising and mineralising fluids have been interpreted to originate can be located either within a single FAUTLZON or in an adjacent FAULTZON to the mineralised jasperoid, with the listwanite / dolomite contact providing a low-resistance fluid pathway between the diorite and the jasperoid.

•Diorite contact-related mineralisation

The second most common style of mineralisation is related directly to the contact of a diorite intrusion with the host strata; most-commonly within the listwanite and in the absence of jasperoid.

This style of mineralisation tends to be parallel to the diorite intrusion, but generally limited to within approximately 20 m of the diorite contact.

•Contact-parallel mineralisation

Both the well-mineralised jasperoid / listwanite domains and the less well-mineralised dolomites show contact-parallel mineralisation.

In the jasperoids, this often manifests as higher grade results (within an overall well-mineralised zone) occurring at a similar distance from the contact (i.e., contact-parallel).

The mineralisation is likely related to a sedimentary / depositional feature such as bedding within the unit. Like the lithological contact being preferentially used as a low resistance fluid pathway for the mineralising fluids, internal variation within the units may offer additional low resistant pathways for the fluids.

•Complex domains mineralisation

There are a small number of FAUTLZONs that are more structurally and lithologically complex than the majority of the FAULTZONs in the deposit. The mineralisation within these complex domains does still appear to adhere to the jasperoid-related, contact-parallel, and diorite contact-related mineralisation models. However, the number of lithology contacts and the geometry of these units appears to be more complex than in other parts of the deposit.

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11.3.5 Ardich Database Extract

A MS Access database ’Ardich_12072021.mdb’ (the July database) was supplied to OreWin on 13 July 2021. The database contained collar, survey, assay, lithology, and density data for drilling completed up to 29 May 2021. The tables contained within the database and used in this study are:

•tblDHColl

•tblDHSurv

•tblDHLithology2017

•tblVWDHAssays_ALL

•tblDHSpecGrav

The July database tables were exported to comma delimited text format (csv) then imported in to Datamine software for validation and to create the working drillhole files.

The tables were reviewed extensively to identify any erroneous or unusual data. A comparison was also made between the July database and the previous database extract.

11.3.6 Ardich 2021 Resource Modelling Dataset Summary

A total of 427 diamond drillholes (DD) have been drilled at Ardich since late-2017, (see Table 11.11 and Figure 11.9). Exploration drilling at Ardich utilised surface PQ and HQ triple tube diamond core drilling. No RC drilling has occurred to date at Ardich.

After the initial discovery of the Ardich deposit, DD drilling programmes have continued to improve confidence in the interpretation.

Table 11.11 Ardich Drilling History in 2021 Resource Modelling Dataset

Year	Number of Drillholes	Drilled Metres
2017	9	1,374.10
2018	91	14,216.40
2019	133	27,821.20
2020	147	35,146.65
to 29 May 2021	45	8,479.90
Total	427	87,038.25

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Figure 11.9 Ardich Drill Collar Location Plan

OreWin, 2021

11.3.7 Ardich Exploratory Data Analysis

In-depth analysis of the grade distribution and continuity of the assay data was undertaken both visually and statistically.

11.3.7.1 Ardich Summary Statistics

The raw statistics of the key elements in the global dataset are shown in Table 11.12.

Table 11.12 Key Element Statistics of Uncomposited Drillhole Data (length weighted)

Element	No. Samples	Min.	Max.	Mean	Std. Dev.	Variance	CV
Au (ppm)	72,432	0.0025	44.7	0.26	1.12	1.26	4.25
Ag (ppm)	72,432	0.25	246	0.68	2.95	8.67	4.32
Cu (%)	72,432	0.00005	19.4	0.005	0.11	0.01	22.87
S (%)	72,432	0.01	32.5	0.52	1.23	1.50	2.34

Statistics were generated for all lithologies present at Ardich.

It can be observed from the summary statistics that the mean grades, the number of samples and the range of grades varies significantly across the lithologies and FAULTZONs.

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Histograms of Au data at the domain level show that most domains have a lognormal grade distribution. Also seen in the histograms for many of the domains is evidence of mixed grade populations. The presence of higher grade samples proximal to the lithology contacts provides a geological explanation to the mixed grade distributions seen in the some of the histograms.

Sulfur content is more strongly related to the lithological unit than the gold mineralisation. While some increase in sulfur grades was observed proximal to the lithological contacts, this phenomenon was not of the same tenor or uniformity as is evident in the gold mineralisation.

Gold Statistics

As can be seen in Table 11.13 and Figure 11.10, the highest mean Au grades are in the jasperoid and listwanite.

Table 11.13 Au Summary Statistics by Lithology based on Uncomposited Samples (length weighted)

Lithology		No. of Samples		% of all Samples	Au (ppm)
	Min.		Max.		Mean
Ophiolite		29,425		41%	0.0025	3.7	0.01
Listwanite		8,723		12%	0.0025	25.4	0.86
Dolomite		15,070		21%	0.0025	24.1	0.36
Cataclasite		8,990		12%	0.0025	30.3	0.16
Jasperoid		2,058		3%	0.0025	44.7	3.21
Limestone		1,852		3%	0.0025	1.5	0.01
Diorite		2,203		3%	0.0025	11.9	0.25
Hornfels		3,447		5%	0.0025	9.3	0.15
Silica Cap		664		1%	0.0025	0.8	0.02
ALL		72,432		100%	0.0025	44.7	0.31

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Figure 11.10 AU_PPM Box and Whisker Plot by Lithology (MODLITH)

OreWin, 2021

A detailed visual assessment of the Au grade distribution showed a preference for higher Au grades to be concentrated on or near a lithological contact. The following observations were made:

•For the jasperoid and listwanite, the higher Au grades were predominately located on or near the lower lithological contact. In addition to the preferential occurrence of grades proximal to contacts, a grade trend with a contact-parallel orientation was also noted.

•For the dolomite and cataclasite, higher Au grades were associated with the upper lithological contact. These lithologies also displayed areas of contact-parallel mineralisation within the main body of the unit.

•Higher Au grades were also noted in the host rock surrounding some, but not all, of the diorite intrusions and occasionally proximal to the faults. Consistent with the mineralisation observed on the main lithology contacts, there appears to be a contact-parallel nature to the mineralisation surrounding the diorites.

•Areas of strongest diorite intrusion often coincide with the areas of greatest jasperisation and gold mineralisation. The presence of the highest Au grades proximal to the lithological contacts suggests the preference of these contact boundaries as a low resistance pathway for the diorite derived mineralising fluids.

•Gold mineralisation in the listwanite is most-commonly located in the lower part of the unit, occurring on either the listwanite / jasperoid, or listwanite / dolomite contact.

•Gold mineralisation in the dolomite is most-commonly located in the upper part of the unit, close to the dolomite / jasperoid or dolomite / listwanite contact. Where the listwanite, jasperoid, and dolomite are all mineralised within a single FAULTZON, the Au grade tenor is often (but not always) highest within the jasperoid.

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Differences in Au grade were observed between the different lithologies at the deposit scale. In addition to this, there were also noticeable differences in Au grade for the same lithologies in different FAULTZONs. As an example, Figure 11.11 shows the range and mean Au grades of the jasperoid for the various FAULTZONs.

This variation in Au grade between FAULTZONs is illustrated in Figure 11.12, which shows a very large difference in grade tenor and location in the jasperoid in AR361 in FAULTZON 10.2 (S1) relative to the jasperoid in AR416 in FAULTZON 10.2 (S3). The jasperoid hosts only 3% of the total samples collected.

Figure 11.11 Jasperoid AU_PPM Box and Whisker Plot by FAULTZON

OreWin, 2021

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Figure 11.12 Example Cross-Section (looking north-west) through FAULTZONs 19.1 and 10.2 showing AU_PPM Differences in Adjacent FAULTZONs

OreWin, 2021

Both the visual and statistical observations of the Au grade distribution in the different FAULTZON blocks informed the decision to estimate each lithology within each FAULTZON independently.

Sulfur Statistics

Table 11.14 and Figure 11.13 show the shows the basic statistics for sulfur in each lithology at the deposit scale. This shows that the lithology with the highest mean S grade is the jasperoid (3.3% S). There is a strong correlation between lithology and sulfur content.

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Table 11.14 Sulfur Summary Statistics by Lithology based on Length Weighted Uncomposited Samples

MODLITH		No. of Samples		% of all Samples	S (%)
	Min.		Max.		Mean
Ophiolite		29,425		41%	0.005	8.5	0.16
Listwanite		8,723		12%	0.005	21.9	0.52
Dolomite		15,070		21%	0.005	20.5	0.28
Cataclasite		8,990		12%	0.005	19.6	1.44
Jasperoid		2,058		3%	0.005	32.5	3.30
Limestone		1,852		3%	0.005	4.6	0.11
Diorite		2,203		3%	0.005	18.9	1.67
Hornfels		3,447		5%	0.005	29.3	1.52
Silica Cap		664		1%	0.005	1.6	0.03
ALL		72,432		100%	0.005	32.5	0.52

Figure 11.13 S_PCT Box and Whisker Plot by MODLITH

OreWin, 2021

11.3.8 Ardich Core Recovery

Exploration drilling at Ardich utilised surface PQ and HQ triple tube diamond core drilling.

Overall, Ardich drill core recovery is good with a mean recovery over 92%.

Uncomposited data statistics were compared to core recovery collected during geotechnical logging. No correlation is seen between Au grade and core recovery.

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11.3.9 Ardich Top Cutting

Statistical analysis, using log. probability plots, mean variance plots, log. histograms, and percentage change statistics, was undertaken for each of the domains. Once identified, apparently outlying samples were reviewed visually in 3D to determine whether there was sufficient local support to allow the sample to retain its un-cut grade, or if grade cutting was required. Often, samples that appeared as statistical outliers at the domain scale transpired to be well-supported at the local scale and conformed to the mineralisation model of higher grades located proximal to lithological or fault contacts, indicating that top cutting may not be justified.

This review process resulted in top cuts being applied to only 41 samples from 24 domains. The domains where top cutting was applied are summarised in Table 11.15. This table shows the effect of the top cutting process on the statistics of the cut samples. Sulfur grades were also assessed for potential top cutting with no top cuts applied

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Table 11.15 Au Top Cut Values and Effect on Statistics

LITHNUM		Lithology		FAULTZON		No. Samples				Top Cut Value			Top Cut Percentile		Mean <br>(Au ppm)				CV<br> (Au)
	Total		Cut			Uncut		Cut			Uncut			Cut
2		Listwanite		8.2		256		1		8.0			99.6%		0.88		0.86		1.73	1.57
	9		491		2		9.5		99.6%			0.61		0.58		2.36		2.01
	19.2		93		3		1.2		96.7%			0.22		0.14		3.20		2.19
3		Dolomite (upper)		8.2		1144		3		6.0			99.7%		0.28		0.27		2.87	2.26
	13.2		66		1		4.4		98.4%			0.69		0.62		2.01		1.65
	14		40		2		1.3		94.9%			0.34		0.21		2.54		1.63
	29.1		240		1		7.0		99.5%			0.74		0.68		2.19		1.41
3.1		Dolomite (lower)		51		292		1		2.0			99.7%		0.07		0.06		3.95	2.71
4		Cataclasite (upper)		2		503		2		12.0			99.6%		0.54		0.50		3.81	3.40
	10.1		713		1		5.0		99.8%			0.17		0.16		3.38		3.06
4.1		Cataclasite (lower)		13.1		269		2		0.3			99.2%		0.04		0.03		3.94	1.89
	29.2		33		2		0.4		93.8%			0.21		0.14		1.70		0.99
5.1		Jasperoid		2.1		296		3		24.0			99.0%		6.73		6.60		0.79	0.71
	8.1		221		3		12.0		99.0%			2.25		2.16		1.29		1.13
	13.2		23		2		9.0		91.0%			5.52		4.45		0.98		0.52
	14		107		2		6.5		98.1%			1.19		1.13		1.35		1.18
	19.1		9		1		1.0		88.3%			0.85		0.55		1.18		0.45
	29.3		15		1		3.5		92.9%			2.19		2.01		0.59		0.41
5.2		Jasperoid		19.1		197		2		18.0			98.9%		3.90		3.77		1.15	1.01
7.2		Diorite		25		28		1		2.0			97.2%		0.52		0.44		1.51	0.97
7.3		Diorite		8.1		64		1		2.0			98.7%		0.30		0.26		2.13	1.37
	13.1		11		1		2.0		90.0%			1.40		0.84		1.63		1.00
7.4		Diorite		8.2		48		1		0.7			98.1%		0.18		0.16		1.46	1.08
7.5		Diorite		51		73		2		7.0			97.3%		2.17		2.05		0.96	0.78

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11.3.10 Ardich Drillhole Compositing

A 1 m sample interval was the most prevalent raw sample length in the assay database, accounting for 76.95% (55,733 samples) of all samples. The next most-prevalent sample length was 2 m, accounting for 11.17% (8,090 samples) of all samples. See Figure 11.14.

Figure 11.14 Histogram of Raw Sample Lengths

OreWin, 2021

Drillholes containing samples of 2 m length are located within almost every FAULTZON. However, the location of these samples was heavily biased to the ophiolite lithology, with 99.25% (8,029 of 8,090 samples) of the 2 m samples located in ophiolite.

In addition, there is a temporal bias in the collection of the 2 m sample lengths with all but three (3) of the 2 m samples being collected from hole AR194 and onwards. This indicates an understanding of the lithology / mineralisation relationship developed over time, for example, it became understood that the ophiolite contained little or no mineralisation and a 2 m sample interval could be used in this material without compromising the sample support of the mineralised areas.

After due consideration of the sample length analyses, compositing to a length of 1 m was determined to be optimal as this length would preserve the integrity of the majority of samples in the raw database, with minimal compromise to the modified samples.

During compositing, samples were not permitted to composite across any of the boundaries defined by the FAULTZON or LITHNUM attributes. Compositing to 1 m had a very little impact on the statistics of the data. Table 11.16 and Table 11.17 shows the effect of the compositing process for the jasperoid and listwanite within each FAULTZON area. The full summary of the effect of compositing across all domains can be found in Table 11.16.

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Table 11.16 Drillhole Au Statistics for Uncomposited (Raw) and Composited (Comp.) Sample Data – Jasperoid

LITHNUM		FAULTZON			No. Samples						Max. Length (m)						Max. Grade <br>(Au ppm)						Mean Grade<br> (Au ppm)						CV<br> (Au)
	Raw			Comp.			Raw			Comp.			Raw			Comp.			Raw			Comp.			Raw			Comp.
5.1		1			74			75			1.5			1.0			8.69			8.69			2.04			2.05			0.98		0.95
	2		161			166			1.6			1.0			14.90			14.90			3.04			3.05			0.91			0.88
	2.1		292			296			1.8			1.0			44.70			42.00			6.74			6.73			0.82			0.79
	8.1		218			221			1.6			1.0			22.90			22.90			2.26			2.25			1.34			1.29
	8.3		23			23			1.3			1.0			8.99			8.99			3.75			3.75			0.72			0.73
	9		119			121			1.8			1.0			6.32			6.32			1.53			1.54			0.96			0.93
	10.1		92			94			1.5			1.0			24.50			23.15			3.90			3.86			1.11			1.09
	10.2		35			36			2.2			1.0			2.70			2.70			0.86			0.86			0.93			0.89
	13.2		21			23			1.7			1.0			26.20			26.20			5.26			5.52			1.04			0.98
	14		104			107			1.6			1.0			10.60			10.60			1.20			1.19			1.42			1.35
	19.1		9			9			1.0			1.0			3.63			3.63			0.85			0.85			1.19			1.19
	19.2		208			216			2.9			1.0			20.30			16.15			2.91			2.95			1.06			1.01
	29		82			84			1.5			1.0			10.50			10.50			3.00			3.01			0.80			0.78
	29.1		122			122			1.4			1.0			8.67			8.67			1.86			1.88			0.87			0.87
	29.3		16			15			1.6			1.0			6.15			6.15			2.18			2.19			0.58			0.59
	51		64			64			1.1			1.0			5.84			5.84			1.10			1.11			1.11			1.11
5.2		1			28			28			1.2			1.0			6.37			6.37			2.64			2.65			0.58		0.59
	10.1		13			13			1.4			1.0			9.40			9.40			3.18			3.08			0.82			0.86
	19.1		192			197			1.8			1.0			35.10			35.10			3.91			3.90			1.18			1.15
	51		168			169			1.5			1.0			14.80			14.08			3.01			3.05			0.96			0.92
5.3		51			17			17			1.2			1.0			10.35			10.35			2.82			2.84			0.74		0.74

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Table 11.17 Drillhole Au Statistics for Uncomposited (Raw) and Composited (Comp.) Sample Data – Listwanite

FAULTZON		No. Samples				Max. Length (m)				Max. Grade (Au ppm)				Mean Grade<br> (Au ppm)				CV<br> (Au)
	Raw		Comp.		Raw		Comp.		Raw		Comp.		Raw		Comp.		Raw		Comp.
1		203		208		1.6		1.0		8.01		8.01		0.86		0.87		1.70		1.65
2		229		234		1.5		1.0		15.10		15.10		1.12		1.13		1.87		1.83
2.1		130		132		3.0		1.0		4.99		4.99		0.47		0.45		2.14		2.17
6		944		959		1.7		1.0		25.40		24.80		1.08		1.09		2.56		2.47
8.1		142		150		1.9		1.0		11.40		8.31		0.86		0.85		1.85		1.72
8.2		253		256		1.6		1.0		14.55		14.55		0.87		0.88		1.78		1.73
8.3		198		202		1.7		1.0		11.10		9.13		0.66		0.64		2.20		2.17
9		480		491		2.5		1.0		16.55		16.55		0.59		0.61		2.33		2.36
10.1		318		330		1.8		1.0		19.30		17.79		1.46		1.45		1.64		1.55
10.2		198		204		1.7		1.0		12.25		12.25		1.42		1.44		1.68		1.65
13.1		279		289		2.3		1.0		8.84		7.89		0.70		0.70		1.90		1.79
13.2		58		59		1.5		1.0		6.41		6.41		1.56		1.55		1.10		1.09
14		66		67		1.6		1.0		8.60		8.60		0.87		0.83		2.05		2.08
18		8		11		2.7		1.0		0.05		0.05		0.02		0.02		1.25		1.08
19.1		117		117		1.4		1.0		5.73		5.73		0.51		0.51		1.74		1.72
19.2		92		93		1.6		1.0		4.81		4.81		0.21		0.22		3.23		3.20
25		459		475		2.0		1.0		16.95		16.95		1.03		1.03		1.64		1.61
28		783		802		2.0		1.0		25.20		25.20		1.69		1.69		1.82		1.74
29		281		283		1.8		1.0		14.35		14.35		1.02		1.01		1.67		1.62
29.1		68		69		1.7		1.0		2.63		2.63		0.27		0.25		1.85		1.85
29.2		59		62		3.0		1.0		9.88		5.61		0.96		0.96		1.79		1.46
29.3		25		25		1.6		1.0		2.24		2.24		0.28		0.30		1.73		1.65
38		88		89		1.6		1.0		7.40		7.40		0.49		0.50		2.48		2.31
48		27		29		1.5		1.0		0.60		0.54		0.12		0.12		1.43		1.38
51		3,028		3,079		2.0		1.0		17.15		17.15		0.58		0.59		1.75		1.75

11.3.11 Ardich Resource Model Estimation

A cell model with 10 m x 10 m x 5 m parent cells was constructed to cover the entire Ardich deposit. The cell model parameters are shown in Table 11.18.

Table 11.18 Ardich Cell Model Prototype Parameters

Direction	Minimum<br>(m)	Maximum<br>(m)	Range<br>(m)	Cell Size<br>(m)	No. of Cells
X (East)	461,800	466,000	4,200	10	420
Y (North)	4,365,500	4,368,500	3,000	10	300
Z (RL)	600	1,650	1,050	5	210

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Sub-celling to 2.5 m x 2.5 m x 1.25 m was permitted to honour interpreted boundaries. Further sub-celling to a minimum of 0.25 m in the Z (RL) direction was permitted at the topographic surface.

The model was not rotated.

This model was populated with the same domain fields as the sample files; these being FAULTZON, MODLITH, LITHNUM, and for some domains the field ‘SUBDOM’, were used to generate the estimation flagging domains required for the subsequent grade estimation processes.

The combination of the interpretated lithological and fault block attributes resulted in the creation of 218 unique domains.

The concatenated field ‘AUESTFLG’ was used to define the estimation domains for the Au and other elements estimation, and ‘SESTFLG’ for the sulfur estimation. The unique estimation domain values were compiled using the attributes FAULTZON, LITHNUM, and SUBDOM (where relevant).

The AUESFLG and SESTFLG identifiers were calculated using the following formula:

AUESTFLG / SESTFLG = (FAULTZON x 10,000) + (LITHNUM x 100) + (SUBDOM)

A suite of 13 elements were estimated using ordinary kriging (OK). Au(cut) plus Au(uncut), Ag, As, C, Ca, Cu, Fe, Mg, Mn, Pb, Sb, and Zn (referred to as Au and other elements) were estimated using the estimation, search, and variogram parameter inputs developed for Au. In addition, a nearest neighbour Au(cut) estimation was undertaken, to be used for validation. A separate OK grade estimation for sulfur (S) was undertaken using a unique set of estimation, search, and variogram parameters developed specifically for that element. Sulfide sulfur (SS) was not estimated; rather, it was calculated using linear regressions based on the estimated sulfur grades.

Several methods were used to honour the observed styles of mineralisation and local variability. These included a dynamic search ellipse orientation for the mineralised domains, and the use of sub-domaining to implement of hard / soft boundaries proximal to contact mineralised zones (SUBDOM) and forced sub-celling with sub-cell estimation at lower contacts in Jasperoids.

11.3.11.1 Search Parameters

The search parameters used for the estimation were developed based on several criteria relating to the nature of the mineralisation and the composition of estimation domains.

Each of the 218 individual domains across the full spectrum of MODLITH and FAULTZON combinations were able to be categorised into two broad domain types. These were ‘thick’ and ‘thin’ domains. Thick domains are comprised mainly of whole parent cells and long drillhole intercepts. Thin domains comprise predominately sub-cells and have shorter drillhole intercepts.

For both the thick and thin domains the nature of the mineralisation is the same. The grades are more consistent laterally and change much quicker in the vertical direction. Lateral change in grade can still be significant, however, and the local grade conditions should be emphasised in the estimation. This was achieved by selecting search parameters that directed the estimation to select samples close to the cell and reduce the number of samples used from more distant locations.

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The singular factor differentiating the two domain types (thick and thin) was the length of the drillhole intercept and the number of samples present in each drillhole. This factor was of primary concern for appropriately selecting the minimum and maximum sample attributes, to ensure that the goal of producing an estimation that represented the local grade conditions was met.

These concepts regarding the nature of the mineralisation and the sample data led to the following concepts being incorporated into the search volume criteria:

•The orientation of the search volume will be rotated to the local conditions by the dynamic anisotropy function. However, the Z axis search distance was to be reduced compared to the X and Y axes to honour both the variography and the visual observations of grade continuity being lower in the Z direction.

•A minimum of two drillholes are required for a cell to estimate in the first, second and third estimation passes. A single drillhole may be used for the fourth estimation pass.

•No more than six samples from any individual drillhole to be used in the thick domains and no more than three samples from any individual drillhole to be used in the thin domains. Due to the angle of the drillholes, 6 m of drillhole generally represents 5 m vertical metres, or a full parent cell of the model. Therefore, capping the samples from any individual drillhole at six ensures that a full 5 m high parent cell is using approximately five vertical metres of drilling data.

•The maximum number of samples used to estimate a cell is 14 for the thick domains and 12 samples for the thin domains. The selection of these values, in conjunction with the maximum samples per hole values was designed to limit the number of distant holes / samples used in the estimation in preference to local grade conditions. This, while still ensuring that two, but preferably three holes were used to estimate most cells.

•No octant search restrictions were used.

The search parameters for grades estimation are shown in Table 11.19.

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Table 11.19 Ardich Search Volume Criteria for Grades Estimation in All Domains

LITHOLOGY		DOMAIN TYPE		Search Ellipse Rotation							Search Volume				Search Distance (m)						No. of Samples		Max. <br>Samples<br>Per Hole
	Rotation 1<br>Axis 3		Rotation 2<br>Axis 1		Rotation 3<br>Axis 3			Search Pass		Volume Factor		X		Y		Z		Min.		Max.
Jasperoid, Listwanite, Dolomite, <br>Cataclasite, <br>Hornfels, Ophiolite, <br>and Silica Cap		THICK		DA		DA			DA		1		1		40		40		10		8	14	6
	2		2				80			80		20		8		14		6
	3		3				120			120		30		8		14		6
	4		3				120			120		30		4		14		6
Cataclasite (FAULTZON 2)		THICK		DA		DA			DA		1		1		40		40		10		8	12	6
	2		2				80			80		20		8		12		6
	3		3				120			120		30		8		12		6
	4		3				120			120		30		4		12		6
Diorite		THICK		045		0			0		1		1		40		40		10		8	14	6
	2		2				80			80		20		8		14		6
	3		3				120			120		30		8		14		6
	4		3				120			120		30		4		14		6
Jasperoid and Listwanite		THIN		DA		DA			DA		1		1		40		40		10		4	12	3
	2		2				80			80		20		4		12		3
	3		3				120			120		30		4		12		3
	4		3				120			120		30		3		14		3
Diorite and Ophiolite		THIN		045		0			0		1		1		40		40		10		4	12	3
	2		2				80			80		20		4		12		3
	3		3				120			120		30		4		12		3
	4		3				120			120		30		3		14		3

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11.3.11.2 Variography

Variograms were generated for Au and S using the 1 m composites (Table 11.20 and Table 11.21).

Due to the structural complexity of the deposit, many domains are relatively small and are often offset vertically from the same lithological unit in neighbouring domains. This often resulted in insufficient samples to generate a variogram at the individual domain level and the dislocation from domain-to-domain also hampered the development of a deposit-wide variogram model for each lithology.

FAULTZON 2.1 (SC1) was selected as a representative domain within the deposit that contained sufficient data in each of the main mineralised lithology units to generate robust variograms. Variograms were developed for the jasperoid, listwanite, and dolomite using the FAULTZON 2.1 (SC1) data, and these were applied across the deposit.

The use of the dynamic anisotropy function allowed the search volume orientation as well as the variogram to be rotated to the local conditions in the other domains.

The grade continuity identified in the gold variograms was consistent with the visual observations of the mineralisation with grade continuity highest perpendicular to the lithology surfaces and lowest in the Z direction. The sulfur variograms showed more grade continuity than the gold variograms with ranges often twice as long as the gold variograms. This higher continuity observed in the sulfur variogram data is supported by visual observations of the surface data.

Table 11.20 Gold Variogram Parameters

MODLITH		Rotation Angles						Nugget		Structure	Structure Variance		Range
	Axis 3 (Z)		Axis 1 (X)		Axis 3 (Z)		Axis 1		Axis 2			Axis 3
Jasperoid and Diorite		–110		20		–25		0.38		1	0.62		85	106	10
Listwanite		–105		5		–160		0.06		1	0.94		150	60	10
All Other		–100		5		–5		0.13		1	0.87		123	84	12

Table 11.21 Sulfur Variogram Parameters

MODLITH		Rotation Angles							Nugget		Structure		Structure Variance		Range
	Axis 3 (Z)		Axis 1 (X)			Axis 3 (Z)		Axis 1		Axis 2				Axis 3
Jasperoid		–105			10		–50		0.1		1		0.90		195	140	20
Listwanite		90			5		175		0.15		1		0.85		134	78	20
Dolomite		80			90		180		0.18		1		0.24		66	14	54
	2			0.58				179		23		89
Cataclasite		–140			170		0		0.04		1		0.96		153	96	12

11.3.12 Ardich Density Model

The density estimation was undertaken within a similar framework to the grade estimations, likewise, using ordinary kriging. The sub-domaining routines employed for the grade estimations were not used for the density estimation.

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The model field ‘SGESTFLG’ was created using the LITHNUM and FAULTZON fields to control the estimation domaining.

Due to the lower number of samples present in the density dataset compared to the grade datasets (12,419 density samples vs. 72,432 Au assays), specific density estimation search parameters were developed. Density variogram models were also developed utilising samples from FAULTZONs SC1, C4, and EC2.

The density variogram models showed a longer range of continuity than was shown in the gold or sulfur variograms. The density variograms were different to the grade variogram models, however overall the density variogram models could also be described as having greatest continuity parallel to the lithology contacts in the same way as the gold and sulfur variograms do. Therefore, the TRDIPDIR and TRDIP values estimated into the model and used for the grade estimation were also used for the density estimation.

Minimum and maximum sample numbers were developed based on the density dataset. In a similar manner to the grade search parameters a ‘thick’ and ‘thin’ domain type designation was applied depending on the domain.

There are a few notable differences between the density and grade estimations. First is removal of the fourth search volume, with the density estimation utilising only three search volumes. Secondly, cells in the ‘thick’ estimation domains can estimate from a single drillhole in the third search volume, whilst in the ‘thin’ domain’s cells can estimate from a single drillhole in all search volumes. These conditions reflect the lower number of density samples and the distribution of these samples within these domains.

The search parameters for grades estimation are shown in Table 11.22.

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Table 11.22 Ardich Search Volume Criteria for Density Estimation in All Domains

LITHOLOGY		DOMAIN TYPE		Search Ellipse Rotation							Search Volume				Search Distance (m)						No. of Samples		Max. <br>Samples<br>Per Hole
	Rotation 1<br>Axis 3		Rotation 2<br>Axis 1		Rotation 3<br>Axis 3			Search Pass		Volume Factor		X		Y		Z		Min.		Max.
Jasperoid, Dolomite, Cataclasite, <br>Ophiolite, and Listwanite		THICK		DA		DA			DA		1		-		40		40		10		4	8	3
	2		2				80			80		20		4		8		3
	3		5				200			200		50		3		8		3
Hornfels and Silica Cap		THICK		DA		DA			DA		1		-		40		40		10		4	8	3
	2		2				80			80		20		4		8		3
	3		5				200			200		50		3		8		3
Diorite		THICK		045		0			0		1		-		40		40		10		2	6	3
	2		2				80			80		20		2		6		3
	3		5				200			200		50		2		6		3
Jasperoid and Listwanite		THIN		DA		DA			DA		1		-		40		40		10		2	6	3
	2		2				80			80		20		2		6		3
	3		5				200			200		50		2		6		3
Diorite, Hornfels, <br>and Silica Cap		THIN		045		0			0		1		-		40		40		10		4	8	3
	2		2				80			80		20		4		8		3
	3		5				200			200		50		3		8		3

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11.3.13 Ardich Resource Classification

The grade estimates classifications were stored in the model field ‘RESCAT’.

The values used to indicate classification in the RESCAT field are as follows:

•1 = Measured

•2 = Indicated

•3 = Inferred

•0 = Unclassified

RESCAT was considered in detail on the basis of the unique properties of each lithological unit within each domain.

Emphasis was placed on those lithologies that host the most-significant mineralisation in each domain (generally jasperoid and listwanite).

Unique polygons were digitised in plan view around areas considered to have higher confidence interpretations and estimates within each domain.

Because of the high degree of complication in the domain interpretations, and the variable nature of the mineralisation, the classification method was not based on any single attribute or parameter (such as drillhole spacing, search volume, distance to nearest sample, or number of samples). Rather, a more interactive and holistic approach to the classification was adopted.

The key factors in assessing classification were observations related to the geological and grade continuity of each domain. This, coupled with an assessment of the performance of the estimation, formed the primary basis for the classification. Additional information such as the drill spacing and the search volume were also considered during the classification process, however they were not used categorically or rigidly, and were not the primary drivers for defining the RESCAT shapes.

For example, a domain with well-understood geology and highly consistent mineralisation, but with relatively wide drill spacing, may receive a similar or a better RESCAT than a domain where the drill spacing is closer, but confidence in the geological interpretation is lower and the mineralisation is less consistent.

The most prominent role that the distribution of the drilling played in the classification was related to the overall number of drillholes informing the interpretation in a domain, rather than the drillhole spacing specifically.

A detailed log of the decision making process for each RESCAT polygon was developed and retained for future reference.

Figure 11.15, Figure 11.16, and Figure 11.17 show the model coloured by RESCAT for the jasperoid, listwanite, and upper dolomite lithologies.

For unmineralised domains, such as the ophiolite, lower dolomite, and lower cataclasite, default RESCAT values were applied.

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Figure 11.15 Plan View of Jasperoid Resource Classification

OreWin, 2021

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Figure 11.16 Plan View of Listwanite Resource Classification

OreWin, 2021

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Figure 11.17 Plan View of Upper Dolomite Resource Classification

OreWin, 2021

11.3.14 Ardich Model Validation

Model validation was approached in several ways:

•The estimated Au grades in the model were compared to the drillhole grades by visual inspection in plan views, sectional views, and in 3D. In general, the model and composite grades visually compared well, and the estimates were considered to have honoured the interpreted mineralisation styles.

•The cell model was checked for global bias by comparing the Au and S statistics of the model estimates compared to the input sample file and a nearest neighbour (NN) estimation. The NN estimation produces a theoretically unbiased (declustered) estimate of the mean, offering an alternative metric to sample grade for comparing the estimation. The cells used to calculate these mean values were restricted to those cells that were classified as Measured, Indicated, or Inferred, thus removing any poorly estimated or default grade cells on the periphery of the domains. The means comparison shows that for the jasperoid and listwanite there is good correlation between the OK and NN estimations. Correlation is lower for the dolomite and cataclasite, which is not considered remarkable given the more-variable nature of those domains.

•Local trends in the grade estimates (also known as drift analysis) were assessed by plotting the mean values from the NN estimate versus the kriged results for Indicated model cells in east–west, north–south and vertical directions (swath plots). There is good correlation between the OK and NN estimations in all orientations. However, the correlation diverges at the model edges, especially in the north-eastern corner where data density becomes reduced and a single drillhole or sample can have a disproportionate effect on the NN estimates.

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11.3.15 Ardich Change of Support

Mining has not occurred at the Ardich project area and therefore no production data is available. A 5 m mining bench is anticipated, with 5 m blastholes likely to be used. Grade estimation at Ardich is based on 1 m assay composites and interpreted geological (structural and lithological) boundaries to estimate resource model tonnes and grade using ordinary kriging.

11.3.16 Ardich Assessment of Reasonable Prospects of Eventual Economic Extraction

Refer to Section 11.5 for the assessments for Mineral Resource estimates meeting reasonable prospects for eventual economic extraction.

11.3.17 Ardich Mineral Resource Tabulation

Ardich Mineral Resources have been tabulated by Mineral Resource classification and oxidation state in Table 11.57. Mineral Resources are presented showing only the SSR attributable proportion.

Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

The overall tonnage and grade estimate have increased for oxide and sulfide material from the previously reported estimate in 2019. This change is predominantly due to the change in the conceptual model, supported by the additional drilling obtained since 2020.

Resource pit shells were generated by OreWin using a metal price assumption of $1,750/oz gold. Gold mineralisation modelled at Ardich is primarily oxidised with a smaller portion of sulfur mineralisation having estimated total sulfur grades >2%. Low-sulfur (LS) oxide is defined as material with <1% total sulfur. High-sulfur (HS) oxide is material with total sulfur >1% and <2%. Sulfide material has ≥2% total sulfur. The Mineral Resources are shown in Table 11.57.

Internal cut-off grades for oxide material range from 0.30–0.55 g/t Au. Sulfide is material with >2% total sulfur above a 1.1 g/t Au cut-off.

11.4 Bayramdere Deposit

The Bayramdere deposit is located approximately 6.3 km east of the Çöpler mine and 5 km south-east of İliç. Bayramdere is within the Kartaltepe Mining Licence 7083. This licence is an operational licence and is 50% SSR-held.

Soil samples have been collected across the prospect on a 100 m x 100 m grid. Soil copper and gold anomalies are identified as coincident with each other, but the copper anomaly covers a larger area.

The Bayramdere mineralisation has an overall strike length of approximately 300 m. Mineralisation is localised within three stacked, shallow-dipping lodes that are very close to the surface, varying in depth 30–40 m below topography. Mineralisation appears to be open to the east and south.

The mineralisation has formed at the contacts of limestone and ophiolite lithologies with mineralisation replacing limestone along the contacts. The limestone to ophiolite contacts are low-angle thrusts, with limestone typically being trapped as wedges of material within a dominantly ophiolite stratigraphy. Mineralisation occurs within iron-rich gossan horizons.

Although a small deposit, Bayramdere is relatively high grade and can support a high stripping ratio to access mineralisation.

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Small-scale open pit iron ore mining has occurred historically at Bayramdere. Iron mineralisation can be associated with gold mineralisation.

11.4.1 Bayramdere Domains

The geological interpretation was represented in the geological model through the creation of mineralised domains based on the continuity of the geology and mineralisation identified specific to each deposit and mineralised zone within the deposit. Separate domains were created for gold, silver, copper, and sulfur. In the creation of mineralised domains, a minimum mining width of 2.5 m was used based on anticipated open pit mining methods.

11.4.2 Bayramdere Geological Model

The Bayramdere deposit is a structurally controlled gold±minor copper±minor silver deposit displaying both epithermal and replacement mineralisation styles. At this stage of exploration, the deposit is dominantly represented by near-surface oxide mineralisation to a depth of up to 180 m below surface. Mineralisation is primarily associated with jasperoid and iron-rich gossan. Secondary pyrite is a commonly visible component within the jasperoids.

At depth, mineralisation transitions below the base of complete oxidation to disseminated pyrite, vein sulfides, and massive sulfide horizons generally occurring within shear zones, along shallow thrusts and diorite sill and dyke margins. The extent of sulfide mineralisation has not been tested.

As with the other Çöpler district deposits, Bayramdere is considered to be the result of a mineralised intrusion generating suitable conditions for mineralisation to be localised into a favourable geological setting of ophiolite, limestone, and hornfels lithologies (see Figure 11.18). A complex system of faults and thrusts have allowed mineralised fluids and diorite dykes and sills associated with the epithermal system to permeate into the stratigraphy.

Like the Çakmaktepe deposit, Bayramdere is associated with flat thrust structures. Key to each structurally associated style of mineralisation is the juxtaposition of ophiolites against limestone + hornfels to create suitable geochemical conditions for gold and other metals deposition. Ophiolite is not associated with mineralisation at Çöpler, this association at present is considered to be unique to Bayramdere and Çakmaktepe.

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Figure 11.18 Bayramdere Geology Schematic Section

Anagold, 2017

11.4.3 Bayramdere Data Summary

The Bayramdere deposit has been drilled on 25 m lines with 20–25 m spaced holes on each line.

A total of 120 resource definition drillholes have been drilled at Bayramdere for a total length of 10,734.2 m, inclusive of metallurgical holes. The assay database includes 8,283 sample intervals for a total assayed length of 10,483.4 m.

When categorised according to type of drilling (excluding geotechnical and metallurgical drillholes), 30% are RC samples, 65% DD core samples, and 6% are a combination of RC and DD core.

Drilling has been completed on drill grids aligned at right angles to mineralisation trends or lithology dip and strike. Several areas contain scissor holes that test mineralisation at 180° from each other.

11.4.4 Bayramdere Drillhole Compositing

Sample compositing has not been applied. The predominant sample length is 1.0 m (52%), followed by 2 m as the next most prevalent length (17%).

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11.4.5 Bayramdere Top Cutting

High-grade top cuts were applied after selecting appropriate limits based on cumulative frequency plots and value grade curves of the upper portion of the sample population.

11.4.6 Bayramdere Cell Model

The Bayramdere cell model parameters are shown in Table 11.23.

Sub-celling was permitted to 2 m x 2 m x 1 m to better honour the domain boundaries.

Table 11.23 Bayramdere Cell Model Prototype Parameters

Direction	Minimum<br>(m)	Maximum<br>(m)	Range<br>(m)	Cell Size<br>(m)	Number of Cells
East	466,000	466,600	600	10	60
North	4,363,800	4,364,100	300	10	30
RL	1,250	1,420	170	5	34

11.4.7 Bayramdere Estimation Method

Estimation was limited to the interpreted domains, with each domain informed only by samples contained within that domain. Outside the mineralised domains a ‘mineralised waste’ estimate was completed.

Mineralisation domains were also developed for silver, copper, and sulfur.

Lithological domains were used for estimates outside of the mineralisation domains.

Ordinary kriging was used to estimate Au, Ag, and Cu into parent cells. Variography was completed to inform estimation.

11.4.8 Bayramdere Density Model

Density has been assigned as a default for each of the mineralisation and lithological domains (see Table 11.24 and Table 11.25 respectively). The assigned densities reflect the arithmetic average of the domain-relevant data taken from DD core samples.

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Table 11.24 Bayramdere Density Values for Mineralisation Domains

Domain	Density<br>(t/m3)
mz100	2.69
mz101	2.49
mz200	2.55
mz201	2.79
mz300	2.29
mz301	2.29
mz400	2.49
mz500	2.42
mz600	2.49
mz700	2.49

Table 11.25 Bayramdere Density Values for Lithology Domains

Domain	Weathering State		Density<br>(t/m3)
Gossan	Weathered		2.50
Diorite		2.44
Limestone		2.54
Ophiolite		2.36
Gossan	Fresh		2.50
Diorite		2.44
Limestone		2.54
Ophiolite		2.36
Overburden	All		1.40

11.4.9 Bayramdere Resource Classification

Grade estimates were classified using the following Anagold guidelines:

•Indicated Mineral Resource should be quantified within relative ±15% with 90% confidence on an annual basis, and

•Measured Mineral Resources should be known within ±15% with 90% confidence on a quarterly basis.

Drillhole spacing for support of classification of Inferred Mineral Resources was required to be 50 m x 25 m spacing. For Indicated Mineral Resource classification, the drillhole spacing requirement was reduced to 25 m x 25 m spacing. Appropriate drillhole pattern spacing selection was based on the understanding of the nature of the mineralisation being structurally controlled, mineral continuity, and assessment of data quality.

The drillhole spacing at Bayramdere is considered sufficient to support grade continuity, geological continuity, depth, and lateral extents of mineralisation.

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No Bayramdere estimates were classified in the Measured category.

Mineral Resources were tabulated using multiple cut-off grades due to variable recoveries and based on gold price only. Cut-off grades vary from 0.35–0.50 g/t Au and are calculated based on the equation:

Xc = Po / (r * (V-R))

where Xc = Cut-off Grade (g/t), Po = processing cost of ore (USD/tonne of ore), r = recovery, V = gold sell price ($/g), R = refining costs ($/g).

Mineral Resources are reported inclusive of Mineral Reserves.

11.4.10 Bayramdere Validation

Bayramdere grade estimates were validated against alternate interpolation methods. Estimated grades were compared to an ID2 model to check for global bias. Swath plots were used to check for a local bias. The estimated Au grades in the model were compared to the composite grades by visual inspection in plan views and cross-sections. Composite samples were queried by domain to confirm appropriate sample flagging.

11.4.11 Bayramdere Assessment of Reasonable Prospects of Eventual Economic Extraction

Refer to Section 11.5 for the assessments for Mineral Resource estimates meeting reasonable prospects for eventual economic extraction.

11.4.12 Bayramdere Mineral Resource Tabulation

Bayramdere Mineral Resources are reported exclusive of Mineral Reserves and have been tabulated by Mineral Resource classification and oxidation state in Table 11.57. Mineral Resources are presented showing only the SSR attributable proportion. Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

11.5 Assessment for Reasonable Prospects for Economic Extraction

Mineral Resources in the 2021MR were assessed for reasonable prospects for eventual economic extraction by one of two ways. For existing operations, by reporting only material that fell within conceptual pit shells based on metal prices of $1,750/oz for gold, or as otherwise specified, and an Initial Assessment has been prepared to demonstrate the further economic potential of the Mineral Resources at the Çöpler deposit with the inclusion of a Copper Concentrator.

The 2021MR is an independent Technical Report Summary prepared to provide a preliminary technical and economic study of the economic potential of the Çöpler District mineralisation to support the disclosure of Mineral Resources. The 2021MR includes an Initial Assessment that has two cases:

•Initial Assessment MII Case – the base case that includes Measured, Indicated and Inferred Mineral Resources in the analysis and

•Initial Assessment MI Case – this case includes only Measured and Indicated Mineral Resources

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Content for this section has been sourced from the published 2021MR report. The following sections discuss the assessments for Mineral Resource estimates meeting reasonable prospects for eventual economic extraction.

Macroeconomic trends, taxes, royalties, data, and assumptions, interest rates, marketing information and plans, legal matters such as statutory and regulatory interpretations affecting the mine plan and environmental matters are outside the expertise of the QP and are within the control of the registrant (see Section 25).

However, as significant environmental and social analysis has been conducted for the Project over an extended period, the Project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas, and following a review of the current supplied information, the opinion of the QP is it reasonable to rely on the information provided by SSR.

11.5.1 Gold Mineral Resources

11.5.1.1 Çöpler Assessment of Reasonable Prospects of Eventual Economic Extraction

Mineral Resource estimates were shown to meet reasonable prospects for eventual economic extraction criteria by reporting only material that was contained within a conceptual pit shell using metal prices of $1,750/oz for gold and the parameters summarised in Table 11.26. The gold price of $1,750/oz was selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The metal price is representative of the range of price estimates publicly reported for Mineral Resource cut-offs. The Çöpler Mineral Resource is assumed to be mined by open pit.

Table 11.26 Summary of Key Parameters Used in 2021 Conceptual Pit Shell at Çöpler

Description	Unit	Minimum	Maximum
Heap Leach Gold Recovery	%	62.3	78.4
POX Gold Recovery	%	91.0	91.0
Mining Cost per tonne mined	$/t	1.49	2.78
Process Costs Heap Leach	$/t	9.30	9.30
Process Costs POX	$/t	34.88	34.88
Site Support and G&A	$Mpa	15	15
Internal Au Cut-off – Heap Leach	g/t	0.19	0.24
Internal Au Cut-off – POX	$/t NSR	34.88	34.88
Internal Au Cut-off – Cu Conc.	$/t NSR	7.68 + 34.88 x Pyrite Mass Pull
Royalty	%	2.0	2.0

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11.5.1.2 Çakmaktepe Assessment of Reasonable Prospects of Eventual Economic Extraction

Mineral Resource estimates were shown to meet reasonable prospects for eventual economic extraction criteria by reporting only material that was contained within a conceptual pit shell using metal prices of $1,750/oz for gold with the parameters summarised in Table 11.27. The gold price of $1,750/oz was selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The metal price is representative of the range of price estimates publicly reported for Mineral Resource cut-offs. The Çakmaktepe Mineral Resource is assumed to be mined by open pit.

Table 11.27 Summary of Key Parameters Used in Conceptual Pit Shell at Çakmaktepe

Description	Unit	Minimum	Maximum
Heap Leach Gold Recovery	%	38.0	80.0
Mining Cost per tonne mined	$/t	1.59	1.59
Process Costs Heap Leach	$/t	14.16	14.16
Site Support per tonne processed	$/t	3.17	3.17
Internal Au Cut-off – Heap Leach	g/t	0.36	0.76
Royalty	%	4.0	4.0

11.5.1.3 Ardich Assessment of Reasonable Prospects of Eventual Economic Extraction

Mineral Resource estimates were shown to meet reasonable prospects for eventual economic extraction criteria by reporting only material that was contained within a conceptual pit shell using metal prices of $1,750/oz for gold with the parameters summarised in Table 11.28. The gold price of $1,750/oz was selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The metal price is representative of the range of price estimates publicly reported for Mineral Resource cut-offs. The Ardich Mineral Resource is assumed to be mined by open pit.

Table 11.28 Summary of Key Parameters Used in Conceptual Pit Shell at Ardich

Description	Unit	Minimum	Maximum
Heap Leach Gold Recovery	%	40.0	73.0
POX Gold Recovery	%	82.9	82.9
Mining Cost per tonne mined	$/t	1.82	1.82
Process Costs Heap Leach	$/t	10.68	10.68
Process Costs POX	$/t	36.25	36.25
Site Support and G&A	$Mpa	15	15
Internal Au Cut-off – Heap Leach	g/t	0.25	0.46
Internal Au Cut-off – POX	$/t NSR	0.77	0.77
Internal Au Cut-off – Cu Conc.	$/t NSR	9.05 + 36.25 x Pyrite Mass Pull
Royalty	%	2.0	2.0

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11.5.1.4 Bayramdere Assessment of Reasonable Prospects of Eventual Economic Extraction

Table 11.29 Summary of Key Parameters Used in Conceptual Pit Shell at Bayramdere

Description	Unit	Minimum	Maximum
Heap Leach Gold Recovery	%	75.0	75.0
Mining Cost per tonne mined	$/t	1.75	1.75
Process Costs Heap Leach	$/t	9.99	9.99
Site Support per tonne processed	$/t	3.19	3.19
Internal Au Cut-off – Heap Leach	g/t	0.35	0.50
Royalty	%	2.0	2.0

11.5.2 Çöpler District Initial Assessment

•Initial Assessment MII Case - the base case that includes Measured, Indicated and Inferred Mineral Resources in the analysis, and

•Initial Assessment MI Case - this case includes only Measured and Indicated Mineral Resources

The Initial Assessment has been prepared on a 100% project basis to analyse the impact of changes in processing method for the Çöpler Mineral Resource.

The existing Çöpler project currently has two processing methods:

•Sulfide process plant

•Heap leach oxide processing facility

The sulfide plant includes the crushing, grinding, flotation and pressure oxidation to produce gold and small amounts of silver The heap leach facility produces gold and small quantities of silver and copper.

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The scenario for the Initial Assessment analyses includes additional processing options to recover copper from the sulfide Mineral Resource. The two processing options are:

•Copper concentrator producing a copper concentrate and a pyrite concentrate.

•Sodium hydrosulfide (NaSH) copper recovery circuit to be installed in the current sulfide plant.

The copper concentrator would make a copper concentrate for sale to smelters and a pyrite concentrate to be fed into the autoclaves in the sulfide plant. The pyrite concentrate would have a high gold content and provide sulfur as a source of fuel for the autoclaves. The copper concentrator capacity is 1.8 Mtpa.

The Çöpler Mineral Resource has been selected for the copper concentrator analysis because the other Mineral Resources at the Project do not have significant amounts of copper.

An economic analysis was performed on a 100% project basis for the expected life of mine of the Project. Two economic models were prepared for the Project: the base case includes Inferred Mineral Resources in the analysis (MII Case), and the second excludes the Inferred material (MI Case). The economic model results are based on Mineral Resources that, unlike Mineral Reserves, do not have demonstrated economic viability.

Implementation of the copper recovery options will require capital expenditures and will also provide opportunities for operational cost and productivity improvements. The Initial Assessment shows the results of a shorter term analysis using metal prices used for the CDMP21TRS Reserve cut-off grade definition (refer Section 13.2.1) and the impact of the estimated capital and potential cost savings from economies of scale and reallocation of shared and fixed costs.

For the Initial Assessment economic analysis, the Ardich and Çakmaktepe pits have been included in the cash flow analysis without change from the CDMP21TRS Reserve Case. This allowed the analysis to quantify the impact of the copper concentrator and NaSH circuit and demonstrate the potential (or otherwise) of the additional Mineral Resources.

11.5.2.1 Copper Recovery Processing Assumptions

A simplified process flow diagram after the addition of the copper concentrator and NaSH circuit is shown in Figure 11.19.

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Figure 11.19 Çöpler Project Initial Assessment Simplified Process Flow Diagram

OreWin, 2022

Copper Concentrator

A preliminary design was developed for a concentrator to process copper and pyrite bearing ore at the Çöpler mine site. The design throughput rate was assumed to be 1.8 Mtpa of ROM ore producing saleable copper concentrate with gold and silver and a sulfur-rich pyrite concentrate containing gold for fuel in the autoclave.

The copper concentrator plant includes a ROM pad and a comminution circuit consisting of a sizer and single stage SAG mill. The concentrator consists of a copper flotation circuit inclusive of roughers, copper concentrate regrind, and cleaner flotation to produce a copper concentrate. The copper concentrate is thickened, filtered, and concentrate loadout into bulk bags for shipment to a smelter. Figure 11.20 shows the copper concentrator process flow diagram.

The pyrite flotation circuit treats the copper flotation tails and consists of roughers, a concentrate regrind mill, and cleaner flotation to produce a pyrite concentrate. The pyrite concentrate is thickened and stored in agitated tanks and pumped to the existing gold POX circuit as required to supplement autoclave sulfur requirements.

Final Tailings are thickened and pumped to the TSF. The circuit is inclusive of associated plant services including reagents, air, and water.

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Figure 11.20 Çöpler Preliminary Copper Concentrator Process Flow Diagram

Ausenco, 2021

NaSH Copper Recovery Circuit

When originally constructed, the POX circuit did not include extraction of copper. A new circuit assuming precipitation from the pregnant leach solution by sodium hydrosulfide (NaSH) sulfidisation would be expected to achieve high copper recoveries.

Allowance was made in the existing plant for space to install a copper recovery circuit in the future.

NSR Inputs and Cut-off Grade Calculations

The cut-off grades for the CDMP21TRS Mineral Reserve are presented as a gold only cut-off grade because the majority of the model cell value is derived from gold. For the analysis of the copper recovery, the copper proportion of the plant feeds have a significant revenue, therefore it is necessary to determine the cell values from the three revenue elements: gold, copper, and silver. For this reason, a NSR was calculated for each cell in the mining model for the Initial Assessment for the sulfide and copper concentrator scenario.

Three processing material types were defined for the Initial Assessment. These three processing material types are shown in Table 11.30.

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Table 11.30 Initial Assessment Case Process Types

Ore Type	Defining Criteria
Oxide	S% < 2
Sulfide	S% >= 2
Copper Concentrator	Cu% >= 0.1

Initial Assessment Processing Recoveries

The processing parameters used for the calculation of the NSR and cut-off grades are shown in Table 11.31 to Table 11.34.

Table 11.31 Heap Leach Recoveries

Item	Units	Amount
Au Recovery	%	62.3%–78.4%

Table 11.32 Copper Concentrate Recoveries

Item	Units	Amount
Au Recovery	%	55
Ag Recovery	%	45
Cu Recovery	%	84
Mass Pull = ( 2 x Cu + 0.15 ) / 100

Table 11.33 Pyrite Concentrate Recoveries

Item	Units	Amount
Au Recovery	%	15
Ag Recovery	%	15
Cu Recovery	%	2
Mass Pull = ( SS + 0.4 ) / 100

Table 11.34 POX Plant Recoveries

Item	Units	Amount
Au Recovery	%	91
Ag Recovery	%	10
Cu Recovery	%	98

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Costs

Costs have been estimated using actual costs from the Project, review of plans for productivity and cost savings, previous capital estimates for a NaSH circuit, and the copper concentrator scoping study.

The accuracy of the estimates are within ±50%. A 25% contingency has been added to the direct capital costs of the copper concentrator and NaSH cost estimates.

Operating and capital costs are shown in Table 11.35 to Table 11.37.

Table 11.35 Operating Costs

Item		Cost
Heap Leach Oxide		9.26
	Mpa
Sulfide Plant		34.67
Copper Concentrator		7.29
Site Support		26.78 to 19.31<br><br>8.55 after mining is completed
G&A		5.00
Mining Costs		1.49 to 3.53

All values are in US Dollars.

Table 11.36 Copper Concentrator and NaSH Capital Cost

Item	Factor	Cost $M
Cu Concentrator	–	100.2
Sulfide Cu Recovery Circuit	–	33.1
Direct Costs	–	133.3
EPCM	18%	24.6
Owner's Costs	20%	26.7
Contingency	25%	33.3
Total Capital	–	217.9

Table 11.37 Other Capital Costs

Item	Units	Amount
Closure	$M	114
Heap Leach Sustaining	$/t	0.15

Metal Prices and Selling Costs

Metal prices for the economic analysis were estimated after analysis of consensus industry metal price forecasts and compared to those used in other published studies. The metal prices used for the economic analysis, shown in Table 11.38, are considered to be representative of industry forecasts.

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Reserve cut-off metal prices are lower than the long-term forecasts and represent a conservative view of the long-term gold price. Metal prices used in cut-off’s for Mineral Resources were selected to be higher than the long-term consensus prices and are in line with other price estimates used for Mineral Resources.

Realisation assumptions are shown in Table 11.39 and Table 11.40.

Table 11.38 nomic Analysis Metal Prices Assumptions

Metal	Units	2022	2023	2024	2025	Long-Term
Gold Price	$/oz	1,800	1,740	1,710	1,670	1,600
Silver Price	$/oz	24.00	23.00	22.00	21.00	21.00
Copper Price	$/lb	4.00	3.80	3.80	3.80	3.40

Table 11.39 Transport and Treatment Charges

Item	Units	Amount
Concentrate Moisture	%	12
Concentrate Transport	$/t wet	25
Concentrate Treatment	$/t concentrate	80

Table 11.40 Payable Metal Assumptions

Item	Units	Amount
Payable Au	%	97.5
Payable Ag	%	90.0
Payable Cu	%	96.0

Çöpler Cut-off Grades

Cut-off grades in the Initial Assessment for oxide used gold cut-offs of 0.47–0.59 g/t Au.

In the sulfide Mineral Resource an NSR was calculated using the parameters discussed above.

For the copper concentrator, the cut-off applies to Mineral Resource with a Cu > 0.10%. The cut-off used was as follows:

$7.68/t NSR + ( Pyrite Mass Pull ) x $34.88/t NSR

For the remaining sulfide Mineral Resource, the cut-off used was $34.88/t NSR.

11.5.2.2 Mining

Mining in the Initial Assessment is planned to be the same as the current operation. A plan and section showing the Initial Assessment pit shell and the CDMP21TRS Reserve Case pit design are shown in Figure 11.21 and Figure 11.22

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Pit optimisation was prepared using the assumptions described above to generate pit shells. Mineral Resources classified as Measured, Indicated, and Inferred were used in the optimisation. Previous work on the Çöpler pits prepared by OreWin produced pit shells that were used for designs on the Çöpler pit. The pit optimisation work generated pit shells that were considered a close match to the CDMP21TRS Reserves Case designs. Based on this experience it was considered reasonable to use the pit shells from the optimisation work for production analysis in the Initial Assessment.

11.5.2.3 Capital and Operating Costs

Capital and operating cost estimates have been developed based on the current project costs, the mine and process designs, and discussions with potential suppliers and contractors.

Capital Costs

The following section describes the costs from the Çöpler District Master Plan 2021, which was used as the starting point for the Initial Assessment.

As the Project has been in operation for a number of years, the level of project definition for the capital cost estimate is very high. A contingency of 10% was included in the cost estimate. The QP considers the capital estimate to be in the accuracy range of +/-15%.

Growth capital costs in the CDMP21TRS Reserve Case includes costs for:

•Ardich establishment and mine development

•Heap leach phase 5 and phase 6

•Road relocation, studies, and project management

•Explosives magazine relocation

Sustaining capital in the CDMP21TRS Reserve Case includes costs for:

•Tailings storage facility (TSF)

•Project team

•Technical services

•Administration

•Assay laboratory

•Mining

•IT

•Sulfide processing

•Oxide processing

•Environment

•Mineral / lands rights

•Health and safety

•Security

•Supply chain

•Reclamation

Capital costs assumptions to the end of 2021 and for the life-of-mine (LOM) are shown in Table 11.41.

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Table 11.41 Capital and Reclamation Costs

Description	Unit	Total LOM
Oxide
Growth	$M	69
Sustaining	$M	29
Sulfide
Growth	$M	–
Sustaining	$M	413
Reclamation and Other
Reclamation	$M	114
Working and Other	$M	–37
Total	$M	588

Operating Costs

The following section describes the costs from the Çöpler District Master Plan 2021, which was used as the starting point for the Initial Assessment.

Operating costs were estimated based on current site cost performance and contract costs including actual operational costs for labour, consumables, contracts and the Anagold budget assumptions. As the Project has been in operation for a number of years, the level of project definition for the operating cost estimates is very high. Given the available project performance data and the high project definition, no contingency was included in the cost estimate. The QP considers the operating cost estimate to be in the accuracy range of 15%. The projected LOM unit operating cost estimate is summarised in Table 11.42 and the average costs are shown in Table 11.43.

Table 11.42 Average Operating Costs Unit Rates

Activity	Unit	LOM Average Unit Cost
Mining	$/t mined	1.62
Processing – Heap Leach	$/t HL processed	14.45
Processing – Sulfide	$/t sulfide processed	35.91
Site Support and Office	$/t ore processed	5.21

Table 11.43 Summary of LOM Average Operating Costs

Cost	Total LOM<br>($M)	5-Year Average<br>per year<br>($/t ore)	LOM Average<br>per year<br>($/t ore)
Mining	766	14.98	10.15
Process	2,225	27.79	29.49
Site Support and G&A	473	7.14	6.27
Operating Costs	3,346	49.91	45.91

Mining costs include waste stripping costs

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Mining Cost Summary

The mining costs were applied to the financial model as operating costs or capital costs. In the mining cost model, costs are broken down into specific areas including drill and blast, load and haul and rehabilitation.

Mining operations for the mine are currently contracted to a Turkish mining contractor. No capital cost is included for mining equipment or facilities. All such costs are built into the unit rate for mining operations included in the operating cost estimate.

Mining operating costs include:

•Drill and blast

•Load and haul

•Labour

•Dewatering

•Other indirects

Mining capital costs include:

•Fixed equipment

•Mobile equipment

•Office and supply

•Mine rehabilitation

•Studies

Processing and Infrastructure Cost Summary

The following has been included in the costs for processing:

•Oxide processing

•Sulfide processing

•Waste management

•TSF

•Utilities and services

•Reagents

•Plant infrastructure

•Plant mobile equipment

The following has been included in the capital costs for infrastructure cost estimates:

•Bulk services

•Site preparation

•Buildings and structures (new and refurbished)

•Communications

•IT hardware and software

•Security and access control

•Site costs

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•Mobile equipment

•Services contracts

•Community support

The following has been included in the operating cost estimates:

•Plant consumables

•Crusher consumables

•Screens and filters

•Grinding media

•Packaging plant bags

•Plant reagents

•Plant mobile equipment

•Plant maintenance

•Power

•Labour

•Production and dispatch

•Plant and infrastructure day work services

•Plant technical services

•Shift maintenance

•Laboratory service level agreement

•TSF water treatment

General and Administration Cost Summary

The General and Administrative (G&A) costs include costs not directly attributable to operational output such as the mining and processing operations. The following costs have been included in total G&A cost:

•Office and general expenses

•Site support costs

•Off-site Anagold offices

•Internal and external consultants

•Maintenance and inspection contracts

•Equipment and sundry, fuels and utilities

•Rentals and leases

•Insurance and insurance taxes

•IT hardware and software

•Personnel transport

•Communications

•Licences and land fees

•Labour

•Accommodation and messing

21007CDMP21TRS220927Rev0.docx Page 174 of 303

•Medical support

•Flights

•Light vehicles

•Environmental, community development and engagement

•Banking and audit fees

•Legal

11.6 Initial Assessment Inputs

The Çöpler District Master Plan 2021 costs were used as the starting point for the Initial Assessment analysis.

The Initial Assessment has been prepared on a 100% project basis to analyse the impact of changes in processing method for the Çöpler Mineral Resource. Implementation of the copper recovery options will require capital expenditures and will also provide opportunities for operational cost and productivity improvements from economies of scale and reallocation of shared and fixed costs.

Mining in the Initial Assessment is planned to be the same as the current operation.

Costs have been estimated using actual costs from the existing operations, review of plans for productivity and cost savings, previous capital estimates for a NaSH circuit, and the copper concentrator scoping study.

The accuracy of the estimates are within ±50%. A 25% contingency has been added to the direct capital costs of the copper concentrator and NaSH cost estimates.

Operating and capital costs are shown in Table 11.44 to Table 11.46.

Table 11.44 Operating Costs

Item		Cost
Heap Leach Oxide		9.26
	Mpa
Sulfide Plant		34.67
Copper Concentrator		7.29
Site Support		26.78 to 19.31<br><br>8.55 after mining is completed
G&A		5.00
Mining Costs		1.49 to 3.53

All values are in US Dollars.

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Table 11.45 Copper Concentrator and NaSH Capital Cost

Item	Factor	Cost $M
Cu Concentrator	–	100.2
Sulfide Cu Recovery Circuit	–	33.1
Direct Costs	–	133.3
EPCM	18%	24.6
Owner's Costs	20%	26.7
Contingency	25%	33.3
Total Capital	–	217.9

Table 11.46 Other Capital Costs

Item	Units	Amount
Closure	$M	114
Heap Leach Sustaining	$/t	0.15

Results of the Initial Assessment are discussed in Section 11.5.2. The assessment is preliminary in nature, it includes Inferred Mineral Resources that are considered too speculative geologically to have modifying factors applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that this economic assessment will be realised.

11.6.1.1 Production Schedule and Cash Flow

The annual mining production and process production for oxide heap leach, sulfide, and copper concentrator are shown in Table 11.47 and Table 11.50.

The Initial Assessment MII Case annual cash flow is shown in Table 11.51 and Table 11.52.

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Figure 11.21 Çöpler Plan Initial Assessment Pit Shell, Resource Shell, and Reserve Pit Design

OreWin, 2022

Figure 11.22 Çöpler Long-Section Initial Assessment Pit Shell, Resource Shell, and Reserve Pit Design

OreWin, 2022

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Table 11.47 Çöpler Initial Assessment MII Case Production Schedule

Description	Units	TOTAL	Year
			2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	2035	2036	2037	2038	2039	2040	2041	2042	2043
Total Movement	kt	721,847	26,798	36,490	52,473	50,053	53,177	51,541	45,392	45,590	46,872	46,719	46,893	41,577	25,116	20,655	22,616	23,239	22,159	21,630	20,758	18,741	2,829	529
Waste	kt	595,461	23,828	32,015	44,655	43,362	44,843	45,329	37,456	39,722	39,649	41,003	40,115	33,191	18,062	16,033	16,033	16,033	16,033	16,033	16,033	16,033	–	–
Plant Feed	kt	126,386	2,970	4,475	7,818	6,691	8,334	6,212	7,936	5,869	7,223	5,716	6,778	8,386	7,053	4,621	6,583	7,206	6,126	5,597	4,725	2,708	2,829	529
Heap Leach Stacked	kt	41,792	263	2,080	5,167	2,637	3,823	1,825	2,923	963	2,324	1,241	2,318	3,476	1,585	325	1,955	2,562	1,920	2,639	1,766	–	–	–
Au Feed Grade	g/t	1.26	0.88	1.22	1.12	1.67	1.47	1.22	1.99	2.21	1.38	0.66	1.21	1.62	1.88	0.76	0.85	0.77	0.81	0.67	0.72	–	–	–
Ag Feed Grade	g/t	1.63	4.90	8.28	0.59	1.11	1.02	0.77	0.47	2.05	1.36	1.46	0.92	0.59	1.22	5.48	2.44	2.38	2.11	1.55	1.65	–	–	–
Cu Feed Grade	%	0.09	0.10	0.07	0.04	0.02	0.04	0.07	0.01	0.03	0.09	0.17	0.11	0.07	0.04	0.08	0.16	0.14	0.17	0.18	0.17	–	–	–
Au Recovered	koz	1,068	20	46	98	94	112	55	108	43	64	21	54	96	61	10	36	43	36	40	30	2	0.5	–
Ag Recovered	koz	607	11	133	7	31	38	15	14	19	30	18	21	20	19	18	45	58	41	40	29	–	–	–
Cu Recovered	klb	9,875	24	87	12	218	460	379	121	85	596	640	777	786	254	96	916	1,048	996	1,418	963	–	–	–
Sulfide Plant Feed	kt	59,654	2,708	2,395	2,650	2,704	2,711	2,587	3,212	3,106	3,099	2,675	2,660	3,110	3,668	2,497	2,828	2,844	2,406	2,769	2,959	2,708	2,829	529
Au Feed Grade	g/t	2.45	3.16	2.69	2.87	3.05	2.55	3.08	2.85	3.54	2.53	2.10	1.80	2.04	3.12	2.07	2.26	2.30	2.08	2.11	1.83	1.51	1.77	1.77
Ag Feed Grade	g/t	4.08	3.82	4.71	1.96	3.59	3.70	5.45	4.89	7.57	5.94	3.74	5.89	4.18	3.66	2.95	2.11	1.48	4.12	6.24	3.81	3.88	1.90	1.92
Cu Feed Grade	%	0.13	0.08	0.07	0.09	0.07	0.13	0.16	0.14	0.15	0.18	0.23	0.20	0.18	0.13	0.14	0.12	0.11	0.17	0.17	0.10	0.07	0.10	0.11
Au Recovered	koz	4,078	248	192	222	240	187	211	244	301	215	153	128	172	327	140	175	180	134	159	157	119	146	27
Ag Recovered	koz	645	10	11	5	31	30	41	43	65	54	27	46	38	38	19	13	8	26	50	36	34	17	3
Cu Recovered	klb	71,961	–	–	–	4,083	2,565	2,706	3,747	4,230	6,159	6,567	4,811	6,115	4,143	1,339	918	623	2,617	3,667	5,910	4,187	6,370	1,204
Cu Concentrator	kt	24,939	–	–	–	1,350	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	189	–	–	–	–
Au Feed Grade	g/t	0.50	–	–	–	0.69	0.77	0.74	0.71	0.45	0.37	0.43	0.43	0.44	0.40	0.41	0.41	0.41	0.41	0.41	–	–	–	–
Ag Feed Grade	g/t	2.12	–	–	–	1.11	1.59	2.84	4.25	1.91	1.93	1.52	1.56	1.90	1.97	2.20	2.20	2.20	2.20	2.20	–	–	–	–
Cu Feed Grade	%	0.28	–	–	–	0.39	0.46	0.44	0.40	0.25	0.20	0.24	0.24	0.24	0.20	0.21	0.21	0.21	0.21	0.21	–	–	–	–
Au Recovered	koz	222	–	–	–	17	24	23	23	14	12	14	14	14	13	13	13	13	13	1	–	–	–	–
Ag Recovered	koz	135	–	–	–	8	9	9	9	9	11	10	9	9	10	10	10	10	10	1	–	–	–	–
Cu Recovered	klb	92,339	–	–	–	5,711	6,208	6,351	6,362	6,418	7,353	6,811	6,175	6,342	6,819	6,769	6,769	6,769	6,769	711	–	–	–	–
Total Feed	kt	126,386	2,970	4,475	7,818	6,691	8,334	6,212	7,936	5,869	7,223	5,716	6,778	8,386	7,053	4,621	6,583	7,206	6,126	5,597	4,725	2,708	2,829	529
Au Feed Grade	g/t	1.67	2.96	2.01	1.71	2.03	1.67	1.85	2.05	2.37	1.62	1.26	1.23	1.52	2.15	1.33	1.34	1.29	1.19	1.37	1.42	1.51	1.77	1.77
Ag Feed Grade	g/t	2.88	3.91	6.37	1.06	2.11	2.02	3.32	3.12	4.93	3.47	2.55	3.04	2.20	2.68	2.84	2.23	1.98	2.92	3.89	3.01	3.88	1.90	1.92
Cu Feed Grade	%	0.15	0.08	0.07	0.06	0.12	0.16	0.21	0.15	0.16	0.16	0.22	0.18	0.15	0.13	0.17	0.16	0.15	0.18	0.18	0.13	0.07	0.10	0.11
Au Recovered	koz	5,368	268	238	320	351	323	289	375	358	291	188	195	283	401	163	225	236	183	200	187	121	147	27
Ag Recovered	koz	1,387	21	144	12	70	77	66	66	93	95	55	76	67	67	46	68	76	77	91	65	34	17	3
Cu Recovered	klb	174,175	24	87	12	10,013	9,233	9,436	10,230	10,733	14,108	14,018	11,763	13,242	11,216	8,204	8,604	8,439	10,382	5,796	6,872	4,187	6,370	1,204

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Table 11.48 Çöpler Initial Assessment MII Case Copper Concentrator Processing Schedule

Description	Units	TOTAL	Year
			2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	2035	2036	2037	2038	2039	2040	2041	2042	2043
Cu Concentrator	kt	24,939	–	–	–	1,350	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	1,800	189	–	–	–	–
Au Feed Grade	g/t	0.50	–	–	–	0.69	0.77	0.74	0.71	0.45	0.37	0.43	0.43	0.44	0.40	0.41	0.41	0.41	0.41	0.41	–	–	–	–
Ag Feed Grade	g/t	2.12	–	–	–	1.11	1.59	2.84	4.25	1.91	1.93	1.52	1.56	1.90	1.97	2.20	2.20	2.20	2.20	2.20	–	–	–	–
Cu Feed Grade	%	0.20	–	–	–	0.23	0.19	0.19	0.19	0.19	0.22	0.20	0.19	0.19	0.20	0.20	0.20	0.20	0.20	0.20	–	–	–	–
Cu Concentrate	kt	137	–	–	–	8	9	10	10	10	11	10	9	10	10	10	10	10	10	1	–	–	–	–
Au Feed Grade	g/t	50.30	–	–	–	62.97	80.92	76.39	73.76	46.65	34.51	42.52	45.93	45.73	39.11	40.89	40.89	40.89	40.89	40.89	–	–	–	–
Ag Feed Grade	g/t	173.26	–	–	–	81.98	137.10	240.96	359.75	160.96	147.06	122.35	134.96	161.46	158.22	177.84	177.84	177.84	177.84	177.84	–	–	–	–
Cu Feed Grade	%	29.50	–	–	–	29.50	29.50	29.50	29.50	29.50	29.50	29.50	29.50	29.50	29.50	29.50	29.50	29.50	29.50	29.50	–	–	–	–
Pyrite Concentrate	kt	169	–	–	–	11	15	15	14	12	11	11	12	12	11	11	11	11	11	1	–	–	–	–
Au Feed Grade	g/t	11.15	–	–	–	13.13	13.46	13.16	13.38	10.46	9.21	10.16	10.20	10.29	9.93	10.15	10.15	10.15	10.15	10.15	–	–	–	–
Ag Feed Grade	g/t	46.94	–	–	–	20.90	27.88	50.73	79.76	44.11	47.95	35.73	36.64	44.43	49.08	53.99	53.99	53.99	53.99	53.99	–	–	–	–
Cu Feed Grade	%	0.68	–	–	–	0.66	0.50	0.52	0.55	0.68	0.84	0.74	0.67	0.68	0.78	0.76	0.76	0.76	0.76	0.76	–	–	–	–

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Table 11.49 Çöpler Initial Assessment MI Case Production Schedule

Description	Units	TOTAL	Year
	Year To		2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	2035	2036	2037	2038	2039	2040	2041	2042	2043	2044
Total Movement	kt	667,027	26,798	36,745	52,407	48,960	51,430	49,675	41,733	42,669	42,511	42,331	42,267	37,211	21,406	17,104	18,015	17,528	18,464	17,446	18,709	16,719	2,679	2,650	1,570
Waste	kt	565,995	23,828	32,015	44,655	43,362	44,843	45,329	35,351	37,617	37,544	38,899	38,010	31,086	15,958	13,928	13,928	13,928	13,928	13,928	13,928	13,928	–	–	–
Plant Feed	kt	101,032	2,970	4,730	7,752	5,598	6,587	4,346	6,382	5,053	4,967	3,432	4,257	6,125	5,448	3,176	4,086	3,599	4,536	3,517	4,781	2,790	2,679	2,650	1,570
Heap Leach Stacked	kt	39,874	263	2,080	5,613	2,847	3,795	1,631	2,986	2,544	2,467	1,284	2,024	3,316	1,418	437	1,428	994	1,885	804	2,058	–	–	–	–
Au Feed Grade	g/t	1.28	0.88	1.22	1.08	1.65	1.46	1.27	1.95	1.25	1.31	0.63	1.31	1.63	1.98	1.43	0.78	0.91	0.65	0.68	0.75	–	–	–	–
Ag Feed Grade	g/t	1.54	4.90	8.28	0.65	1.09	0.61	0.76	0.99	1.47	1.00	1.12	1.26	0.54	0.99	6.32	3.18	2.55	1.58	1.49	1.01	–	–	–	–
Cu Feed Grade	%	0.08	0.10	0.07	0.05	0.03	0.04	0.06	0.02	0.13	0.11	0.18	0.12	0.02	0.02	0.07	0.08	0.17	0.17	0.17	0.14	–	–	–	–
Au Recovered	koz	1,026	20	46	98	100	111	51	108	65	65	21	51	92	57	17	25	20	28	13	34	2	1	–	–
Ag Recovered	koz	537	11	133	7	33	23	13	28	36	25	15	24	18	14	26	43	26	29	13	20	–	–	–	–
Cu Recovered	klb	8,150	24	87	12	299	412	315	233	923	853	723	711	266	85	100	319	498	961	457	873	–	–	–	–
Sulfide Plant Feed	kt	61,158	2,708	2,650	2,138	2,750	2,792	2,715	3,396	2,509	2,500	2,148	2,232	2,809	4,030	2,739	2,658	2,605	2,651	2,713	2,723	2,790	2,679	2,650	1,570
Au Feed Grade	g/t	2.32	3.16	2.60	2.22	2.65	2.59	2.32	2.60	2.99	2.34	1.60	1.56	1.88	3.02	1.85	1.91	2.27	1.90	2.23	2.22	2.17	2.12	2.13	2.13
Ag Feed Grade	g/t	4.81	3.82	6.52	4.33	5.13	3.44	7.16	4.13	5.34	2.80	5.28	4.03	4.85	3.75	4.77	6.21	5.32	4.32	5.12	5.09	4.92	4.91	4.96	4.96
Cu Feed Grade	%	3.81	–	4.30	3.99	4.25	4.12	4.42	4.21	4.15	4.14	4.88	4.84	4.68	4.89	4.88	4.57	4.27	4.20	1.14	1.29	1.86	4.20	4.20	4.20
Au Recovered	koz	4,187	248	194	229	216	201	164	239	215	168	98	104	158	363	149	151	175	150	178	180	177	167	165	98
Ag Recovered	koz	822	10	11	5	45	29	58	38	37	20	29	28	44	49	42	53	45	37	45	45	44	42	42	25
Cu Recovered	klb	100,365	–	–	–	6,415	6,402	4,659	4,884	7,481	7,607	3,234	6,262	3,423	2,903	2,571	3,991	5,547	6,669	5,421	5,387	5,264	4,725	4,723	2,798
Cu Concentrator	kt	9,987	–	–	–	1,350	1,800	1,800	1,800	1,221	1,800	216	–	–	–	–	–	–	–	–	–	–	–	–	–
Au Feed Grade	g/t	0.48	–	–	–	0.69	0.77	0.74	0.29	0.27	0.15	0.15	–	–	–	–	–	–	–	–	–	–	–	–	–
Ag Feed Grade	g/t	2.09	–	–	–	1.11	1.59	2.84	2.14	1.59	2.75	2.75	–	–	–	–	–	–	–	–	–	–	–	–	–
Cu Feed Grade	%	0.21	–	–	–	0.23	0.19	0.19	0.22	0.24	0.23	0.23	–	–	–	–	–	–	–	–	–	–	–	–	–
Au Recovered	koz	85	–	–	–	17	24	23	9	6	5	1	–	–	–	–	–	–	–	–	–	–	–	–	–
Ag Recovered	koz	57	–	–	–	8	9	9	11	8	11	1	–	–	–	–	–	–	–	–	–	–	–	–	–
Cu Recovered	klb	39,322	–	–	–	5,711	6,208	6,351	7,258	5,362	7,529	903	–	–	–	–	–	–	–	–	–	–	–	–	–
Total Feed	kt	111,019	2,970	4,730	7,752	6,948	8,387	6,146	8,182	6,274	6,767	3,648	4,257	6,125	5,448	3,176	4,086	3,599	4,536	3,517	4,781	2,790	2,679	2,650	1,570
Au Feed Grade	g/t	1.78	2.96	2.00	1.39	1.86	1.69	1.58	1.86	1.76	1.38	1.17	1.44	1.75	2.75	1.79	1.52	1.90	1.38	1.87	1.59	2.17	2.12	2.13	2.13
Ag Feed Grade	g/t	3.39	3.91	7.29	1.66	2.69	1.76	4.20	2.55	3.04	2.13	3.67	2.71	2.52	3.03	4.98	5.15	4.56	3.18	4.29	3.33	4.92	4.91	4.96	4.96
Cu Feed Grade	%	2.14	0.01	2.44	1.14	1.74	1.43	2.02	1.81	1.76	1.63	2.95	2.59	2.16	3.62	4.22	3.00	3.14	2.53	0.92	0.80	1.86	4.20	4.20	4.20
Au Recovered	koz	5,298	268	240	327	333	337	239	356	286	238	120	155	250	420	166	176	195	178	191	213	179	167	165	98
Ag Recovered	koz	1,416	21	144	12	86	61	81	76	81	55	46	52	62	63	68	96	70	66	58	65	44	42	42	25
Cu Recovered	klb	147,837	24	87	12	12,426	13,021	11,325	12,376	13,765	15,989	4,860	6,973	3,689	2,988	2,671	4,309	6,045	7,630	5,878	6,260	5,264	4,725	4,723	2,798

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Table 11.50 Çöpler Initial Assessment MI Case Copper Concentrator Processing Schedule

Description	Units	TOTAL	Year
			2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	2035	2036	2037	2038	2039	2040	2041	2042	2043	2044
Cu Concentrator	kt	9,987	–	–	–	1,350	1,800	1,800	1,800	1,221	1,800	216	–	–	–	–	–	–	–	–	–	–	–	–	–
Au Feed Grade	g/t	0.48	–	–	–	0.69	0.77	0.74	0.29	0.27	0.15	0.15	–	–	–	–	–	–	–	–	–	–	–	–	–
Ag Feed Grade	g/t	2.09	–	–	–	1.11	1.59	2.84	2.14	1.59	2.75	2.75	–	–	–	–	–	–	–	–	–	–	–	–	–
Cu Feed Grade	%	0.21	–	–	–	0.23	0.19	0.19	0.22	0.24	0.23	0.23	–	–	–	–	–	–	–	–	–	–	–	–	–
Cu Concentrate	kt	57	–	–	–	8	9	10	11	8	11	1	–	–	–	–	–	–	–	–	–	–	–	–	–
Au Grade	g/t	45.86	–	–	–	62.97	80.92	76.39	26.84	23.87	13.40	13.40	–	–	–	–	–	–	–	–	–	–	–	–	–
Ag Grade	g/t	163.14	–	–	–	81.98	137.10	240.96	164.79	114.96	205.65	205.65	–	–	–	–	–	–	–	–	–	–	–	–	–
Cu Grade	%	29.50	–	–	–	29.50	29.50	29.50	29.50	29.50	29.50	29.50	–	–	–	–	–	–	–	–	–	–	–	–	–
Pyrite Concentrate	kt	68	–	–	–	11	15	15	10	7	9	1	–	–	–	–	–	–	–	–	–	–	–	–	–
Au Grade	g/t	10.59	–	–	–	13.13	13.46	13.16	7.60	7.36	4.58	4.58	–	–	–	–	–	–	–	–	–	–	–	–	–
Ag Grade	g/t	46.04	–	–	–	20.90	27.88	50.73	57.02	43.30	85.82	85.82	–	–	–	–	–	–	–	–	–	–	–	–	–
Cu Grade	%	0.72	–	–	–	0.66	0.50	0.52	0.89	0.99	1.08	1.08	–	–	–	–	–	–	–	–	–	–	–	–	–

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Table 11.51 Initial Assessment MII Case Cash Flow

Cash Flow Statement ($M)	TOTAL	Year
		2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	2035	2036	2037	2038	2039	2040	2041	2042	2043	2044+
Heap Leach - Gold Revenue	1,737	36	80	167	157	180	87	173	68	102	34	86	154	98	16	58	69	57	65	47	3	1	–	–
Sulfide Plant - Gold Revenue	6,999	447	334	379	429	338	375	427	505	363	267	226	298	543	244	302	309	236	256	252	191	234	44	–
By-Product Revenue	637	5	5	5	40	33	33	36	38	50	49	42	46	40	29	31	30	37	22	25	15	22	4	–
Net Revenue	9,373	488	419	552	625	550	496	637	612	515	350	354	499	680	289	390	408	329	342	324	209	257	48	–
Realisation Costs
Freight and Refining	68	1	1	1	4	5	5	5	4	4	4	4	4	5	3	4	4	4	2	1	0	1	0	–
Royalties	486	23	27	19	29	36	30	25	36	33	26	14	15	27	39	11	18	19	15	14	13	7	10	1
Total - Realisation Costs	554	24	28	21	34	41	35	30	41	38	30	18	20	32	42	15	22	23	16	16	13	7	10	1
Operating Costs
Mining	1,255	47	57	86	74	78	77	70	72	76	77	79	73	48	42	48	51	51	51	50	47	–	–	–
Processing - Heap Leach	550	13	21	64	33	44	25	35	17	30	20	30	41	23	12	27	32	26	33	25	–	–	–	–
Processing - Sulfide Plant	2,084	115	83	91	94	94	89	111	107	107	92	92	107	127	86	98	98	83	96	103	94	98	18	–
Processing - Cu Concentrator	182	–	–	–	10	13	13	13	13	13	13	13	13	13	13	13	13	13	1	–	–	–	–	–
Site Support	453	32	30	29	27	27	27	27	23	23	19	19	19	19	19	19	19	19	19	19	9	9	–	–
G&A	100	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	0	–	–
Total - Operating Costs	4,624	212	196	275	242	261	237	262	237	253	227	238	259	236	178	211	219	197	205	202	154	107	18	–
Operating Surplus / (Deficit)	4,194	252	196	256	350	249	224	345	334	224	93	98	221	413	68	165	167	109	121	106	42	143	20	-1
Capital Costs
Growth	357	4	69	192	22	–	–	–	–	–	–	18	18	18	18	–	–	–	–	–	–	–	–	–
Sustaining	458	32	55	52	12	49	49	12	12	30	30	12	12	12	12	12	12	11	11	10	10	10	–	–
Closure	114	–	–	–	–	–	–	–	–	–	–	7	13	13	13	–	2	3	1	1	1	1	1	58
Working and Other	-37	-12	-21	-4	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–
Total - Capital Costs	893	25	103	239	34	49	49	12	12	30	30	36	42	42	42	12	14	14	12	11	11	11	1	58
Net Cash Flow Before Tax	3,301	227	93	17	315	200	175	333	322	195	63	62	178	371	26	153	153	95	108	95	30	131	18	–60
Tax	344	6	3	4	6	8	7	12	11	25	2	6	31	74	6	28	29	17	20	18	5	26	1	–
Net Cash Flow After Tax	2,958	221	90	13	310	192	168	322	311	170	61	56	147	296	20	124	125	78	88	77	25	106	17	-60

2044+ covers the period from 2044–2052

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Table 11.52 Initial Assessment MI Case Cash Flow

Cash Flow Statement ($M)	TOTAL	Year
		2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	2035	2036	2037	2038	2039	2040	2041	2042	2043	2044
Gold Revenue
Heap Leach	1,671	36	80	168	167	178	82	173	104	105	34	81	147	92	27	40	33	44	21	54	3	1	–	–
Sulfide Plant	6,733	447	334	379	355	314	259	376	339	264	154	162	246	577	237	238	277	236	283	283	283	266	264	157
Cu Concentrator	137	–	–	–	28	39	38	15	9	7	1	–	–	–	–	–	–	–	–	–	–	–	–	–
By-Product Revenue	549	5	5	5	49	46	40	44	49	56	17	25	14	11	11	17	22	27	21	23	19	17	17	10
Net Revenue	9,089	488	419	553	599	577	419	608	502	432	206	268	407	681	275	295	332	308	325	360	305	284	281	167
Realisation Costs
Freight and Refining	44	1	1	1	4	5	4	5	4	4	1	1	2	2	1	1	1	1	1	1	1	1	1	0
Royalties	451	23	27	19	29	34	32	19	34	26	22	4	10	21	38	9	11	14	12	13	16	14	12	13
Total - Realisation Costs	495	24	28	21	34	39	36	23	37	30	23	6	12	23	39	10	12	15	13	15	16	14	13	13
Operating Costs
Mining	1,161	47	57	86	74	78	77	67	69	71	70	71	65	41	35	38	39	42	41	45	42	–	–	4
Ore Rehandle	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–
Processing - Heap Leach	533	13	21	68	35	43	23	36	32	31	20	27	39	21	13	23	18	26	16	27	–	–	–	–
Processing - Sulfide Plant	2,212	115	92	74	105	110	107	131	96	100	76	77	97	140	95	92	90	92	94	94	97	93	92	54
Site Support	453	32	30	29	27	27	27	27	23	23	19	19	19	19	19	19	19	19	19	19	9	9	–	–
Total - Site Operating Costs	4,359	207	200	256	241	258	234	261	219	225	185	195	221	221	163	172	166	179	170	186	147	101	92	59
Grade Control	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–
Corporate Costs	100	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	5	0	–	–
Total - Operating Costs	4,459	212	205	261	246	263	239	266	224	230	190	200	226	226	168	177	171	184	175	191	152	101	92	59
Operating Surplus / (Deficit)	4,136	252	186	271	319	275	144	319	240	172	-7	63	170	431	69	107	149	108	138	154	137	169	177	95
Capital Costs
Growth	357	4	69	192	22	–	–	–	–	–	–	18	18	18	18	–	–	–	–	–	–	–	–	–
Sustaining	457	32	55	52	12	49	49	12	12	29	29	12	12	12	12	12	12	11	11	10	10	10	–	–
Closure	63	–	–	–	–	–	–	–	–	–	–	7	13	13	13	–	2	3	1	1	1	1	1	7
Working and Other	-37	-12	-21	-4	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–	–
Total - Capital Costs	841	25	103	239	34	49	49	12	12	29	29	36	42	42	42	12	14	14	12	11	11	11	1	7
Net Cash Flow Before Tax	3,295	227	83	32	285	225	95	307	228	142	-36	27	127	389	26	95	135	94	125	142	125	157	175	87
Tax	328	6	3	4	5	9	3	11	8	5	–	–	4	52	6	17	25	17	24	27	24	31	33	15
Net Cash Flow After Tax	2,967	221	80	28	280	217	91	296	220	137	-36	27	123	337	20	79	110	77	102	115	101	127	143	72

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11.6.1.2 Initial Assessment Summary Results – MII Case

This section describes the results of the Initial Assessment economic analysis including Inferred Mineral Resources (MII Case). The economic model results are based on Mineral Resources that, unlike Mineral Reserves, do not have demonstrated economic viability.

The Initial Assessment MII Case production is 126.4 Mt at 1.67 g/t Au this includes 24.9 Mt of feed to the copper concentrator at 0.5 g/t Au and 0.20% Cu. The metal production in the Initial Assessment MII Case is 5.4 Moz of gold and 164 Mlb of copper. The Mineral Resource in the Initial Assessment MII Case contains 27% Inferred Mineral Resource. The production schedule was prepared so that only Measured and Indicated Mineral Resources are processed in the first four years of the production schedule and there is only 10% Inferred Mineral Resource in the process feed fifth year of production. An additional cash flow analysis has been prepared using only Measured and Indicated Mineral Resources to show the impact of Inferred Mineral Resource on the economic analysis (Refer to Section 11.6.1.3).

The increase in total production relative to the CDMP21TRS Reserves Case is from expansion of the Çöpler pit deeper than the current Reserve pit design. As well as the copper concentrator feed that is captured in the Initial Assessment pit shell additional oxide and sulfide Mineral Resource is exposed resulting in increased feed for both the heap leach oxide and sulfide processing facilities.

The Initial Assessment MII Case results include:

•After-tax NPV at a 5% real discount rate is $2.00 billion

•Mine life of 22 years

The Initial Assessment MII Case shows an average AISC of $924/oz gold.

The CDMP21TRS Reserves Case after-tax NPV at a 5% discount rate is $1.73 billion. Incremental analysis suggest that the impact from the addition of the copper concentrator and NaSH circuit to increase the after-tax NPV5% by $273M demonstrates the economic potential of the Çöpler Mineral Resource.

The Initial Assessment has been prepared to demonstrate economic potential of the Mineral Resources at the Çöpler deposit. The Initial Assessment is preliminary in nature, it includes Inferred Mineral Resources that are considered too speculative geologically to have modifying factors applied to them that would enable them to be categorised as Mineral Reserves, and there is no certainty that this economic assessment will be realised.

Key results of the Initial Assessment MII Case economic analysis are shown in Table 11.53. The after-tax cash flow is shown in Figure 11.23. The sulfide and oxide production profiles are shown in Figure 11.24 and gold production is shown in Figure 11.25. The NPV results for before and after-tax over a range of discount rates is shown in Table 11.54.

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Table 11.53 Initial Assessment Results Summary – MII Case

Item	Unit	Initial Assessment MII Case
Oxide Processed
Heap Leach	kt	41,792
Au Feed Grade	g/t	1.26
Sulfide Processed
Milled	kt	59,654
Au Feed Grade	g/t	2.45
Cu Concentrator Processed
Milled	kt	24,939
Au Feed Grade	g/t	0.50
Cu Feed Grade	%	0.20
Total Gold Produced
Oxide – Gold	koz	1,068
Sulfide – Gold	koz	4,078
Cu Concentrator – Gold	koz	222
Total – Gold	koz	5,368
Total – Copper	Mlb	164
5-Year Annual Average
Average Gold Produced	kozpa	300
Free Cash Flow	$Mpa	165
Total Cash Costs (CC)	$/oz gold	761
All-In Sustaining Costs (AISC)	$/oz gold	938
Key Financial Results
Life-of-Mine (LOM) CC	$/oz gold	783
LOM AISC	$/oz gold	924
Site Operating Costs	$/t processed	43.79
After-Tax NPV5%	$M	2,004
Mine Life	years	22

5-Year annual average is for the period 1 January 2022 through 31 December 2026

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Figure 11.23 Initial Assessment After-Tax Cash Flow – MII Case

OreWin, 2022

Table 11.54 Initial Assessment Before and After-Tax NPV – MII Case

Discount Rate	Before-Tax NPV ($M)	After-Tax NPV ($M)
Undiscounted	3,301	2,958
5%	2,194	2,004
10%	1,571	1,457
12%	1,398	1,304

21007CDMP21TRS220927Rev0.docx Page 186 of 303

Figure 11.24 Initial Assessment Processing – MII Case

OreWin, 2022

Figure 11.25 Initial Assessment Gold Production – MII Case

OreWin, 2022

21007CDMP21TRS220927Rev0.docx Page 187 of 303

Costs, metal prices, taxation, and royalty assumptions used in the Initial Assessment MII Case economic analysis were the same as in the CDMP21TRS Reserves Case.

The estimates of cash flows have been prepared on a real basis with a base date of Q4’21 and a mid-year discounting is used to calculate NPV. Comparison of the initial years of the Initial Assessment and the MI Case showed only 1.4% of the material processed in the first nine years of the Initial Assessment is Inferred Mineral Resource. Most of the Inferred material is processed in years 10 to 20 and does not exceed 50% of the total processing in any one year. This is shown in Figure 11.26.

Figure 11.26 Initial Assessment MII Case Gold Production

OreWin, 2022

11.6.1.3 Initial Assessment Summary Results – MI Case

This section describes the results of the Initial Assessment economic analysis excluding Inferred Mineral Resources (MI Case). The economic model results are based on Mineral Resources that, unlike Mineral Reserves, do not have demonstrated economic viability.

21007CDMP21TRS220927Rev0.docx Page 188 of 303

The Initial Assessment MI Case production is 111.0 Mt at 1.78 g/t Au this includes 10.0 Mt of feed to the copper concentrator at 0.5 g/t Au and 0.21% Cu. The metal production in the Initial Assessment MI Case is 5.2 Moz of gold and 148 Mlb of copper. The production schedule was prepared so that only Measured and Indicated Mineral Resources are processed to show the impact of Inferred Mineral Resource on the economic analysis in the base case (MII-Case) analysis.

The Initial Assessment MI Case results include:

•After-tax NPV at a 5% real discount rate is $1.87 billion

•Mine life of 22 years

The Initial Assessment Case shows an average AISC of $921/oz gold.

The after-tax NPV5% for this MI Case only analysis is $1,867M this is a reduction of $137M from the MII Case. To mitigate this, additional resource development work and studies will need to be carried out to convert the Inferred Mineral Resources to either Indicated or Measured Mineral Resources or find an alternative feed source.

The Initial Assessment has been prepared to demonstrate the economic potential of the Mineral Resources at the Çöpler deposit. The Initial Assessment is preliminary in nature, it includes Inferred Mineral Resources that are considered too speculative geologically to have modifying factors applied to them that would enable them to be categorised as Mineral Reserves, and there is no certainty that this economic assessment will be realised.

Key results of the Initial Assessment MI Case economic analysis are shown in Table 11.55. The after-tax cash flow is shown in Figure 11.27. The sulfide and oxide production profiles are shown in Figure 11.28 and gold production is shown in Figure 11.29. The NPV results for before and after-tax over a range of discount rates is shown in Table 11.56.

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Table 11.55 Initial Assessment Results Summary – MI Case

Item	Unit	Initial Assessment – MI Case
Oxide Processed
Heap Leach	kt	39,874
Au Feed Grade	g/t	1.28
Sulfide Processed
Milled	kt	61,158
Au Feed Grade	g/t	2.32
Cu Concentrator Processed
Milled	kt	9,987
Au Feed Grade	g/t	0.48
Cu Feed Grade	%	0.21
Total Gold Produced
Oxide – Gold	koz	1,026
Sulfide – Gold	koz	4,135
Cu Concentrator – Gold	koz	85
Total – Gold	koz	5,247
Total – Copper	Mlb	148
5-Year Annual Average
Average Gold Produced	kozpa	297
Free Cash Flow	$Mpa	165
Total Cash Costs (CC)	$/oz gold	753
All-In Sustaining Costs (AISC)	$/oz gold	932
Key Financial Results
Life-of-Mine (LOM) CC	$/oz gold	775
LOM AISC	$/oz gold	921
Site Operating Costs	$/t processed	44.13
After-Tax NPV5%	$M	1,867
Mine Life	years	22

5-Year annual average is for the period 1 January 2022 through 31 December 2026

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Figure 11.27 Initial Assessment After-Tax Cash Flow – MI Case

OreWin, 2022

Table 11.56 Initial Assessment Before and After-Tax NPV – MI Case

Discount Rate	Before-Tax NPV ($M)	After-Tax NPV ($M)
Undiscounted	3,235	2,907
5%	2,029	1,867
10%	1,414	1,325
12%	1,254	1,180

Figure 11.28 Initial Assessment Processing – MI Case

OreWin, 2022

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Figure 11.29 Initial Assessment Gold Production – MI Case

OreWin, 2022

Costs, metal prices, taxation, and royalty assumptions used in the Initial Assessment Case economic analysis were the same as in the CDMP21TRS Reserves Case.

The estimates of cash flows have been prepared on a real basis with a base date of Q4’21 and a mid-year discounting is used to calculate NPV.

11.7 Mineral Resources Statement

Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

Mineral Resources are reported exclusive of Mineral Reserves and have been summarised by project, Mineral Resource classification, and oxidation state in Table 11.57. Mineral Resources are presented showing only the SSR attributable proportion. Mineral Reserves are discussed and disclosed in the separate CDMP21TRS.

Table 11.58 shows the cut-off values, metallurgical recoveries, and SSR ownership percentage associated with the Mineral Resources.

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Table 11.57 Summary of CDMP21TRS Mineral Resources Estimates Exclusive of Mineral Reserves as at 31 December 2021 Based on $1,750/oz Gold Price

Mineral Resource Classification		SSR Tonnage<br>(kt)		Grades						Contained Metal
	Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)		Gold<br>(koz)		Silver<br>(koz)			Copper<br>(klb)
Çöpler Mine – Oxide
Measured		65		1.39		4.67		0.16		3	10		225
Indicated		21,739		0.84		2.30		0.16		589	1,610		77,646
Measured + Indicated		21,803		0.84		2.31		0.16		592	1,619		77,871
Inferred		28,017		0.90		6.87		0.13		813	6,192		78,353
Çöpler Mine – Sulfide
Measured		121		0.83		3.72		0.18		3	15		472
Indicated		37,667		1.06		3.66		0.19		1,286	4,428		158,692
Measured + Indicated		37,788		1.06		3.66		0.19		1,289	4,442		159,164
Inferred		39,838		1.24		13.60		0.17		1,585	17,418		145,512
Çakmaktepe – Oxide
Measured		–		–		–		–		–	–		–
Indicated		1,671		1.55		8.33		–		83	447		–
Measured + Indicated		1,671		1.55		8.33		–		83	447		–
Inferred		602		0.85		4.04		–		16	78		–
Ardich – Oxide
Measured		2,101		1.46		3.06		0.02		99	207		350
Indicated		6,786		0.99		2.45		0.00		215	534		1,147
Measured + Indicated		8,887		1.10		2.59		0.01		314	741		1,497
Inferred		8,980		1.29		3.17		0.02		372	914		3,306
Ardich – Sulfide (Incl. sulfide and sulfide-with-Cu)
Measured		180		5.76		8.41		0.04		34	49		34
Indicated		953		2.07		3.61		0.01		63	111		383
Measured + Indicated		1,133		2.59		4.37		0.02		97	159		416
Inferred		2,209		2.64		4.47		0.01		195	317		367
Bayramdere – Oxide
Measured		–		–		–		–		–	–		–
Indicated		72		2.34		20.82		–		5	48		–
Measured + Indicated		72		2.34		20.82		–		5	48		–
Inferred		4		2.17		19.95		–		0	3		–
CDMP21 Mineral Resources – Oxide Subtotal
Measured		2,166		1.46		3.11		0.02		102	217		575
Indicated		30,267		0.92		2.71		0.12		894	2,640		78,793
Measured + Indicated		32,433		0.95		2.74		0.11		995	2,857		79,368
Inferred		37,603		0.99		5.95		0.10		1,202	7,188		81,659
CDMP21 Mineral Resources – Sulfide Subtotal
Measured		301		3.88		6.52		0.08		38	63		506
Indicated		38,620		1.09		3.66		0.19		1,349	4,538		159,075
Measured + Indicated		38,921		1.11		3.68		0.19		1,387	4,602		159,580
Inferred		42,047		1.32		13.12		0.16		1,780	17,735		145,879
CDMP21 MINERAL RESOURCES – OVERALL TOTAL (Exclusive of Mineral Reserves)
Measured		2,467		1.76		3.53		0.02		139	280		1,081
Indicated		68,887		1.01		3.24		0.16		2,243	7,178		237,867
Measured + Indicated		71,354		1.04		3.25		0.15		2,382	7,458		238,948
Inferred		79,650		1.16		9.73		0.13		2,982	24,923		227,538

1. Mineral Resources are reported based on 31 December 2021 topography surface.

2. Mineral Resources are reported exclusive of Mineral Reserves.

5. Sulfide definitions: At Ardich, sulfide is comprised of standard sulfide material (≥2% total sulfur) and sulfide-with-Cu material (sulfide with Cu>0.10%).

7. All Mineral Resources in the CDMP21TRS were assessed for reasonable prospects for eventual economic extraction by reporting only material that fell within conceptual pit shells ($1,400/oz for gold and $19/oz for silver for Bayramdere, and $1,750/oz for gold, $22/oz for silver for all other projects).

8. The point of reference for Mineral Resources is the point of feed into the processing facility.

9. Tonnage is metric tonnes, ounces represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

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Table 11.58 Summary of Cut-off Values, Metallurgical Recoveries, and SSR Ownership of CDMP21TRS Mineral Resources Estimates Exclusive of Mineral Reserve (as at 31 December 2021) Based on Gold Price $1,750/oz, Silver Price $22.00/oz Ag and Copper Price $3.95/lb

Mineral Resource Classification		SSR Tonnage<br>(kt)		Grades				Cut-off Value/s	Metallurgical Recovery (%)	SSR Ownership (%)
	Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)
Çöpler Mine – Oxide
Measured		65		1.39		4.67	0.16	0.19–0.24 g/t Au	62.3–78.4	80
Indicated		21,739		0.84		2.30	0.16
Measured + Indicated		21,803		0.84		2.31	0.16
Inferred		28,017		0.90		6.87	0.13
Çöpler Mine – Sulfide
Measured		121		0.83		3.72	0.18	34.88/t NSR or>0.10% Cu and 7.68/t NSR	Au 55–91<br><br>Ag 10–45<br><br>Cu 84–98	80
Indicated		37,667		1.06		3.66	0.19
Measured + Indicated		37,788		1.06		3.66	0.19
Inferred		39,838		1.24		13.60	0.17
Çakmaktepe – Oxide
Measured		–		–		–	–	0.36–0.76 g/t Au	38.0–80.0	50
Indicated		1,671		1.55		8.33	–
Measured + Indicated		1,671		1.55		8.33	–
Inferred		602		0.85		4.04	–
Ardich – Oxide
Measured		2,101		1.46		3.06	0.02	0.23–0.41 g/t Au	40.0–73.0	75
Indicated		6,786		0.99		2.45	0.00
Measured + Indicated		8,887		1.10		2.59	0.01
Inferred		8,980		1.29		3.17	0.02
Ardich – Sulfide (Incl. sulfide and sulfide-with-Cu)
Measured		180		5.94		8.41	0.01	36.25/t NSR or>0.10% Cu and 9.05/t NSR	Au 55–91<br><br>Ag 10–45<br><br>Cu 84–98	78
Indicated		953		2.05		3.61	0.02
Measured + Indicated		1,133		2.67		4.37	0.02
Inferred		2,209		2.75		4.47	0.01
Bayramdere – Oxide
Measured		–		–		–	–	0.35–0.50 g/t Au	75	50
Indicated		72		2.34		20.82	–
Measured + Indicated		72		2.34		20.82	–
Inferred		4		2.17		19.95	–
CDMP21 Mineral Resources – Oxide Subtotal
Measured		2,166		1.46		3.11	0.01	As Above	As Above	75
Indicated		30,267		0.92		2.71	0.12
Measured + Indicated		32,433		0.95		2.74	0.11
Inferred		37,603		0.99		5.95	0.10
CDMP21 Mineral Resources – Sulfide Subtotal
Measured		301		3.88		6.52	0.08	As Above	As Above	78
Indicated		38,620		1.09		3.66	0.19
Measured + Indicated		38,921		1.11		3.68	0.19
Inferred		42,047		1.32		13.12	0.16
CDMP21 MINERAL RESOURCES – OVERALL TOTAL (Exclusive of Mineral Reserves)
Measured		2,467		1.76		3.53	0.02	As Above	As Above	76
Indicated		68,887		1.01		3.24	0.16
Measured + Indicated		71,354		1.04		3.25	0.15
Inferred		79,650		1.16		9.73	0.13

All values are in US Dollars.

1. Mineral Resources are reported based on 31 December 2021 topography surface.

2. Mineral Resources are reported exclusive of Mineral Reserves.

5. Sulfide definitions: At Ardich, sulfide is comprised of standard sulfide material (≥2% total sulfur) and sulfide-with-Cu material (sulfide with Cu>0.10%).

6. At Çöpler and Ardich: sulfide cut-off uses an NSR value in $/t based on gold price $1,750/oz, silver price $22.00/oz, and copper price $3.95/lb with allowances for payability, deductions, transport, and royalties.

8. The point of reference for Mineral Resources is the point of feed into the processing facility.

9. Tonnage is metric tonnes and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

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11.8 Comparison with Previous Estimates

The 2021 Mineral Resource inventory is detailed in Table 11.57. Mineral Resources are presented showing only the SSR attributable proportion.

For comparison, a summary of the overall 2020 Mineral Resource inventory is shown in Table 11.59. This comparison compares Mineral Resources exclusive of Mineral Reserves. Mineral Resources are presented showing only the SSR attributable proportion. Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

Table 11.59 EOY 2020 Mineral Resources Summary – Exclusive of Mineral Reserves (as at 31 December 2020)

Mineral Resource Classification		EOY 2020 Mineral Resources Total (Exclusive of Mineral Reserves)
	Tonnage (kt)		Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)		Gold<br>(koz)		Silver<br>(koz)		Copper<br>(klb)
Measured		4,563		1.70		0.20		0.00		250		29		186
Indicated		55,424		1.34		2.99		0.04		2,385		5,334		47,907
Measured + Indicated		59,987		1.37		2.78		0.04		2,634		5,363		48,094
Inferred		58,224		1.30		9.25		0.06		2,442		17,318		75,917

Mineral Resources are reported showing only the SSR attributable tonnage portion.

The factors contributing to the differences between the 2021 Mineral Resources and the previous Mineral Resources reported as at 31 December 2020 are as follows:

Ardich

•A significant proportion of the Measured plus Indicated (M+I) tonnage increase in 2021 is attributed to the updated Ardich model, which incorporates 194 additional drillholes (233 holes in 2020 vs. 427 holes in 2021), with the net result of significantly improving confidence in the interpretation, thereby increasing M+I inventory.

•There was no Mineral Reserve declared for Ardich in 2020. A maiden Mineral Reserve has been declared for Ardich in 2021, and this depletes the report of Mineral Resources exclusive of Mineral Reserves.

Other

•A drop in cut-off grade and the inclusion of copper extraction has resulted in a larger conceptual pit shell for Çöpler that contains additional volume above the cut-off.

•Review of metallurgical recoveries.

•Depletion through mining since 31 December 2020.

There has been an 28% increase in tonnage above the cut-off across all combined Mineral Resource categories, with a corresponding 6% increase in contained gold.

11.9 QP Opinion

The CDMP21TRS QP has not identified any relevant technical and/or economic factors that require resolution with regards to the Mineral Resource estimate.

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Based on the Initial Assessment in this Report, the QP concludes that there are reasonable prospects for economic extraction of the Mineral Resources. The Mineral Resources were prepared in accordance with the definitions and standards in S-K 1300.

11.10 Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects

The Mineral Resources reported in the CDMP21TRS are suitable for reporting as Mineral Resources using the Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

The CDMP21TRS includes an Initial Assessment Case that uses Measured, Indicated, and Inferred Mineral Resources to examine the impact of adding two new processing options to extract copper. The Initial Assessment Case is considered to be the same as a Preliminary Economic Assessment (PEA) under NI 43-101. This is because even though Mineral Reserves have been previously disclosed at the Project, the change in processing method would allow a PEA to be published.

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12 MINERAL RESERVES ESTIMATES

12.1 Summary

Open pit mining at the Çöpler project is carried out by a mining contractor and managed by Anagold. The mining method is a conventional open pit method with drill and blast to facilitate extraction utilising excavators and trucks. Anagold currently operates a sulfide process plant and an oxide heap leach facility. Costs are based on actual operational costs and the Anagold budget assumptions.

The Mineral Reserves were developed based on mine planning work completed in 2021 and estimated based on an end of September 2021 topography surface. Çöpler oxide ore cut-off grades vary from 0.47–0.59 g/t Au. The Çöpler sulfide ore cut-off grade is 1.05 g/t Au. Çakmaktepe oxide cut-off grades vary from 0.52–0.69 g/t Au. There is no Çakmaktepe sulfide Mineral Reserve. Average oxide gold recoveries are 61% and average sulfide gold recoveries are 91%.

The cut-off grades for the Mineral Reserves estimates are based on a gold price of $1,350/oz. There are no credits for silver or copper in the cut-off grade calculations. Economic analysis has been carried out using long-term metal prices of $1,600/oz gold, $20.25/oz silver, and $3.05/lb copper, and average metal prices of $1,658/oz gold, $21.55/oz silver, and $2.95/lb copper. Metal prices were selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The metal prices are representative of the range of price estimates publicly reported for Mineral Reserve cut-offs.

12.2 Mineral Reserves Statement

The Mineral Reserves statement is shown in Table 12.1. Mineral Reserves have been classified in accordance with S-K 1300 and were estimated by Bernard Peters BEng (Mining), FAusIMM (201743), employed by OreWin Pty Ltd as Technical Director – Mining. Mineral Reserves are presented on a SSR attributable tonnage portion basis and have an effective date of 31 December 2021.

Table 12.2 shows the cut-off values, metallurgical recoveries, and SSR ownership percentage associated with the Mineral Reserves.

The CDMP21TRS Reserve Case is at a feasibility level of study. The Mineral Resource estimates have been reported in the CDMP21TRS. The Mineral Resource models include dilution. Measured Mineral Resources were converted to Proven Mineral Reserves, and Indicated Mineral Resources were converted to Probable Mineral Reserves. Inferred Mineral Resources were treated as waste and were not converted to Mineral Reserve. The Çöpler Mineral Reserve has been demonstrated to be viable by the CDMP21TRS.

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Table 12.1 Summary of CDMP21TRS Mineral Reserves Estimates (as at 31 December 2021) Based on $1,350/oz Gold Price

Mineral Reserve Classification		SSR Tonnage<br>(kt)		Grades						Contained Metal
	Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)		Gold<br>(koz)		Silver<br>(koz)		Copper<br>(klb)
Çöpler Mine – Oxide
Proven Mineral Reserve		–		–		–		–		–		–	–
Probable Mineral Reserve		1,763		1.22		11.17		0.13		69		633	5,043
Probable – Stockpile		–		–		–		–		–		–	–
Total Mineral Reserve		1,763		1.22		11.17		0.13		69		633	5,043
Çöpler Mine – Sulfide
Proven Mineral Reserve		326		2.02		6.69		–		21		70	–
Probable Mineral Reserve		28,662		2.13		4.96		–		1,964		4,571	–
Probable – Stockpile		9,974		2.25		–		–		720		0	–
Total Mineral Reserve		38,962		2.16		3.70		–		2,705		4,641	–
Çakmaktepe Mine – Oxide
Proven Mineral Reserve		–		–		–		–		–		–	–
Probable Mineral Reserve		137		1.26		10.91		–		6		48	–
Probable – Stockpile		6		2.69		–		–		1		–	–
Total Mineral Reserve		143		1.32		10.49		–		6		48	–
Ardich – Oxide Reserve
Proven Mineral Reserve		5,279		1.73		2.00		0.00		293		339	158
Probable Mineral Reserve		10,287		1.73		1.91		0.01		572		631	2,231
Probable – Stockpile		–		–		–		–		–		–	–
Total Mineral Reserve		15,566		1.73		1.94		0.01		866		970	2,389
Ardich – Sulfide
Proven Mineral Reserve		1,428		5.67		11.08		–		260		509	–
Probable Mineral Reserve		1,632		3.10		4.15		–		163		218	–
Probable – Stockpile		–		–		–		–		–		–	–
Total Mineral Reserve		3,060		4.30		7.38		–		423		726	–
CDMP21 Mineral Reserves – Oxide Subtotal
Proven Mineral Reserve		5,279		1.73		2.00		0.00		293		339	158
Probable Mineral Reserve		12,187		1.65		3.35		0.03		647		1,312	7,274
Probable – Stockpile		6		2.83		–		–		1		0	0
Total Mineral Reserve		17,472		1.68		2.94		0.02		941		1,651	7,432
CDMP21 Mineral Reserves – Sulfide Subtotal
Proven Mineral Reserve		1,754		4.99		10.26		–		281		579	–
Probable Mineral Reserve		30,294		2.18		4.92		–		2,127		4,788	–
Probable – Stockpile		9,974		2.25		–		–		720		0	–
Total Mineral Reserve		42,022		2.32		3.97		–		3,128		5,367	–
CDMP21 MINERAL RESERVES – OVERALL TOTAL
Proven Mineral Reserve		7,033		2.54		4.06		0.00		574		918	158
Probable Mineral Reserve		42,481		2.03		4.47		0.01		2,774		6,100	7,274
Probable – Stockpile		9,980		2.25		–		–		721		0	0
Total Mineral Reserve		59,494		2.13		3.67		0.01		4,069		7,018	7,432

1. Mineral Reserves are reported based on 31 December 2021 topography surface.

2. The Mineral Reserves were scheduled based on end of August 2021topography surface. Small differences between the Mineral Reserve statement and the production schedule may occur.

4. Mineral Reserve cut-offs are based on $1,350/oz gold price; average oxide recoveries are 61% and average sulfide recoveries are 91%.

5. Cut-off values are shown in Table 12.2. All cut-off values include allowance for royalty payable. There are no credits for silver or copper in the cut-off calculations.

6. There is no Çakmaktepe sulfide Mineral Reserve or Bayramdere Mineral Reserve.

7. Economic analysis has been carried out using a long-term gold price of $1,600/oz.

8. The point of reference for Mineral Reserves is the point of feed into the processing facility.

9. Tonnage is metric tonnes, ounces represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

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Table 12.2 Summary of Cut-off Values, Metallurgical Recoveries, and SSR Ownership of CDMP21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price

Mineral Reserve Classification		SSR Tonnage<br>(kt)		Grades				Cut-off Value/s<br><br>(g/t Au)		Metallurgical Recovery<br><br>(%)	SSR Ownership<br><br>(%)
	Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)
Çöpler Mine – Oxide
Proven Mineral Reserve		–		–		–	–	–		–	–
Probable Mineral Reserve		1,763		1.22		11.17	0.13	0.47–0.59		62.3–78.4	80
Probable – Stockpile		–		–		–	–	–		–	–
Total Mineral Reserve		1,763		1.22		11.17	0.13	0.47–0.59		62.3–78.4	80
Çöpler Mine – Sulfide
Proven Mineral Reserve		326		2.02		6.69	–	1.05		85	80
Probable Mineral Reserve		28,662		2.13		4.96	–
Probable – Stockpile		9,974		2.25		–	–
Total Mineral Reserve		38,962		2.16		3.70	–
Çakmaktepe Mine – Oxide
Proven Mineral Reserve		–		–		–	–	–		–	–
Probable Mineral Reserve		137		1.26		10.91	–	0.52–0.71		14–80	50
Probable – Stockpile		6		2.69		–	–
Total Mineral Reserve		143		1.32		10.49	–
Ardich – Oxide
Proven Mineral Reserve		5,279		1.73		2.00	1.73	0.44–0.80		40–73	77
Probable Mineral Reserve		10,287		1.73		1.91	1.73
Probable – Stockpile		–		–		–	–	–		–	–
Total Mineral Reserve		15,566		1.73		1.94	1.73	0.44–0.80		40–73	77
Ardich – Sulfide
Proven Mineral Reserve		1,428		5.67		11.08	–	1.11		83	78
Probable Mineral Reserve		1,632		3.10		4.15	–		72
Probable – Stockpile		–		–		–	–	–		–	–
Total Mineral Reserve		3,060		4.30		7.38	–	1.11		83	75
CDMP21 Mineral Reserves – Oxide Subtotal
Proven Mineral Reserve		5,279		1.73		2.00	0.00	0.44–0.80		14–80	77
Probable Mineral Reserve		12,187		1.65		3.35	0.03		77
Probable – Stockpile		6		2.83		–	–	0.52–0.71		14–80	50
Total Mineral Reserve		17,472		1.68		2.94	0.02	0.44–0.80		14–80	77
CDMP21 Mineral Reserves – Sulfide Subtotal
Proven Mineral Reserve		1,754		4.99		10.26	–	1.05–1.11		83–85	78
Probable Mineral Reserve		30,294		2.18		4.92	–		79
Probable – Stockpile		9,974		2.25		–	–		80
Total Mineral Reserve		42,022		2.32		3.97	–		79
CDMP21 MINERAL RESERVES – OVERALL TOTAL
Proven Mineral Reserve		7,033		2.54		4.06	0.00	0.44–1.11		14–85	77
Probable Mineral Reserve		42,481		2.03		4.47	0.01		79
Probable – Stockpile		9,980		2.25		–	–		80
Total Mineral Reserve		59,494		2.13		3.67	0.01		78

1. Mineral Reserves are reported based on 31 December 2021 topography surface.

2. The Mineral Reserves were scheduled based on end of August 2021topography surface. Small differences between the Mineral Reserve statement and the production schedule may occur.

4. Mineral Reserve cut-offs are based on $1,350/oz gold price; average oxide recoveries are 61% and average sulfide recoveries are 91%.

5. All cut-off values include allowance for royalty payable. There are no credits for silver or copper in the cut-off calculations.

6. There is no Çakmaktepe sulfide Mineral Reserve or Bayramdere Mineral Reserve.

7. Economic analysis has been carried out using a long-term gold price of $1,600/oz.

8. The point of reference for Mineral Reserves is the point of feed into the processing facility.

9. Tonnage is metric tonnes and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

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Significant factors that could materially affect the Mineral Reserve are:

•Environmental, Permitting, Social, and Community – the Çöpler project is subject to the laws and regulations of Turkey, the mine has several local communities that are nearby. In order to operate the mine, Anagold must maintain appropriate relations with all the authorities and stakeholders. Social, community and government relations are managed by Anagold and include programmes and engagement with the local communities and both local and national governments. Anagold has remained in compliance with all aspects of the Environmental Impact Assessments (EIA) and operating permits throughout the history of the Project.

•Seismic impacts – the Çöpler project is in an area with a history of significant seismic activity that could negatively impact mining operations.

•Metal price impacts – gold is the primary revenue element and silver and copper are produced as by-products. The ore is mined at an elevated cut-off grade and low-grade ore is stockpiled for processing after mining is completed. The use of the elevated cut-off grade serves to mitigate the risks from periods of lower gold prices.

•Mining impacts – the mining equipment is suitable for a selective mining unit (SMU) of approximately 3 m x 3 m x 5 m. This allows for selectivity in mining and enhances the opportunities for blending the feed to the sulfide plant. The total mining rates in the CDMP21TRS mine plan are at 22.5 Mtpa (Çöpler mining only), In the past, total mining rates of 36.5 Mtpa (combination of Çöpler and Çakmaktepe mining) have been achieved, increasing the total mining rate may allow gold to be brought forward in the production schedule but will require additional stockpile storage areas.

•Geotechnical impacts – slope recommendations have significant impacts on the Mineral Reserve and the continued study will allow the Mineral reserves to be maximised.

•Processing impacts – the processing analysis in the Reserve Case includes incorporation of a flotation circuit into the existing sulfide plant to upgrade sulfide sulfur (SS) to fully utilise grinding and pressure oxidation (POX) autoclave capacity. Continued debottlenecking of the sulfide plant and optimisation of the flotation circuit may improve costs and recoveries, changing cut-off grades and impacting the Mineral Reserve.

•The addition of the flotation circuit to the sulfide plant required new grade control protocols and associated stockpile strategies to be implemented to manage the required sulfide plant feed blend. It is likely that there will need to be ongoing modification of the stockpiling cut-offs and procedures for both short-term and longer term blending as the mine progresses. Measures such as increasing the number of active mining areas, increasing the mining rate, and increasing the size or number of run-of-mine (ROM) stockpiles may be required.

12.3 Comparison with Previous Estimates

The 2021 Mineral Reserves inventory is detailed in Table 12.1.

For comparison, a summary of the overall 2020 Mineral Reserve inventory is shown in Table 12.3.

Table 12.3 EOY 2020 Mineral Reserves Summary as at 31 December 2020

Mineral Reserve Classification		EOY 2020 Mineral Reserves Total
	Tonnage (kt)		Au<br>(g/t)		Ag<br>(g/t)		Cu<br>(%)		Gold<br>(koz)		Silver<br>(koz)		Copper<br>(klb)
Proven		1,757		2.31		7.70		0.01		130		435		199
Probable		38,012		2.03		5.76		0.02		2,484		7,040		14,493
Probable Stockpile		5,591		2.54		–		–		457		–		–
Total		45,360		2.11		5.13		0.01		3,071		7,475		14,692

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Mineral Reserves are reported showing only the SSR attributable tonnage portion.

The factors contributing to the differences between the 2021 Mineral Reserves and the previous Mineral Reserves reported as at 31 December 2020 are as follows:

•There was no Mineral Reserve declared for Ardich in 2020. A maiden Mineral Reserve has been declared for Ardich in 2021, and this has added 24Mt at 2.17 g/t Au to the Mineral Reserve, increasing the total by 1.68 Moz of gold.

•New designs for two new phases beneath the Çöpler pit.

•Depletion through mining since 31 December 2020.

There has been an 31% increase in tonnage above the cut-off across all combined Mineral Reserve categories, with a corresponding 32% increase in contained gold.

12.4 Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects

The Mineral Reserves reported in the CDMP21TRS are suitable for reporting as Mineral Reserves using the Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

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13 MINING METHODS

The objective of the CDMP21TRS is to provide a consistent and structured growth plan for the business. Mine plans were updated to improve metal production, through a revised development sequence, available information was then consolidated into a growth strategy, for communication to all levels of the business, using recognised progress reporting systems.

13.1 Geotechnical

13.1.1 Pit Slope Stability – Çöpler

This section contains a summary of the feasibility study level mining geotechnical investigation and design conducted for the Çöpler mine. Much of this work has been prepared prior to 2020, however the work and the recommendations are still applied to the mine designs and workings.

The Çöpler mine maintains an on-site geotechnical monitoring programme that consists of 58 prisms, 33 extensometers, a long-range synthetic aperture radar, and daily data and field monitoring. Additional work is currently in progress to implement pit slope depressurisation. It is expected that pit slope depressurisation will be used extensively throughout the Main pit as the sulfide pit phases are progressed.

In April 2015, Golder Associates (Golder) completed a pit slope optimisation study intended to further optimise the pit slope angles as defined in their earlier study completed in April 2014. This programme included the drilling of five oriented geotechnical core holes to identify any prevalent jointing throughout the Çöpler deposit.

Golder completed the 2015 pit slope optimisation study using recommendations from the 2014 Golder pit slope review with the intention of identifying opportunities to increase definition of potential problem areas within the Çöpler pit. This would allow for mine planning and designs to take advantage of steeper slope angles in some areas. No material changes in-pit slope recommendations were made with the updated report. Anagold chose to continue using the more conservative slope angle recommendations made by Golder in 2014.

The results of the study have provided Anagold with a much better understanding of potential highwall conditions. Not all slope angle recommendations made by Golder were able to be fully followed due to a lack of data and modelling of alteration zones within the Çöpler deposit. Where slope angles were not able to be further refined, Golder recommended that Anagold follow the recommendations set forth in the 2014 geotechnical review.

13.1.2 Review of 2021 Geotechnical Studies

Golder completed three additional Geotechnical Studies / Reports in 2021 which include:

•Golder, Data Review and Geotechnical Model, Çöpler Geotechnical Design Review (October 2021), (PowerPoint Presentation).

•Golder, Çöpler Pit Slope Design Review (November 2021).

•Golder, 2021 Ardich Project Slope Stability Study, Geotechnical support for the Pre Feasibility Study (December 2021).

A geotechnical review of the 2021 Golder reports was completed in December 2021. Key areas of concern and recommendations, where applicable, are provided in Section 13.1.2. The Golder reports focus on the following keys areas:

•Update on structural data

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•Update on rock and soil strengths

•Update on rock mass quality

•Comments on the geological surfaces

•Comments on geotechnical studies commissioned by Anagold

•Geometric review of design

•Review of piezometer data

13.1.2.1 Structural Data

The outcomes of the update on structural data, with focus of areas where faults provided potential control on overall slopes (circled areas in Figure 13.3), are considered appropriate.

13.1.2.2 Rock and Soil Strengths

Whilst the November 2021 report suggests limited ‘new’ laboratory testing of rock strength materials, there was significant testing by Golder in 2020 to justify a review of the 2019 intact rock strengths. Table 13.1 provides an overview and indicates significantly lower intact strengths to those reported by Golder in 2019 (Pit Slope Stability Evaluation, Çöpler Open Pit Mine, November 2019) and with the bolded values in Table 13.1 highlighting the significant changes. Based on Hoek & Brown envelopes and with comparison over a normal stress range of 40–700 kPa the 2021 strengths are somewhat lower than previous Golder studies and with reductions of nominally 35% for Diorite and 12% for Metasediments.

Table 13.1 Golder Intact Strength Estimates

Rock Type		Golder 2019				Golder 20211
	UCS (MPa)		mi		UCS (MPa)		mi
Metasediments		49		22		41		14
Diorite		42		28		22		24
Carbonates		41		10

1 35% percentile values suggested by Golder

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13.1.2.3 Rock Mass Quality

It is understood that Anagold has undertaken mapping of rock mass quality, GSI, within the pit between 2015 and 2017, (Figure 13.1).

Figure 13.1 Rock Mass Quality Mapping Locations (2015–2017)

The mapping indicates GSI values typically below 40. Figure 13.2 provides the data from mapping of the Main pit, predominantly of Diorite and Metasediments and where the majority of mapping took place (225 measurements). Whilst additional mapping of the Marble and Manganese pits also took place, mapping was limited to 42 data points. There is no legend provided for alteration types so it is assumed the alteration noted as OX relates to altered materials and SU relates to presence of sulphides and hence fresh rock.

The data in Figure 13.2 infers median GSI values of nominally:

•19 for altered diorite

•Less than 20 for fresh diorite

•23 for altered metasediments

•38 for fresh metasediments

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Figure 13.2 Rock Mass Quality Mapping of Main Pit

The Anagold mapping indicates significantly lower rock mass quality than what has been adopted in the Golder studies and with the following values noted for rock like materials by Golder:

•41 for diorite

•52 for metasediments

•61 for marble

Golder assign the lower GSIs mapped by Anagold, largely as a result of two factors:

•Observed intact strengths and RQD in the field being lower than typically seen in the core which had been utilised by Golder in assessing rock mass quality of the rock like materials.

•Greater percentage of soil-like material in the exposures and hence significantly poorer GSIs.

The last item is considered as significant and particularly in view that Golder has assumed the percentages of rock-like material comprise 60% of the diorite and 75% of the metasediments. This aspect is further discussed in the following Section.

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13.1.2.4 Geological Surfaces

Golder discusses the Anagold geological models that comprise major lithological units and an RQD block model, the latter utilised by Anagold to provide alteration state. Anagold has utilised comments in the Barr 2012 study (Pit Wall Stability Analysis, Çöpler Mine, August 2012) whereby an RQD of less than 15% implied as signifying altered material, Anagold nominally accepting an RQD of 15 to 25% as being potentially altered and an RQD greater than 25% as unaltered. Figure 13.3 provides the Anagold geotechnical units and which suggest alteration in the final pit occurs in limited areas.

Figure 13.3 Geotechnical Units

Golder comments that the RQD ranges selected by Anagold are too narrow and suggest altered material could have RQD of up to 40% and unaltered material with RQD greater than 60%. These comments by Golder appear appropriate if one considers the location of the rock mass quality mapping discussed above, compare Figure 13.1 and Figure 13.2, which is largely in unaltered materials according to the Anagold model but with the mapping results suggesting largely soil-like materials.

It’s recommended a rock mass quality model be created to allow the slope design recommendations to be appropriately implemented. Such a model needs to take into consideration the Anagold rock mass quality mapping, which suggests there is a higher proportion of soil like materials in the slopes than the Golder estimates.

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13.1.2.5 Geotechnical Reporting Commissioned by Anagold

Golder notes several studies, including both external and internal reports. Two reports relate to internal design checks, which indicate a requirement for localised flattening of designs in areas of altered materials. For the two examples presented, the issues relate to potential impacts on slopes over two benches high and with very high Factors of Safety (FOS) for global stability. Golder also notes “local instabilities may occur where pockets of altered rock mass are exposed on benches. If the altered rock mass area is significant (i.e., exposed over more than two benches) or if the local instabilities cannot be managed by operations, then the mine may need to locally adapt the slope configuration to the altered rock mass slope design or change the whole domain to the shallower altered rock mass design angles”.

Golder notes Anagold had engaged a third party, Professor Tamer Topal, to address the south wall failure in the Marble pit. The scope appears to be limited to a single stability analysis (i.e., one cross-section) based on piezometer data from specific drilled boreholes and utilising all available monitoring (inclusive of inclinometers). However, the scope does not comprise slope recommendations for the failure. Whilst the study has merit, with three large failures and with the Marble pit south wall having failed twice previously it would be prudent that the failures be appropriately backanalysed to confirm that the Geotechnical Model and design assumptions remain appropriate as a key check on inputs for the designs. Without consideration of all failures the results of the Topal study, limited to the current failure, may not provide robust feedback on geotechnical parameters for slope designs going forward.

13.1.2.6 Geometric Review of Design

Golder notes the following regarding the 2021 LOM design “the geometric design for the Çöpler pits conforms with Golder pit slope design recommendations, however, some critical areas have been identified and further design evaluations are recommended”. Following is a summary of the key aspects noted by Golder:

•“Details of the operational pit slope performance is currently unknown to Golder. The pit slope performance would provide valuable information”.

•Requirement for slope stability analyses to address “maximum vertical height for uninterrupted interramp slopes and geotechnical berm widths”.

•“Anagold has recently developed 3D solids for altered and potentially altered rock masses based on the RQD evaluation from the entire database. Golder currently is not completely aware of the details of how these models were developed and how they are representative of the field conditions”.

•“Adequacy of the planned set back distance from designed pit crest to the existing waste rock dump should be evaluated with the slope stability analyses”.

For items 2 and 4 above, Golder does not clarify if these stability analyses would utilise revised geotechnical parameters. It is considered these would be best addressed once the rock mass quality model is created and would include and consider results of backanalyses and slope performance.

Item 3 above requires appropriate interaction between Anagold and Golder such that an appropriate rock mass quality model is developed either “back-boned” to existing Anagold models or appropriately utilising geotechnical borehole logging data and rock mass quality mapping to develop a model. The primary aim, regardless of approach is that the Golder slope design recommendations relate to an Anagold model so that designs can be appropriately implemented.

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13.1.2.7 Piezometer Data

The available seepage locations and piezometer data are provided by Golder but there is key information that is not addressed by Golder and these comprise:

•The seepage is focused at the contact between Diorite and Metasediments, compare Figure 13.3 and Figure 13.4.

•No clear indication of significance of the piezometer data.

The nominal piezometer locations are presented in Figure 13.4 (red dots) and the compiled data in form of a hydrograph in Figure 13.5. The followings trends observed are as follows:

•Significant compartmentalisation in piezometer Pa3.

•Lower groundwater level near limestone contact, piezometer Pa1.

•Groundwater conditions elsewhere indicating phreatic surface at the mined slope and with a Hu of nominally 70% (i.e., 70% of hydrostatic) and indicating depressurisation of the slopes.

Figure 13.4 Locations of Seepage and Ponding

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Figure 13.5 Çöpler VWP Data

13.1.2.8 Review Summary of 2021 Çöpler – Golder Geotechnical Reports

The review of the 2021 Çöpler – Golder Geotechnical Reports has highlighted potential issues of concern. The following are key recommendations from the Geotechnical review.

•Appropriate interaction between Anagold and Golder is required such that an appropriate rock mass quality model is developed either “back-boned” to existing Anagold models or appropriately utilising geotechnical borehole logging data and rock mass quality mapping to develop a model. The primary aim, regardless of approach is that the Golder slope design recommendations relate to an Anagold model so that designs can be appropriately implemented.

•A Feedback loop of appropriate revision of strengths based on backanalysis of failure and review of slope performance.

•Stability analyses once the above components are completed and with appropriate revision of slope design parameters.

13.1.3 RQD Model

RQD is used as a simple and inexpensive indication of rock mass quality. RQD does not account for joint orientation, continuity, or gouge material. Joints sets parallel to the core axis will not intersect the core and therefore is it recommended to use RQD in combination with other geotechnical inputs. RQD is a measure of percent core recovery with artificial fractures ignored.

At the Çöpler project, it has been determined that RQD is a generally reliable indicator of alteration. Therefore, areas with RQD modelled as being less than 15% are considered altered.

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Standard testing of RQD was collected on 661 diamond core holes, 30 of which were drilled within the pit for metallurgical purposes. The 661 holes represent approximately 34% of all drilling in the Çöpler deposit. The Main pit contains RQD measurements for holes evenly spaced with data gaps occurring in the Manganese, Marble, and West pits.

RQD was interpolated in the resource model using the inverse distance method, weighted to the power of two (ID2) with 2 m drillhole composites. A total of six domains were used to estimate RQD values and included a distinction between oxide and sulfide material. To account for the variance in sample spacing, a two-pass approach was used to capture available samples. Model cell estimates were limited to the search distances used with no attempt to assign RQDs to unestimated cells.

13.1.4 Pit Slope Design Parameters

The pit slope design parameters remain unchanged and those applied for each deposit are shown in Table 13.2. Note that for Çakmaktepe, design parameters are in relation to the Central pit and based on the 2018 Golder study which are defined based on azimuth (i.e., direction the slope faces).

Table 13.2 Çöpler Mine Pit Slope Parameters

Çöpler Rock Type		Interramp Slope Angle Çöpler Pits
	Altered RQD<15%		Unaltered (Fresh) RQD>15%
Diorite		23		38
Metasediment		32		43
Marble		50.5		50.5
Gossan Massive Sulfides		40		40
Çakmaktepe Slope Direction		Interramp Slope Angle Çakmaktepe Central Pits
0° to 180° (south-west wall)		34
180° to 360° (all other walls)		40

Golder site review, Çöpler and Golder 2018 for Çakmaktepe

13.1.5 Mine Operations Monitoring and Management

Pit slopes in the Çöpler pit are monitored daily to ensure safety and stability. Daily inspections of the active mining areas are conducted by shift engineers to identify hazards such as unstable rock on benches above, excessive water in and around the highwalls, and any visible cracking and movement of the highwalls. In addition, Anagold employs a geotechnical management team consisting of surveyors, geologists, and geotechnical engineers. This team conducts regular highwall inspections, measurement of movement through extensometers and prism surveys, and data collection and interpretation of the long-range synthetic aperture radar measurements.

Mining at Çöpler utilises perimeter pre-split blasting techniques in areas where competent rock is encountered (typically, limestone/marble, unaltered metasediment, and unaltered diorite). The pre-split holes are drilled according to the bench face angle recommendations as shown in Table 13.2. Blasting is conducted in a manner to minimise back-break through usage of delays and providing adequate relief. A typical pre-split highwall at Çöpler is shown in Figure 13.6.

Where pre-splitting is not practical, highwalls are sloped by excavator to the recommended bench face angle.

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Figure 13.6 Typical 15 m Pre-Split at Çöpler Mine

Anagold, 2020

13.1.6 Geotechnical Domains

Based on the 2014 Golder geotechnical site review, the following geotechnical domain categories are considered appropriate for design recommendations to be founded upon:

•Marble / limestone – characterised by competent rocks and marbleised near the Çöpler intrusion.

•Fresh diorite – characterised as a fresh to slightly weathered or altered moderately strong rock.

•Hydrothermally altered diorite – alteration sufficient to significantly reduce strength relative to fresh diorite, but without the shearing and intense clay alteration of contact and fault zones.

•Weathered diorite and metasediment – highly weathered, extremely weak rock and soil that occurs in the oxidised zone (depth typically to 30 m).

•Fresh metasediment – fresh to slightly weathered, weak to moderately strong rock consisting of a turbidite sequence that may also be structurally complex near faults.

•Hydrothermally altered metasediment – alteration sufficient to significantly reduce strength relative to fresh metasediment, but without the shearing and intense clay alteration of contact and fault zones.

•Fault gouge including intrusive contact and intense sulfide alteration – slicken sided plastic clay with rock fragments that occurs in fault zones including the intrusive contacts.

The character and extent of the hydrothermal alteration beyond the fault zones is poorly defined. Where data are lacking within the alteration zones the most conservative pit slope angle is assumed, representing upside potential should the alteration zone be further defined in the geological model.

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The above listed geotechnical domains are mostly well known and modelled in a geologic model. The alteration zones, however, vary significantly and have not been modelled to an extent to where variations by alteration type are well defined. It has been recommended by Golder that the best way to identify alteration zones is by modelling RQD in the geologic model. For this purpose, RQD values of 15% and less are considered altered and RQD values greater than 15% are considered unaltered, or fresh.

13.1.7 Pit Dewatering

Earlier studies have predicted the formation of pit lakes at various stages of mining. Golder’s hydrogeological study was used to predict pit lake formation. The groundwater flow model predicted that a pit lake would form over time after mining. These results, in conjunction with the acid rock drainage (ARD) work being conducted by SRK Turkey, are being used to predict pit lake water quality.

Sources of groundwater recharge include direct infiltration of precipitation and/or infiltration during storm water run-off events throughout the entire site. Fractured or karstic openings in the bedrock and alluvial sediments along drainages are considered to be the predominant pathways for infiltration. The main hydrogeological units and features considered in the groundwater model were:

•Limestone (modelled hydraulic conductivity = 0.6 m/day)

•Diorite (modelled hydraulic conductivity = 0.0002 m/day)

•Metasediments (modelled hydraulic conductivity = 0.0002 m/day)

•Alluvium (modelled hydraulic conductivity = 10 m/day)

•Various fault systems (Sabirli, Çöpler, and Other) (modelled hydraulic conductivity = 6.1 m/day)

The calibrated groundwater model was used to predict pit inflows and pit lake development based on a pit design with a maximum depth to 875 m. This analysis estimated pit inflow at less than approximately 1,100 m3/day. Estimations of pit lake formation suggest that over a 100-year scenario, based on a pit design with a maximum depth to 875 m, pit lake water elevations are projected to reach the 906 m elevation (±20 m). Modelling results indicate that water from beneath the Lower Çöpler West waste rock dump (WRD) will take more than 1,000 years to flow to the Karasu River. Groundwater located beneath the Lower Çöpler East WRD is estimated to discharge to the Karasu River within approximately 300-years.

Revisions to the pit design since the groundwater model was constructed and calibrated (in 2012) show that the minimum pit elevation (895 mRL) will be higher than the minimum pit elevation simulated in the model (875 mRL). Additionally, the area on the north side of the pit and the southern and south-eastern portions of the pit will be mined to a lower elevation than simulated in the model. Limestone in these areas may increase discharge to the pit during dewatering and may impact the formation of a pit lake following closure. Updating and possibly recalibrating the model based on the revised ultimate pit configuration and available data since 2012 would be required to better quantify the magnitude of the increase or impact.

13.2 Mine Plan

Open pit mining at the Çöpler project is carried out by a mining contractor and managed by Anagold. The mining method is a conventional open pit method with drill and blast and utilising excavators and trucks operating on bench heights of 5 m. The mining contractor provides operators, line supervisors, equipment, and ancillary facilities required for the mining operation. SSR provides management, technical, mine planning, engineering, and grade control functions for the operation.

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SSR currently operates a sulfide process plant and an oxide heap leach facility. Costs are based on the actual operational costs and project budget assumptions. Production schedules and costs are based on current site performance and contracts.

The parameters, costs and throughput assumptions used to prepare cut-off grades and the production schedule are listed in the following sections.

13.2.1 Ore Definition

A revised set of processing parameters was used to calculate the internal Au cut-off grades for ore definition. The cut-off grades for the CDMP21TRS were calculated using the parameters described in the following sections.

13.2.1.1 Oxide Heap Leach Parameters

Table 13.3 details the gold recovery parameters by material type and location.

Table 13.3 Heap Leach – Gold Recoveries

Location		Unit		Material Types
	LMS		M/SED		GOSS		JASP		DIO		MNDIO		OPH
Çöpler Manganese		%		78.4		66.8		71.2		–		71.2		71.2	–
Çöpler Main		%		68.6		66.8		71.2		–		71.2		71.2	–
Çöpler Marble		%		75.7		66.8		65.1		–		62.3		62.3	–
Çakmaktepe Central		%		70.0		80.0		–		73.0		61.0		–	70.0

Table 13.4 details the silver recovery parameters by material type and location.

Table 13.4 Heap Leach – Silver Recoveries

Location		Unit		Material Types
	LMS		M/SED		GOSS		JASP		DIO		MNDIO		OPH
Çöpler Manganese		%		27.3		32.5		27.5		–		37.8		37.8	–
Çöpler Main		%		24.6		32.5		27.5		–		37.8		37.8	–
Çöpler Marble		%		34.0		32.5		27.5		–		32.0		32.0	–
Çakmaktepe Central		%		17.0		28.0		–		17.0		24.0		–	19.0

Table 13.5 details the copper recovery parameters by material type and location.

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Table 13.5 Heap Leach – Copper Recoveries

Location		Unit		Material Types
	LMS		M/SED		GOSS		JASP		DIO		MNDIO		OPH
Çöpler Manganese		%		3.5		13.8		3.3		–		15.8		15.8	–
Çöpler Main		%		3.5		13.8		3.3		–		15.8		15.8	–
Çöpler Marble		%		3.5		13.8		3.3		–		15.8		15.8	–
Çakmaktepe		%		–		–		–		–		–		–	–

Table 13.6 details the operating costs by location.

Table 13.6 Oxide Operating Costs

Parameter	Unit	Çöpler	Çakmaktepe
Rehandle Cost	$/t	0.32	0.64
Processing – Fixed	$/t	3.05	3.05
Processing – Variable	$/t	8.94	8.94
G&A (Process and Site)	$/t	3.17	3.17
Ore Haulage	$/t	–	1.53
Mining Cost	$/t mined	1.89	1.59

13.2.1.2 Sulfide Plant Parameters

The following sections outline the processing parameters for the sulfide plant. Average life-of-mine (LOM) sulfide gold recoveries are 91%.

Throughput

Total Plant Throughput = Direct POX Feed + Float Plant Feed

POX Plant Throughput = Direct POX Feed + Float Plant Concentrate

Table 13.7 details the maximum plant throughputs for each part of the plant. The front-end limit of 400 tph means when the flotation plant is running at full capacity (i.e., 150 tph), the direct feed to the pressure oxidation (POX) circuit will be limited to 250 tph.

Table 13.7 Plant Throughput Limits

Parameter	Unit	Maximum Throughput
Float Plant	t/hr	150
POX Plant	t/hr	280
Total	t/hr	400

Float Plant Throughput = 216345 x Feed SS%2 – 30592 x Feed SS% + 980.24

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Float Concentrate Mass Pull = 277.09 x Feed SS%2 – 15.17 x Feed SS% + 0.33

The POX circuit throughput is also limited by the sulfide sulfur (SS) in the feed to the autoclave, which must be less than 13.75 tph. If the SS content is too high, then the POX circuit throughput will need to be reduced until the rate is less than 13.75 tph SS.

Recovery – POX Gold

POX Gold Recovery = a x (1 - EXP( - b x (Au(g/t) - c ))) + d.

Table 13.8 details the POX gold recovery factors by material type.

Table 13.8 POX – Gold Recovery Parameters

Material Type	a	b	c	d
Limestone / Marble	98.3	1.4	–1.5	–1.00
Metasediment	97.7	1.4	–1.4	–1.00
Gossan	98.3	1.4	–1.5	–1.00
Jasperoid	98.3	1.4	–1.5	–1.00
Diorite	98.3	1.4	–1.5	–1.00
Mn Diorite	96.7	1.2	–1.4	–1.00
Ophiolite	98.3	1.4	–1.5	–1.00

Recovery – Float Plant

Float Concentrate Gold Recovery = 55%.

Float Tails Gold Recovery = 43%.

Float Concentrate SS Recovery = 75%.

Table 13.9 details the operating costs by location.

Table 13.9 Sulfide Operating Costs

Parameter	Unit	Amount
Rehandle Cost	$/t	0.90
Processing – Fixed	$/t	8.32
Processing – Variable	$/t	19.10
Processing – Variable (SS)	$/t SS	2.68
G&A (Process and Site)	$/t	6.60

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13.2.1.3 Metal Prices and Realisation Assumptions

Cut-off grades were determined using a gold price of $1,350/oz. There are no credits for silver or copper in the cut-off grade calculations. Table 13.10 details revenue and realisation assumptions for the Au cut-off grades.

Table 13.10 Au Cut-off Grade Revenue and Realisation Assumptions

Parameter	Unit	Au Cut-off Assumption
Payment and Deductions
Gold	$/oz	1,350
Payable	%	100
Treatment and Refining
Selling	$/oz	8.54
Royalties
Çöpler	%	2
Çakmaktepe	%	4

13.2.2 Ore Cut-off Grades

Internal cut-off grades have been calculated for each of the material types based on the economic inputs and assumptions are shown in Table 13.11. Internal cut-off grades have been used to calculate process quantities within the CDMP21TRS Reserve Case pit stages.

The addition of the flotation circuit to the sulfide plant required new grade control protocols and associated stockpile strategies to be implemented to manage the required sulfide plant feed blend. It is likely that there will need to be ongoing modification of the stockpiling cut-offs and procedures for both short-term and longer term blending as the mine progresses. Measures such as increasing the number of active mining areas, increasing the mining rate, and increasing the size or number of ROM stockpiles may be required.

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Table 13.11 Internal Au Cut-off Grades

Mining Area			Ore Type			Rock Type			Zone		COG<br>(Au g/t)
Çöpler			Oxide			Limestone / Marble			Manganese		0.47
	Main				0.53
	Marble				0.48
	Metasediment				Manganese			0.55
		Main
		Marble
	Gossan				Manganese			0.51
		Main				0.51
		Marble				0.56
	Diorite				Manganese			0.51
		Main				0.51
		Marble				0.59
	Mn Diorite				Manganese			0.51
		Main				0.51
		Marble				0.59
	Sulfide			All			All			1.05
Çakmaktepe			Oxide			Limestone / Breccia			Central		0.60
	Jasperoid				0.57
	Diorite				0.69
	Metasediment				0.52
	Ophiolite				0.60

13.2.3 Pit Design

New pit designs for the CDMP21TRS were created in 2021 based on updated metal prices and costs.

The key aims of the optimised pit designs are:

•Minimise mining costs and maximise economic return by exposing the highest value ore with minimum waste mining.

•Address operational requirements for loading, hauling, slope stability, and rockfall, as follows:

•Loading – the phases were designed with a minimum operational width of 15–30 m between phases (depending on bench configuration) to allow efficient mining for the equipment scale.

•Hauling – generally, two exit haul roads per phase were included: the west bound exit to the crusher, low-grade stockpile, and west dump; and the east bound exit to the potentially acid forming (PAF) and non-acid forming (NAF) dumps. Haul roads are generally 15 m wide at a 10% gradient. Single lane haulage traffic is allowed in the lower benches of the mine and is set at 10 m wide.

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Pit designs for the Çöpler pit were updated in 2021. Ardich pit designs were prepared in 2021 and updated in 2022. The Çöpler, Ardich, and Çakmaktepe pit design for 2034, when in-pit mining is completed for the CDMP21TRS Reserve Case, is shown in Figure 1.4. Following completion of in-pit mining, the sulfide plant will be fed from stockpiles until 2043.

Figure 13.7 Ultimate Pit Designs – CDMP

Anagold, 2022

13.2.4 Waste Dump and Stockpile Design

The CDMP21TRS mine plan allows for the use of five WRDs to store mined waste rock and sulfide ore that is extracted during mining operations. These five WRDs are Lower Çöpler East, Lower Çöpler West, Upper Çöpler, West, and Marble Backfill WRDs. Current operations do not use the Lower Çöpler West and Marble Backfill WRDs. The Lower Çöpler East and Upper Çöpler WRDs will primarily be utilised as sulfide ore stockpile areas, with the Upper Çöpler WRD being mined out to allow for future pushback extension of the Marble pit towards the north and allow for leach pad extensions to the west. Figure 13.8 shows the site layout.

The Lower Çöpler East WRD has a capacity of 14.9 Mm3 (26.8 Mt) of mine waste and 5.5 Mm3 (9.9 Mt) of sulfide ore. The total surface area impacted by the Lower Çöpler East WRD is 51.5 ha. The Lower Çöpler West WRD has a capacity of 94.6 Mm3 (170.3 Mt) of mine waste and 12.4 Mm3 (22.3 Mt) of sulfide ore. The total surface area impacted by the Lower Çöpler West WRD is 206.5 ha. The Upper Çöpler WRD has a capacity of 7.6 Mm3 (13.6 Mt) of sulfide ore. The total surface area impacted by the Upper Çöpler WRD is 26.1 ha. The West WRD complex has a capacity of 34.4 Mm3 (61.9 Mt) of mine waste. The total surface area impacted by the West WRD is 108.9 ha.

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An estimated 69.8 Mt of waste rock will be consumed in the construction of the tailings storage facility, haul road, and tailings pipeline corridor. Total constructed waste rock storage capacity is 155.0 Mm3 (279.1 Mt). The total surface area impacted by all WRDs and stockpiles is 366.9 ha. When possible and economically preferable, waste rock will be backfilled within mined out areas of the pits as they become available.

13.2.4.1 Waste Rock Dump (WRD) Geotechnical Design

The WRDs will generally consist of 15 m tall lifts deposited at the waste material’s angle of repose of approximately 1.33H:1V. The typical bench width will be 17 m and 15 m wide haul roads will be used to construct the WRDs. The WRDs will have overall slopes ranging from 2.5H:1V to 2.6H:1V.

In February 2014, Golder completed an evaluation of the geotechnical stability of the four WRD designs (Golder, 2014a), later updated in May 2015 (Golder, 2015b) to account for the updated material properties developed by Golder during the pit slope optimisation study and the updated waste dump designs and layouts developed by Anagold. Six of the most critical cross-sections were evaluated to determine the minimum Factor of Safety (FOS) for the proposed waste dumps. The sections were aligned to pass through the highest part of the waste piles, the steepest waste pile slopes, and the steepest foundation grades.

In addition to static stability analyses, pseudo-static stability analyses were performed to account for seismic loading conditions for the WRDs. The pseudo-static analyses were conducted based on the procedure proposed by Hynes-Griffin and Franklin (1984) in which a horizontal acceleration equal to 50% of the peak ground acceleration at bedrock is applied to the model. The design criteria peak ground acceleration is 0.30 g for the magnitude 7.0 operating basis earthquake (OBE). Therefore, a horizontal pseudo-static acceleration of 0.15 g was applied to the WRD sections in the seismic stability analyses.

The results of the stability analysis are summarised in Table 13.12.

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Figure 13.8 CDMP21TRS Reserve Case Site Plan

Anagold, 2022

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Table 13.12 Waste Rock Dump (WRD) Design Factor of Safety (FOS)

WRD			Section			Loading Condition			Failure Surface Location		Minimum Computed FOS
Lower Çöpler East			A			Static			Shallow		1.4
	Pseudo-static				1.1
	Static				Deep			1.9
	Pseudo-static					1.3
	B			Static			Shallow			1.7
		Pseudo-static			1.3
		Static			Deep			1.9
		Pseudo-static				1.3
Lower Çöpler West			C			Static			Shallow		1.7
	Pseudo-static				1.3
	Static				Deep			1.9
	Pseudo-static					1.3
	D			Static			Shallow			1.6
		Pseudo-static			1.2
		Static			Deep			1.8
		Pseudo-static				1.3
West Çöpler			E			Static			Shallow		1.6
	Pseudo-static				1.1
	Static				Deep			1.9
	Pseudo-static					1.3
	F			Static			Shallow			1.6
		Pseudo-static			1.2
		Static			Deep			2.0
		Pseudo-static				1.4

The Lower Çöpler East WRD facility will be constructed over a portion of the existing North-east WRD. Foundation conditions underlying the existing North-east WRD and the proposed Lower Çöpler East facility consist of Munzur Limestone. Minimum computed factors of safety for the Lower Çöpler East facility are 1.4 and 1.1 for static and seismic loading conditions, respectively.

The Lower Çöpler West WRD facility will be founded on Munzur limestone. Limit equilibrium stability analyses indicate minimum computed FOS of 1.6 and 1.2 for static and seismic loading conditions, respectively (Golder, 2015b).

The West WRD is to be constructed adjacent to the Çöpler open pit and will be founded on Munzur Formation limestone and metasediment with sporadic diorite intrusions. Minimum computed FOS are 1.9 and 1.3 for static and seismic loading conditions, respectively.

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13.2.4.2 Waste Rock Geochemical Review

Anagold mines and monitors the waste rock types to determine PAF and NAF material according to the Çöpler waste rock management plan to ensure proper disposal of PAF material as it is encountered during the ore control process. SRK established the criteria for identifying PAF and NAF material as shown in Table 13.13.

Table 13.13 Waste Rock Geochemical Classification

Lithology		Sulfide Sulfur (SS%) Cut-off Grade		Waste Rock Groups	Descriptions
Diorite		0.8		PAF/High-sulfide diorite	Diorite with SS ≥0.8%
	NAF/Low-sulfide diorite		Diorite SS <0.8%
Metasediment		0.8		PAF/High-sulfide MTS	Metasediment with SS ≥0.8%
	NAF/Low-sulfide MTS		Metasediment with SS <0.8%
Limestone / Marble		2		High-sulfide LMS	Limestone with SS ≥2%.
	Low-sulfide LMS		Limestone with SS <2%.
Gossan		–		Gossan – NAF	All Gossan unit
MnOx		–		MnOx – NAF	All MnOx unit
Massive Pyrite		–		Massive Pyrite – PAF	All Massive Pyrite unit

In September 2015, SRK completed a geochemical impact assessment for the Çöpler WRD facilities. The key findings from the SRK report suggests that all WRD facilities at Çöpler, except one, have a neutralising potential (NP) to acid potential (AP) ratio of greater than 20:1; indicating that the Çöpler material has excellent neutralisation capacity for ARD. The one exception to this was the West WRD, which was estimated to have a NP:AP ratio 1:3. It was recommended that Anagold optimise the WRD construction sequencing in order to take advantage of the neutralisation potential of the other WRD facilities by blending higher quantities of NAF material into the West WRD. Anagold anticipates that this will be a readily achievable solution that will not add any additional costs to the Project.

A series of waste rock samples representing the LOM distribution were tested by SRK in order to measure the immediate reactivity, future acid potential, and long-term acid potential of the waste rock.

Regarding immediate reactivity, a paste pH test was conducted that resulted in all samples generating near-neutral and slightly alkaline paste pH.

Regarding future acid potential, a large majority of all samples taken reside above the NP:AP 1:1 boundary. The remainder of the samples that fall below the 1:1 boundary are extremely close to the 1:1 boundary and should only pose a minimal risk to ARD generation. In terms of long-term acid potential, only two samples registered below the 1:1 NP:AP ratio.

13.2.5 Ore Stockpiles, Rehandle and Blending

For the CDMP21TRS, oxide and sulfide ore are processed through separate crushing circuits.

Oxide ore that is unable to be directly dumped into the crushing circuit is placed on the appropriate stockpile for processing at a later time. Oxide ore is typically segregated according to clay content and grade. The processing engineer determines the desired blend on a daily basis in order to maintain a consistent feed grade and rock type blend going to the heap leach pad.

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All sulfide ore is currently placed in one of three primary stockpiles: High-grade, medium-grade, and low-grade. Sulfide ore is directed to the primary stockpiles or to the crusher pad. There is no allowance for material to be directly dumped into the sulfide crushing circuit. All material is rehandled by a loader from the crushing pad into the crushing circuit.

The following Au grade bin assumptions were used for the CDMP21TRS Mineral Reserves:

•High-grade Au >4.0 g/t Au

•Medium-grade Au 2.5–4.0 g/t Au

•Low-grade Au 1.05–2.5 g/t Au

Currently site typically experience a lack of sulfide sulfur (SS) feed to the POX, requiring additional cost to run the POX plant. The flotation plant was designed to upgrade (increase) the SS feed into the POX circuit. For the POX autoclave to operate autogenously, SS feed must be above 10.20 tph and less than 13.75 tph to achieve target oxidation with current oxygen availability. If the SS feed rate is too high, then the feed to the plant will need to be reduced until the POX SS feed rate is less than 13.75 tph limit. Operating performance of the autoclaves indicates that higher than design oxygen utilisations efficiencies are possible, which may allow greater than 13.75 tph sulfide sulfur to be treated. This oxygen utilisation efficiency along with increased oxygen availability is upside to the CDMP21TRS Reserve Case.

Plant feed will therefore need to be blended to achieve the target SS feed range of 10.20–13.75 tph into POX.

To blend on SS feed, new grade control protocols have been developed and implemented on site. Site grade control is currently being done on Au and SS grades to aid in achieving the ideal range for SS feed into the plant and assist with the development of a new stockpile strategy.

The following SS grade bin assumptions were used for the Mineral Reserves inside each Au grade bin:

•High-grade SS >4.8% SS

•Medium-grade SS 3.2% to 4.8% SS

•Low-grade SS <3.2% SS

The effectiveness of these new grade bins in controlling the SS blend will need to be monitored on an ongoing basis as the plant matures and adjustments to the grade bin parameters (and size of stockpiles) may be required. This work will need to continue as the mine progresses and new mining areas are included.

The smallest parcel size for plant feed considered for the CDMP21TRS Mineral Reserves was one month.

The operation will need to be in control of the plant feed blend at a more granular level than was modelled for the Mineral Reserves. If maintaining a plant feed blend at a more granular level was found to be problematic, there are several measures that site could implement to manage any short-term and longer term concerns:

•Mine working areas

Given the relatively small size of the mining fleet, the number of active mining working areas could be increased, increasing mining selectivity, and therefore improving the blending capacity from the mine.

•Stockpile size

The size of stockpiles could be adjusted to reduce feed impacts from short-term fluctuations coming from the mine.

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•Mining rate

Given the current site contract mining arrangement, site could ramp up the mining rates to reach sufficient material (of the required type) to maintain the required blend.

•Variation of grade bins

Grade bin designations could be adjusted to have better control of the grade bands that are causing problems in the plant feed blend.

13.2.6 Grade Control

All ore control operations are managed by Anagold technical staff. Anagold maintains an on-site laboratory with the capacity to assay an average of 600 blasthole samples per day.

Prior to sampling, blastholes are identified as ‘potential ore’ (oxide or sulfide) or ‘potential waste’ (oxide or sulfide) based on grade control data from the bench above and the mining model prediction. A 10 m outside buffer is then applied to the potential ore areas to ensure appropriate sampling density. All potential ore blastholes are sampled for AuFA (fire assay for Au). Approximately 50% of potential ore blastholes are sampled for AuCN (cyanide soluble Au), total carbon, and total sulfur. Additionally, all potential sulfide ore blastholes are sampled for SS. Approximately 25% of potential waste blastholes are sampled for AuFA, AuCN, total carbon, and total sulfur.

Sampling of the blasthole drill cuttings is performed according to the formal procedure by using a sample scoop to extract a complete cross-section of the cutting pile. The sampled cuttings are deposited into a canvas bag, which is labelled with the drillhole identifier (ID) and with a laboratory information management system (LIMS) bar code tag inserted into the bag with the cuttings. Sample bags are sealed and sent to the on-site laboratory for analysis. The sample scoop is cleaned prior to collecting each sample to avoid contamination between samples.

Assay results are uploaded to the ore control database with reference to each specific drillhole ID. The assay results are then estimated into a cell model with parent cell sizes of 3 m x 3 m x 5 m using ordinary kriging (OK) to estimate ore grade and type. The ore control geologist will then digitise mining shapes with a minimum width of 3 m (to match the SMU) and minimum tonnage of 500 t. These mining shapes are then sent to the survey group for layout in the mine using colour coded flagging under the supervision of the ore control geologist.

To effectively blend the sulfide feed on SS content, new grade control protocols were developed and implemented on site in 2021. They are undergoing further review to optimise and improve production.

13.3 Mine Production Schedule

The CDMP21TRS Reserve Case has examined production from three open pit mining locations at the Çöpler mine, the Çöpler deposit, the Ardich deposit and the Çakmaktepe deposit. The Çakmaktepe pit, which contains only oxide ore, is almost exhausted. Anagold has prepared the open pit production schedules. The case adopted for the CDMP21TRS Reserve Case is based on Mineral Reserves only and does not include Inferred Mineral Resources. Figure 13.9 shows total mine production and the tonnages and grades for each ore type on a 100% project basis.

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Figure 13.9 CDMP21TRS Reserve Case Mining Production

OreWin, 2022

13.3.1 Scheduling Assumptions

The following scheduling methodology was used for the CDMP21TRS to balance mine, mill, and stockpile quantities:

•Heap leach:

•Oxide ore is not limited by processing capacity.

•Oxide ore that is unable to be directly dumped into the oxide crushing circuit is placed in the appropriate stockpile for future processing.

•Oxide ore is segregated dependent on clay content and average grade.

•Sulfide plant:

•All sulfide ore is segregated into one of three primary gold stockpiles: high-grade, medium-grade, and low-grade, which are each further split by SS grade.

•Existing stockpiles are mined at the average grade of each stockpile.

•All material is rehandled by a loader from the crushing pad into the crushing circuit (no direct tipping).

•The flotation circuit was commissioned in December 2021 with circuit ramp up and a transition to stable operations expected in early-2022.

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•Plant throughput capacity is calculated from the available mill hours and varies by material type.

•The production schedules are based on Proven and Probable Mineral Reserves only. No Inferred Mineral Resources were used.

•The open pit schedules were based on mining inventories by bench reported within the pit stages.

•Low-grade stockpiling was used to balance the mining rate where necessary.

13.3.2 Production Schedule

The input assumptions for CDMP21TRS Reserve Case were adjusted based on current mine and production performances including throughput rates and recoveries.

All throughput rates are reported inclusive of all availability and utilisation factors on a calendar year. Total mine production is limited to an annual average of 22.5 Mtpa. The throughput assumptions are supported by current mining rates including productivity allowances for winter and summer conditions. Mining rates are limited based on vertical advance and bench configuration in order to ensure that the schedule is achievable. Production is not limited by the mining rate and increases in rate would be possible to bring forward oxide ore or increase stockpiling to bring higher grade feed to the sulfide plant.

Mining in the CDMP21TRS Reserve Case is completed in 2034, after which the sulfide plant is fed from stockpiles.

The objective of the production schedule is to maximise the early cash flow by delaying costs and bringing revenue forward with ore feed to meet concentrator throughput capacity. Considerations for the LOM scheduling include:

•Ensuring continuous ore supply to the concentrator by delivering the highest value ore first and meeting physical mining and milling hours capacity constraints.

•Achieving excavator productivities and sinking rates to deliver ore at maximum utilisation of milling hours available at the concentrator.

•Maximising annual utilisation hours for the mine loading equipment.

•Maintaining a balance of ore throughput rates (material types) and mill cut-off grades that allows milling hours to be maximised.

The mine schedule incorporates strategic stockpiling considerations by optimising the number of excavators on the benches of early phases, increasing the opportunity to raise mill cut-off grades. This leads to stockpiling medium-grade and low-grade material and sending higher grade ore to the mill sooner. The open pit total movement is shown in Table 13.14 on a 100% project basis.

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Table 13.14 CDMP21TRS Reserve Case Mining Schedule

Mining Year		Total Tonnes <br>(kt)		Oxide Ore								Sulfide Ore						Waste<br>Tonnes<br>(kt)
	Tonnes<br>(kt)		Au <br>(g/t)		Ag<br>(g/t)		Cu<br>(%)		Tonnes<br>(kt)		Au<br>(g/t)		Ag<br>(g/t)		SS<br>(%)
2022		28,065		353		0.95		5.74		0.11		3,884		2.41		4.27	–	23,828
2023		36,516		1,920		1.22		9.43		0.06		2,582		2.32		8.73	–	32,014
2024		52,435		3,980		1.18		2.04		0.00		3,814		2.39		2.33	–	44,641
2025		52,375		2,196		1.72		2.04		0.00		6,817		2.13		5.32	–	43,362
2026		52,331		2,880		1.66		2.05		0.01		4,607		1.81		3.93	–	44,843
2027		52,340		1,266		1.39		2.04		0.01		5,745		2.09		4.47	–	45,329
2028		52,389		3,136		1.97		2.72		0.01		4,614		2.16		4.70	–	44,639
2029		46,088		1,230		1.99		5.57		0.04		5,988		2.48		6.37	–	38,870
2030		25,000		992		2.27		2.04		–		392		4.23		4.23	–	23,616
2031		25,000		30		1.23		2.04		–		–		–		–	–	24,970
2032		25,000		918		2.03		2.04		–		–		–		–	–	24,082
2033		20,000		2,504		1.95		2.04		–		338		3.16		4.23	–	17,158
2034		4,496		1,083		2.37		2.04		–		1,384		5.22		4.23	–	2,029
Total		472,035		22,486		1.69		3.02		0.01		40,167		2.34		4.90	–	409,382

Table shows mining schedule does not show processing or existing stockpile rehandle

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13.3.3 Processing Schedule

The CDMP21TRS processing schedule was balanced to meet the maximum build rates for the oxide heap leach pads, or available mill hours for the sulfide plant.

Sulfide ore production throughputs are limited dependent on ore tonnage, SS tonnage, and carbonate content, (expressed as C). The sulfide plant crusher / grinding circuit is limited to 400 tph, while the limitations on SS tonnage exist due to the consumption of oxygen by SS in the POX circuit and carbonate content to maintain an operable acid balance through the acidulation and POX circuits. The process facilities are limited by the amount of oxygen that can be provided to the POX process. Based on current performance, high-SS is unlikely to be a problem, and any higher material would be blended down using low-SS material. The carbonate:SS ratio will potentially be an issue with declining SS grades. The main issue currently appears to be a lack of SS in the feed, forming the justification for the flotation circuit. The flotation circuit upgrades the SS content into the autoclave feed and rejects carbonate.

In order to target the highest value material, the sulfide production schedule is therefore required to target the highest value material, while also balancing the plant throughput rates and required range of sulfide sulfur into the autoclave.

The CDMP21TRS Reserve Case production is predominantly from sulfide ore. The oxide heap leach and sulfide plant processing schedules feed type, Au grade, and gold production are shown in Figure 13.10. Gold production and recovery is shown in Figure 13.11. The annual production schedule is in Table 13.15. Schedules are on a 100% project basis.

The CDMP21TRS Reserve Case production includes 22.5 Mt at 1.69 g/t Au oxide ore processed by heap leaching and 52.9 Mt at 2.33 g/t Au processed in the sulfide plant. Total ore production is 75.4 Mt at 2.14 g/t Au. Total gold production is 4.4 Moz. Mining at the Çöpler pit is completed in 2029 and at Ardich in 2034. Oxide heap leach stacking is completed in in 2034, while sulfide processing will continue from stockpiles until 2042. Sulfide processing will continue from stockpiles until 2042 for a 21-year mine life. The production schedule is for the period 1 January 2022 through 2042.

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Figure 13.10 CDMP21TRS Reserve Case Processing Schedule

OreWin, 2022

Figure 13.11 CDMP21TRS Reserve Case Gold Production and Recovery

OreWin, 2022

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Table 13.15 CDMP21TRS Reserve Case Production Schedule

Description		Units		Total	Year
	2022		2023		2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	2035	2036	2037	2038	2039	2040	2041	2042
Heap Leach Stacked		kt		22,557	263	2,080	3,980	2,196	2,880	1,266	3,136	1,230	992	30	918	2,504	1,083	–	–	–	–	–	–	–	–
Au Feed Grade		g/t		1.69	0.88	1.22	1.18	1.72	1.66	1.39	1.97	1.99	2.27	1.23	2.03	1.95	2.37	–	–	–	–	–	–	–	–
Ag Feed Grade		g/t		3.04	4.90	9.41	2.04	2.04	2.05	2.04	2.72	5.57	2.04	2.04	2.04	2.04	2.04	–	–	–	–	–	–	–	–
Cu Feed Grade		%		0.01	0.10	0.07	0.00	0.00	0.01	0.01	0.01	0.04	–	–	–	–	–	–	–	–	–	–	–	–	–
Sulfide Plant Feed		kt		52,892	2,708	2,395	2,635	2,551	2,569	2,442	2,378	2,490	2,504	2,269	2,538	2,718	2,101	2,607	2,730	2,739	2,305	2,623	2,759	2,476	2,356
Au Feed Grade		g/t		2.33	3.16	2.69	2.87	2.97	2.25	2.77	2.75	3.58	2.18	1.70	1.61	1.91	3.99	1.80	2.10	2.14	1.86	1.96	1.80	1.49	1.69
Ag Feed Grade		g/t		3.89	3.82	4.71	1.99	3.96	3.67	4.72	4.48	7.24	4.89	4.80	5.39	4.11	4.34	2.14	1.33	0.68	3.08	5.99	4.13	4.49	2.34
Total Feed		kt		75,448	2,970	4,475	6,614	4,747	5,449	3,709	5,514	3,720	3,495	2,299	3,456	5,222	3,184	2,607	2,730	2,739	2,305	2,623	2,759	2,476	2,356
Total Metal Recovered
Au Recovered		koz		4,369	268	238	318	302	264	245	310	314	203	118	152	228	301	139	164	167	126	148	143	107	115
Ag Recovered		koz		663	21	156	55	38	43	28	66	74	26	12	23	38	22	5	3	2	7	15	11	11	5
Cu Recovered		klb		161	24	87	12	3	17	1	5	12	1	0	–	–	–	–	–	–	–	–	–	–	–

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14 PROCESSING AND RECOVERY METHODS

The following section describes the Projects existing processing operations.

14.1 Sulfide Ore Processing

The sulfide plant commenced commissioning in Q4’18. The basic flow sheet is shown in Figure 14.1 and comprises:

•Crushing and ore handling

•Grinding

•Acidulation

•Pressure oxidation

•Iron / arsenic precipitation

•Counter current decantation (CCD)

•Gold leach, carbon adsorption, and detoxification

•Carbon desorption and refining

•Neutralisation and tailings

•Tailings storage facility (TSF)

Figure 14.1 Çöpler Process Flow Sheet for Sulfide Plant

Anagold, 2020

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The incorporation of a flotation circuit into the existing sulfide plant to upgrade sulfide sulfur (SS) to fully utilise grinding and pressure oxidation (POX) autoclave capacity has completed construction and commenced commissioning on ore in January 2022. This addition to the sulfide plant is incorporated between grinding and acidulation, as shown in Figure 14.2, by taking a bleed / slip stream from the grinding thickener feed, floating sulfides, and returning the sulfide concentrate to the grinding thickener to be combined with direct feed. Gold not recovered to flotation concentrate will report with flotation tails to the gold leaching and recovery circuit and combined with material process through the POX autoclave circuit to recover gold.

The flotation circuit will also reject carbonates to flotation tails, bypassing acidulation and POX, providing additional benefits in the acid balance through POX.

Figure 14.2 Flotation Block Flow Diagram

Anagold, 2020

The existing sulfide circuit, before the addition of flotation, has demonstrated additional latent capacity in throughput controlling sections of the circuit, crushing/grinding and autoclaves. The incorporation of flotation will allow the POX autoclaves to maximise throughput and sulfide sulfur oxidation capacity, utilising latent capacity in the process plant, in particular the grinding and pressure oxidation circuits. Fully utilising this latent capacity with the addition of a small flotation plant allows with minimal capital cost the increase in overall plant throughput.

The throughput from crushing and grinding was designed with a nominal volumetric capacity of 306 tph will increase up to a maximum of 400 tph. Additionally, the POX autoclave circuit has demonstrated it can process up to a long-term average maximum of 280 tph feed (two autoclave operation) and 13.75 tph sulfide sulfur, compared to design of 245 tph and 12.5 tph respectively. The limit of 13.75 tph sulfide sulfur is dictated by the capacity of the oxygen supply to effect oxidation of the sulfides, design 96%.

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The flotation plant feed rate is variable between 50–150 tph based on sulfide sulfur feed grade and the oxidation capacity of the POX autoclaves to oxidise sulfides. Operating performance of the autoclaves indicates that higher than design oxygen utilisations efficiencies are possible, which may allow greater than 13.75 tph sulfide sulfur to be treated. Alternatively, increased autoclave throughput with reduced sulfide oxidation is possible, with a resultant reduction in overall gold recovery. This oxygen utilisation efficiency, along with increased oxygen availability, is upside to the CDMP21TRS Reserve Case.

14.1.1 Sulfide Plant Performance

The sulfide plant commenced commissioning in Q4’18.

The operating performance is summarised in Figure 14.3 for throughput and recovery against the design for the period 2020 and 2021.

Since completing ramp-up of the sulfide plant in June 2020, POX throughput has progressively improved to exceed design up to a monthly average peak of 330 tph and at the maximum SS of 13.7 tph. The gold recovery has remained at around 91%, lower than design, with the tailings grade remaining stable between 0.25–0.30 g/t Au.

Further improvements have been implemented during 2020–2021. The includes the installation of oxygen to leach to supplement air to maintain sufficient oxygen levels for gold leaching has led to improved recoveries.

Figure 14.3 Gold Recovery and Throughput Comparison

Anagold, 2020

14.1.2 Sulfide Plant Description

A detailed sulfide flow sheet is shown in Figure 14.4. The following description of the sulfide plant includes the existing operating circuits and the flotation circuit.

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14.1.2.1 Crushing and Ore Handling

Haul trucks from the mine tip ore onto designated stockpile fingers. The ore is withdrawn from stockpiles by front end loader (FEL) and deposited into the run-of-mine (ROM) dump hopper. A static grizzly is fitted to the top of the ROM bin to remove coarse oversize.

ROM ore is reclaimed from the bin by the sizer apron feeder, which discharges material into the mineral sizer. The sizer is a tooth roll unit which crushes the ore from a feed top size of 500 mm to a nominal top size of 250 mm. Discharge from the sizer drops down a chute onto the sizer discharge conveyor.

The sizer teeth are configured in a manner to direct oversize rocks to one end where they pass through a spring-loaded oversize rejection gate and fall to a reject bunker. The crushed product is carried by the sizer product conveyor to the semi-autogenous grind (SAG) mill feed conveyor. The SAG mill feed conveyor has a belt scale to monitor the ore flow to the SAG mill and this information is used to control the sizer apron feeder speed.

14.1.2.2 Grinding

The SAG milling stage consists of a high aspect SAG mill with water cannon pebble recycle. The SAG mill grinds the crushed ore to produce a discharge particle size distribution P80 of approximately 1,400 µm.

Large ore particles are retained in the SAG mill by the internal SAG discharge grate. Particles too large for ball milling are retained as oversize on the SAG mill trommel screen and this oversize is washed by trommel sprays. The trommel screen oversize is either projected back into the SAG mill using a high-pressure water cannon or rejected via a conveyor. Slurry that passes through the trommel screen discharges into the grinding cyclone feed pump box where it mixes with the ball mill discharge slurry and density control water.

Slurry collected in the grinding cyclone feed pump box from the SAG mill and ball mill is fed to the grinding cyclone cluster. The cyclones produce an overflow product with a P80 of 100 µm, which is screened to remove any trash (organic material, etc.) by the grinding trash screen. Coarse particles report to cyclone underflow, which is returned to the ball mill for further size reduction until it is fine enough to report to cyclone overflow and leave the circuit.

The slurry product from the grinding circuit, trash screen undersize, is currently thickened in a high rate thickener and excess water reports to the thickener overflow for immediate re-use within the grinding circuit. The thickened slurry discharging from the thickener underflow is pumped to the grinding thickener underflow storage tanks.

To provide for the flotation circuit, a portion of the trash screen undersize, dependent on POX autoclave sulfide sulfur requirements, will be diverted to the flotation circuit where the remaining slurry continues to the thickener.

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Figure 14.4 Process Flow Sheet for Sulfide Plant

Anagold, 2020

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14.1.2.3 Flotation

A portion of the grinding trash screen undersize will be diverted to the flotation circuit and pumped to the conditioning tanks. This proportion, between 50 tph and 150 tph, will depend on SS feed grade and POX autoclave SS requirements. The flotation circuit can operate as a single or dual train, each train will have a maximum throughput of 75 tph.

The flotation plant consists of two equally sized conditioning tanks, in series, for copper sulfate, if required, and potassium amyl xanthate (PAX) conditioning with a nominal residence time of seven minutes each tank. From conditioning, the slurry is pumped to two equally sized flotation trains consisting of six 50 m3 tank cells with a residence time of nominally 60 minutes at maximum throughput (75 t/h each). Frother dosing and supplemental collector dosing will occur down the trains in every second cell. The plant is designed to handle high mass pull to maximise sulfide recovery, with preference to high recovery over high selectivity.

The flotation concentrate is pumped to the grinding thickener feed mixing with slurry directly from the grinding circuit upgrading the sulfide sulfur material fed to the acidulation and POX circuit. The flotation tail is pumped to the gold leach tanks for recovery of gold present in the non-sulfidic portions of the ore.

14.1.2.4 Acidulation

The grinding thickener underflow storage tanks provide process surge and effectively decouple the upstream crushing, grinding and flotation, when operating, circuits from the downstream hydrometallurgical circuit. If the acidulation feed tanks reach their high-level limit then ore feed to the upstream circuits will be stopped. If the tanks are approaching their low-level limit then the upstream circuit feed rate can be increased to compensate.

The tanks are agitated for solids suspension and mixing and have a total residence time of 12 hours. Agitation achieves short term blending of the incoming feed from the upstream circuits, and this provides a relatively slow changing feed composition to the downstream hydrometallurgical circuit. Antiscalant can be added to these tanks if necessary, to reduce scale build-up in the downstream acidulation circuit.

The acidulation circuit uses recycled solution, containing free acid, from the decant thickener to leach the carbonate minerals in the ore. Supplemental concentrated sulfuric acid can also be added, when required, to meet total acid addition demand. The total acid addition targets nearly complete destruction of acid soluble carbonates in the acidulation tanks. Acidulation is conducted in two reaction tanks. The acidulation tanks are agitated to disperse the slurry, acid and decant thickener overflow recycle throughout the tank and ensure the carbonates in the ore react with the acid in solution.

Depending on the ore type being processed the slurry from the grinding thickener underflow storage tanks is split between acidulation and the POX feed tanks. The proportion of this split is determined by how much carbonate in the feed material requires destruction to achieve the target of 22.5 g/L free acid content in the POX autoclave discharge slurry. This free acid level favours the formation of an iron mineral reaction product which exhibits better settling behaviour in downstream thickeners (hematite favoured over jarosite), while also reducing the potential for excessive CO2 gas evolution and gypsum scaling in the POX autoclaves.

Additional concentrated sulfuric acid is added if required to maintain the targeted acid soluble carbonate destruction in the acidulation tanks. When there are low carbonate levels in the feed, and little or no acidulation is required, POX feed thickener overflow solution is recycled to the acidulation tanks (instead of decant thickener overflow solution) to limit the maximum concentration in the tanks to 30% solids.

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Slurry overflows from acidulation tank 1 into acidulation tank 2 and then discharges into the POX feed thickener mix tank. Either of the acidulation tanks can be bypassed, if required. The diluted slurry from acidulation requires thickening prior to storage in the POX feed tanks. The POX feed thickener recovers excess solution and advances it to the decant thickener (as wash water) and/or to the iron / arsenic precipitation circuit (to maintain the water balance in the acidulation circuit) or recycles it to acidulation tank 1.

POX thickener underflow slurry is pumped to the POX feed thickener underflow surge tank. The storage in the surge tank allows blending in the correct proportions of the acidulated slurry with the un-acidulated grinding thickener underflow slurry in the POX feed tank to ensure the total level of acid soluble carbonates in the POX feed slurry is within target levels.

The decant thickener recovers acid (that is generated in the POX autoclaves) from the POX discharge slurry and recycles it to the acidulation circuit for carbonate destruction. The underflow slurry is pumped from the thickener to the iron / arsenic precipitation circuit by the decant thickener underflow pumps. Thickener overflow gravitates to the decant thickener overflow tank from where it is pumped to the acidulation tanks by the decant thickener overflow pumps. Solution is bypassed to the POX feed thickener overflow tank when processing low-carbonate ores.

14.1.2.5 Pressure Oxidation

The POX feed surge tanks 1 and 2 are a common feed system that services both POX autoclave trains (T1 and T2). The tanks are agitated to mix / blend the incoming slurry and provide approximately 18 hours of slurry storage to minimise disruptions to the POX circuit. For simplicity, where only POX T1 is discussed in this document it is assumed that both T1 and T2 have identical configurations and controls.

Slurry is pumped to the POX trains 1 and 2 low-temperature heaters by the POX heating feed pumps. The low-temperature (LT) heater receives incoming feed slurry and vent gas (predominantly steam) recovered from the LT flash vessel. The gas heats the slurry to approximately 95ºC before being transferred to the high-temperature (HT) heater. The steam in the gas condenses and any excess is vented to the wetted elbow of the POX T1 Venturi scrubber.

The HT heater receives slurry from the LT heater and vent gas (predominantly steam) recovered from the HT flash vessel. The gas heats the slurry to approximately 150ºC before being pumped to the POX autoclave. The steam in the gas condenses and any non-condensing gases accumulate in the vapor space at the top of the vessel, prior to being vented.

Slurry is pumped to the autoclave by two pumping trains.

If one full autoclave train is offline, the remaining autoclave train can operate at 150% of normal capacity, provided both of its feed pumping trains are operating.

A horizontal multi-compartment autoclave is used to oxidise the sulfides in the ore at high temperature and pressure using gaseous oxygen. The oxidation of sulfide material in the autoclave generates heat and when the rate of heat generation exceeds that required to achieve the target temperature of 220°C quench water is added. Sufficient quench water is added to control the temperature to the target. The quench water is pumped through the same sparge pipe that introduces gaseous oxygen addition into the autoclave. There is one sparge pipe underneath each autoclave agitator.

A vent controls the pressure in the autoclave to prevent the water boiling. This pressure is called overpressure and results from the presence of gases such as oxygen, nitrogen, and CO2.

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Slurry discharges from the autoclave through a severe service let down valve to the HT flash vessel. The HT flash vessel operates at a lower pressure than the autoclave and the resulting pressure drop for the discharge slurry entering the HT flash results in steam being flashed from the slurry. The flashing of steam cools the slurry to the equilibrium temperature corresponding to the pressure in the flash vessel.

Steam vented from the HT flash is sent to the HT heater to heat the feed to the autoclave, excess steam is vented to the venturi scrubber for treatment prior to discharge.

Slurry discharges from the HT flash vessel through a severe service let down valve to the LT flash vessel. The LT flash vessel operates at a lower pressure than the HT flash vessel, the resulting pressure drop for the discharge slurry entering the LT flash results in steam being flashed from the slurry. The flashing of steam cools the slurry to approximately 100°C at a pressure just above atmospheric. Slurry is forced from the HT flash vessel to the LT flash vessel by the pressure difference between the two vessels.

Steam vented from the LT flash is sent to the LT heater to heat the feed to the HT heater, excess steam is vented from the LT heater to the Venturi scrubber for treatment prior to discharge.

Steam, entrained slurry, together with gas, including carbon dioxide and unreacted oxygen vented from various points in the autoclave circuit, is scrubbed in Venturi scrubber to remove entrained acidic slurry droplets.

Demineralised water is used in the POX circuit for steam production and for seal water.

Flashed slurry is pumped from the LT flash vessel by decant thickener feed. The decant thickener was described previously and the decant thickener underflow is feed to iron / arsenic precipitation.

14.1.2.6 Fe/As Precipitation

Iron / arsenic precipitation uses limestone slurry addition to the decant thickener underflow slurry to neutralise the free acid and raise the pH to approximately 2.8, which removes ferric iron and arsenic from solution.

The decant thickener underflow duty pump transfers the thickener underflow slurry to iron / arsenic precipitation tank 1. Limestone is added for pH control, and low-pressure air is sparged into the tanks to oxidise any ferrous iron that may be present to ferric iron. The ferric ions combine with the residual arsenic, also leached in the POX circuit, and precipitate together as the pH of the solution is raised. Limestone reacting with the free acid generates carbon dioxide gas and gypsum.

The two iron / arsenic precipitation tanks normally operate in series. The treated slurry overflows from the second iron / arsenic precipitation tank to the CCD 1 Mix Tank.

The low-pressure air and CO2 generated during the limestone neutralisation reactions rise above the slurry surface on top of the tanks and carry some entrained solution / slurry.

These off gases from the iron / arsenic precipitation tanks (1 and 2) are vented via the iron / arsenic precipitation tank fans 1 and 2 and fed to the iron / arsenic scrubber.

The iron / arsenic scrubber is a Venturi type scrubber. The off gases are cooled and scrubbed of the entrained solution / slurry in the scrubber. The clean gases are emitted to the atmosphere.

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14.1.2.7 Counter Current Decantation

Counter current decantation (CCD) washes the iron / arsenic stage discharge slurry with process water using two stages of thickeners operating in counter current mode. The remaining soluble metals in solution exiting the iron / arsenic precipitation circuit are washed from the slurry and report to CCD 1 overflow. The slurry discharging from CCD 2 underflow has the soluble metals washed from the slurry to sufficiently low levels to feed into the cyanide leach circuit.

CCD thickener 1 overflow solution gravitates into the CCD thickener 1 overflow tank. The duty CCD thickener 1 overflow pump transfers the CCD thickener 1 overflow solution to the neutralisation circuit. The CCD thickener 1 underflow pump transfers the thickener underflow slurry to CCD 2 mix tank. Process water is added in the CCD 2 mix tank as wash solution to wash the solids. Diluted flocculant solution is added in the CCD 1 and 2 thickener feeds to aid in the settling of solids in the thickeners. Duty CCD thickener 2 underflow pump transfers the underflow slurry from the CCD thickener 2 to the pre-leach tank.

14.1.2.8 Cyanide Leach, Carbon Adsorption and Detoxification

The cyanide leach circuit consists of one pre-leach tank and two leach tanks. Slurry is received in the pre-leach tank from the duty CCD thickener 2 underflow pump and flotation tails. The pre-leach tank has a residence time of nominally 10 minutes and is used to raise the pH of the slurry to pH 10–11 prior to the slurry entering the leach tanks where cyanide is added for gold leaching.

The leach tanks have a total residence time of up to six-hours and slurry flows through the leach tanks by gravity and discharges the final leach tank to enter the carbon adsorption circuit. The leach tanks operate at 30% solids concentration and have low-pressure air and oxygen, from the Air Liquide oxygen plant, added to maintain sufficient oxygen in solution for gold leaching.

The carbon adsorption circuit consists of six agitated tanks with a total residence time of up to 12-hours. Each tank contains activated carbon to adsorb the leached gold contained in solution. Slurry flows by gravity from tank 1 to tank 6 and discharges into the detoxification circuit. Carbon flow is counter-current to slurry and therefore is transferred stage wise from tank 6 through to tank 1, using dedicated recessed impeller pumps. Each tank has interstage screens installed so that the carbon remains in each tank and does not follow the direction of the slurry flow.

Gold is loaded onto the carbon as it moves from tank 6 to tank 1 and reaches its maximum loading in adsorption tank 1. The loaded carbon is pumped from adsorption tank 1 to the loaded carbon screen where spray water on the screen washes the carbon prior to it entering the elution column for carbon desorption and recovery of gold through the refining circuit.

Slurry exiting adsorption tank 6 flows to the detoxification circuit where destruction of the residual cyanide contained in the slurry occurs. The detoxification circuit consists of one tank with a total residence time of one-hour. Air and sodium metabisulfite are added to the circuit to destroy the residual cyanide down to a concentration of less than 5.0 ppm CNWAD. Residual copper in the slurry catalyses the cyanide destruction process.

14.1.2.9 Carbon Desorption and Refining

The carbon desorption method selected is a split AARL elution. A common stainless steel column is used for acid wash, cold cyanide strip for copper, when required, and a hot gold elution cycle to recover gold. The elution column is a 6 t column and is designed to handle the stripping of three carbon batches per day. Loaded carbon enters the elution column via the loaded carbon screen.

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The first step of stripping the carbon is an acid wash using nitric acid solution to remove loaded impurities such as calcium. After the acid wash, a pre-soak solution is added to the elution column prior to commencement of the eluent recycle for initial stripping of copper, when required, followed by a hot elution cycle to strip gold from the carbon.

Pregnant eluate is collected in the pregnant eluate tank and pumped through electrowinning cells with gold metal plated out onto stainless steel cathodes. Smelting of gold recovered from the stainless-steel cathodes is conducted in the gold refinery.

Desorbed carbon from the elution column is regenerated through a horizontal diesel fired rotary kiln to remove organic material loaded onto the carbon.

14.1.2.10 Neutralisation and Tailings

Slurry from cyanide destruction and the CCD 1 thickener overflow solution are neutralised with lime to precipitate residual metals in solution. Air is added for the oxidation and removal of ferrous iron and manganese.

Normally the two neutralisation tanks operate in series. Discharge from the neutralisation feed box gravity flows into neutralisation tank 1 prior to overflowing into neutralisation tank 2. Discharge from neutralisation tank 2 gravitates into the tailings thickener mix tank.

The first neutralisation tank is equipped with a sodium metabisulfite addition system, and this allows it to be used for the detoxification step when the normal detoxification tank is bypassed for maintenance or descaling. Both neutralisation tanks can also be bypassed as required to allow for maintenance.

The discharge slurry from neutralisation flows by gravity into the tailings thickener mix tank before overflowing into the tailings thickener. Tailings thickener overflow water overflows directly into the process water storage tank. The underflow slurry from the tailings thickener is pumped to the agitated tailings tank. The discharge slurry from the tailings tank is pumped to a TSF on a continuous basis via the 4.3 km long tailings pipeline.

A schematic flow sheet of the process is shown in Figure 14.4 including the flotation circuit addition.

14.1.2.11 Tailing Storage Facility

The process tailings slurry is deposited into the TSF for final storage. Operators will alternate the location within the facility where the tailings are deposited to maximise the storage and dewatering within the facility.

In the TSF the solids compact and reject excess water which is recovered for recycling to the process plant. The controlled deposition of tailings at alternating locations around the perimeter of the TSF creates a pond that collects water, which decants from the tailings slurry as it settles and compacts. This decant water collected within the pond area is recycled to the process water system tank via the tailings water reclaim pumps.

The TSF is developed and constructed in stages ahead of requirements.

14.1.2.12 Reagents

There are ten major reagents used in the process plant, listed as follows:

•Oxygen

•Sulfuric acid

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•Limestone

•Sodium hydroxide

•Flocculant

•Sodium metabisulfite

•Milk of lime

•Sodium cyanide

•Nitric acid

•Antiscalant

The flotation plant has the following main reagents:

•Frother

•Collector

•Copper Sulfate

All reagents are delivered in bulk tankers, containers, or bags with storage on site. Any reagents that require dilution or mixing prior to use are prepared on site on a batch wise basis, as required. Oxygen is produced on-site supplied from an Air Liquide owned and operated oxygen plant under a gas supply agreement. Additional oxygen can be delivered as liquid into on-site storage.

14.1.2.13 Utilities

The major utilities used in the process plant are as follows:

•Iron / arsenic low-pressure air

•CIP leach low-pressure air

•Plant air

•Instrument air

•Raw water

•Fire water

•Potable water

•Process water

•Diesel fuel

These utilities are reticulated throughout the process plant to their end user.

14.2 Oxide Heap Leach Processing

The oxide heap leaching and associated facilities were commissioned in the second half of 2010 and initial gold production was achieved in Q4’10. The process was originally designed to treat approximately 6.0 Mtpa of ore by three-stage crushing (primary, secondary, and tertiary) to 80% passing 12.5 mm, agglomeration and heap leaching on a lined heap leach pad with dilute alkaline sodium cyanide solution. Gold is recovered through a carbon-in-column (CIC) system, followed by stripping of metal values from carbon, electrowinning and melting to yield a doré (containing gold and silver) suitable for sale. Control of copper in leach solutions is undertaken in a sulfidisation, acidification, recycling, and thickening (SART) plant which also regenerates cyanide. The process flow sheet is summarised in Figure 14.5.

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14.2.1 Oxide Heap Leach Performance

Since commissioning through the end-of-December 2021, an estimated 55.1 Mt of oxide ore was placed on the heap at an average grade of 1.35 g/t Au.

At the end-of-December 2021, a total of approximately 1,837 koz had been produced as bullion.

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Figure 14.5 Heap Leach Process Flow Sheet

Anagold, 2016

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15 INFRASTRUCTURE

The following section describes the infrastructure supporting the CDMP21TRS Reserves Case.

15.1 Introduction

The facility infrastructure supports the mine, and process areas of oxide heap leach and sulfide plant. The existing infrastructure, and the tailings storage facility (TSF) and heap leach pad area when the planned expansion are complete will be sufficient for the current Mineral Reserves. The infrastructure for the addition of flotation to the sulfide plant will be supported by the existing facility infrastructure with some components modified to meet the addition of the flotation circuit. The flotation circuit is located within the sulfide plant foot-print adjacent to the grinding circuit building.

The location of the processing facilities, Çöpler mine, Ardich Reserve pit, TSF, and the haul road from Ardich to Çöpler is shown in the site plan in Figure 15.1.

Figure 15.1 Çöpler Project Plan

Anagold, 2022

The current leach pad consists of four phases designed to accommodate approximately 58 Mt of oxide ore heap with a nominal maximum heap height of 100 m above the pad liner. The additional two phases (5 and 6), with a capacity of 20 Mt will be constructed during 2022 to 2024 to accommodate the remainder of the Ardich Reserve. The phase 5 pad construction has been approved by the Ministry of Environment, Urbanisation, and Climate Change (MoEUCC).

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The TSF is developed and constructed in stages. TSF 1 phase 3 has been constructed and approval for use was received in February 2021 by the MoEUCC. Ongoing work in ensuring sufficient long-term capacity for storage of tailings has been undertaken. Studies by Anagold have determined, that the effect of the addition of the flotation plant to the sulfide plant circuit would result in an increase in the solids content and improvement in the final settled density based on an increase in the rate of tailings consolidation.

Construction and development of TSF 1 will provide storage of tailings for up to 70.8 Mt, more than sufficient to accommodate the CDMP21TRS tailings to be produced.

A PFS level study (TSF 2) has been carried out that identifies approximately 13.4 Mt additional tailings storage capacity in a site adjacent to TSF 1, should it be required in the future.

15.1.1 Existing Infrastructure

The existing site infrastructure supporting the existing operation includes the following:

•Site security gate and guard station

•Site administration building

•Site warehouse

•Site assay laboratory

•Container or modular type offices

•Cyanide receiving and mixing system

•Site kitchens and eating areas

•Site single living dormitory with adjacent multi-purpose room

•Site family housing

•Contractor (mining) dormitories, kitchens, and offices

•Site raw water wells, pumping system and storage tanks

•Site potable water treatment and distribution system

•Two sanitary wastewater collection and treatment systems

•Sulfide maintenance building

•Sulfide control rooms

•Combined oxide and sulfide gold refinery building

•Sulfide process buildings:

•Grinding building

•Pressure oxidation (POX) building

•Carbon desorption building

•Tailings pump building

•Main control room and electrical building

•HV switchyard electrical building

•Crusher electrical building

•POX flocculant building

•Limestone building

•Potable water booster pump house

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•Reagent building

•Tailings and process water pump house

•Plant and instrument air compressor building

•Counter current decantation (CCD) electrical building

•Reagent dry storage

•Leach air compressor building

•Aw water pump building

•Lime slaking (MOL) building

•Fe/As air compressor building

•Emergency diesel generators building

•TSF reclaim electrical building

•TSF drainage tank electrical building

•TSF OD-UD pond electrical building

•CIP CCD ablutions block

•Pump shelters with monorails

•Carbon elution building – electrical room

•Raw water bores P/P house and electrical building

•Gatehouse

•Fire water pump house

•Community relations centre

•Raw water wells

15.1.2 Flotation Building

The flotation circuit is an insulated engineered building. The building is equipped with an overhead crane for flotation cell and pump maintenance. Flotation reagent mixing and distribution are contained in a lean-to off the main flotation building.

15.2 Site Water Management

15.2.1 Hydrology Background

The only perennial surface water in the vicinity of the Çöpler Mine is the Karasu River flowing in the northern and western part of the area. All other valleys are either ephemeral streams or dry valleys. The average flow rate of the Karasu River measured at the Bağıştaş / Karasu Gauging Station in the upper Euphrates Basin, is approximately 145 m³/sec, draining an area of 15,562 km². A hydroelectric dam (Bağıştaş -1 Dam) was built on the Karasu River downstream of the mine site. When the reservoir is at high levels the impoundment will extend into the very lower reaches of both the Çöpler and Sabırlı Creeks and the maximum inundation elevation will be 916.5 m as it is released into the spillway. The Çöpler and Sabırlı streambeds in the Project area do not flow perennially. They both discharge into the Karasu River. The drainage area of the Sabırlı Creek is approximately 35 km² and that of the Çöpler Creek is approximately 10 km².

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The project submitted a Five-Year Water Management Report in December 2019, prepared by SRK Danışmanlık ve Mühendislik A.Ş., as part of the Environmental Impact Assessments (EIA) conditions. This report benchmarks the expected results with those achieved. Overall results achieved were generally as predicted. In 2020, as part of updating the EIA, further hydrogeology studies have been undertaken by SRK Danışmanlık ve Mühendislik A.Ş. The report has updated the surface water and hydrological models based on actual data over the operating period of the mine to improve the model.

15.2.2 Site-Wide Surface Water Hydrology

Existing mine site facilities are located primarily within the Çöpler and Sabırlı Creek watersheds immediately upstream of their confluence with the Karasu River. Site-wide surface water management for the included diversion facilities consist of a network of diversion channels and retention structures to minimise storm water run-on to the mine site facilities to prevent mine-impacted storm water run-off from exiting the site and discharging to the Karasu River.

The sub-basin areas, characterisation of the surface run-off conditions, and design rainfall data were used to construct the existing conditions hydrology model. The hydrology analysis utilised HEC-HMS software to develop estimates of the peak flow rates and volumes generated by the existing watersheds.

15.2.3 Surface Water Management Structures

Engineered surface water management structures are constructed to minimise effects of storm water run-off to critical mine facilities and to control the release of mine-impacted water to the environment. A combination of interim and permanent diversion channels and retention ponds are utilised to achieve these goals. Interim structures will be reclaimed at closure while permanent structures will remain in place post-closure. Other flood control structures were developed to control or direct run-off away from pit crests and are planned for run-off that does not discharge to surface water drainages or streams and therefore do not require lining.

Sediment ponds to control run-off and sediment release are lined based on the EIA commitments. Interim diversion channels are designed to convey the 25-year storm event with 1.5 m of freeboard and the 100-year storm with no freeboard. Permanent diversion channels are designed to convey the 100-year storm with 0.5 m of freeboard. Lined sediment ponds are downgradient of the waste dumps and are sized to contain the 100-year run-off volume with an emergency spillway to safely discharge the peak flow. The TSF is designed to contain the volume generated by the 24-hour probable maximum precipitation (PMP) within the operating freeboard.

15.2.4 Fresh Water Supply

Fresh water is supplied by existing wells to the site, supporting the operation. Figure 15.2 shows the location of the mine water extraction wells. An additional three wells were developed in 2018, wells WM-45, WM-46 and WM-47, to increase water supply for the Project. Two raw water storage tanks support the demands of the heap leach and sulfide process equipment and the fire water requirements.

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Figure 15.2 Mine Water Supply Well Locations

Anagold, 2020

15.2.5 Potable Water Treatment

The site is serviced by a potable water treatment system and distribution system. The system consists of multi-media filtration, carbon filtration, ultraviolet (UV) disinfection system (plus further softening and reverse osmosis for water used in the dining room), which directly feeds the site potable water distribution system.

15.2.6 Waste Management

Waste will be generated from multiple sources such as human waste, food spoilage, and process and maintenance wastes.

Hazardous wastes will be contained, packaged, and disposed of in accordance with local, regional, and national regulations. Non-hazardous wastes will either be buried on-site or transported off site to the appropriate processing site in accordance with local, regional, and national regulations.

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15.3 Power to Site

The existing 154 kV line provides power to the mine and process plant. The following structures are associated with site power distribution:

•HV switchyard 154 kV

•Main electrical building

•Oxygen plant substation

•CCD electrical building

•Crushing electrical building

•Grinding electrical building

•Carbon elution electrical room

•TSF area electrical buildings

•Bore field area electrical building

15.4 Emergency Backup Power

Motors and loads for certain critical equipment and systems were identified as requiring power in the event of a utility outage. A load shedding scheme is applied to feed critical electrical users automatically in the event of a utility outage.

Generators are diesel fuelled with a minimum of eight-hours of diesel storage based on generators operating under full load.

15.5 Communications

The Project uses networks for the DCS, precious metals search (PMS), the integrated process related and security CCTV system, security systems (access control / card reader), information technology (IT) and telephones and communication between the DCS and packaged control systems.

Single mode fibre and copper cabling is distributed within the sulfide plant area and selected buildings for the tailing pipeline and dam.

15.6 Site Roads

The Çöpler project has access provided via the main access road and sulfide plant roads.

Generally, site roads have an overall width of 6 m and provide everyday operational access for large trucks or facility access for site personnel vehicles. These roads are limited to a maximum grade of 9%. All roads are compacted hardstand surfaced with 100 mm wearing course and cross-sloped to provide positive drainage.

15.7 Plant Fire Protection System

A separate plant fire protection system is provided for the sulfide facility and will include the flotation building.

A combined sprinkler, hose reel and hydrant underground piping system is provided for the active fire protection of the facility.

A gas-based fire suppression system is used in the main control and electrical building.

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15.8 Site Water Management

15.8.1 Hydrology Background

A hydroelectric dam (Bağıştaş -1 Dam) was built on the Karasu River downstream of the mine site. When the reservoir is at high levels the impoundment will extend into the very lower reaches of both the Çöpler and Sabırlı Creeks and the maximum inundation elevation will be 916.5 m as it is released into the spillway.

The Çöpler and Sabırlı streambeds in the Project area do not flow perennially. They both discharge into the Karasu River. The drainage area of the Sabırlı Creek is approximately 35 km² and that of the Çöpler Creek is approximately 10 km2.

The project submitted a Five-Year Water Management Report in December 2019, prepared by SRK Danışmanlık ve Mühendislik A.Ş., as part of the EIA conditions. This report benchmarks the expected results with those achieved. Overall results achieved were generally as predicted.

In 2020, as part of updating the EIA, further hydrogeology studies have been undertaken by SRK Danışmanlık ve Mühendislik A.Ş. The report has updated the surface water and hydrological models based on actual data over the operating period of the mine to improve the model.

15.9 Heap Leach Facility

The heap leach includes the leach pad and collection ponds that consist of process ponds and a storm pond. The current leach pad consists of four phases and designed to accommodate approximately 58 Mt of oxide ore with a nominal maximum heap height of 100 m above the pad liner. The additional two phases, 5 and 6, with a capacity of 20 Mt were approved in October 2021 and phase 5 has received construction approval from MoEUCC in November 2021. The heap is stacked in 8 m thick horizontal lifts at the natural angle-of-repose with intermediate benches to achieve an overall heap slope of 2H:1V.

15.9.1 Heap Leach Pad Development

The heap leach facility pad development is in six phases, and is in the same geographical area, adjacent to the Çöpler open pit as shown on Figure 13.8. The heap leach phases 1 to 4 are completed.

The remaining phases of pad development 5 and 6 are yet to be constructed and will have a combined capacity of approximately 20 Mt.

The phase 5 (15 Mt capacity) was approved for pad construction in November 2021.

The phase 6 (5 Mt capacity) sits above phase 4B and 5 and will be the last to be constructed and stacked. Approvals and construction will be scheduled well in advance of being required for ore stacking and leaching.

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15.10 Tailings Storage Facility

The existing tailings storage facility (TSF) at the Çöpler mine was designed by Golder Associates Inc. (Golder) with support from Golder Associates Turkey, Ltd (Golder Turkey). The TSF initial design was developed to provide a capacity of 45.9 Mt through six phases with a crest elevation of 1,265 m. The TSF was permitted through submission of a Turkish Design Application Report to the MoEUCC and subsequently approved based on the design through phase 5.

Anagold is advancing the development of the Çöpler Mine. Recently developed a prefeasibility level design for an additional TSF, referred to as TSF 2 in the valley adjacent and to the north of the existing TSF 1. Both TSF 1 and TSF 2 were included in the EIA submitted by Anagold in 2014.

The current designs for TSF 1 and TSF 2 are within the 2014 EIA boundaries, except for a small portion of TSF 1, phase 7. Expansion beyond phase 3 of TSF 1 is currently limited by the construction and re-routing of a new road to Sabırlı Village as well as purchase of some small tracts of private land located within the phase 4 limits on the east side of the existing road to Sabırlı Village. Construction of the new Sabirli Village road commenced in Q3’21 and is on schedule. Acquisition of the private land parcels have substantially progressed through regulatory processes.

Based on the prefeasibility design, TSF 2 has capacity for 13.4 Mt. To maximise capacity of TSF 1, phase 7 was developed as part of the design to a crest elevation of 1,280 m at a conceptual level and to support further planning, including planned updates to the site Environmental Assessment. Select engineering evaluation of phase 7 has been completed to support future planning including updated stability analysis, water balance, and consolidation modelling. Anagold’s preference is to continue with development of TSF 1 phase 4 and to consider other options, if required depending on tailings capacity requirements, due to the higher capital costs related to construction of TSF 2 at this time. Without construction of TSF 2, TSF 1 alone provides for tailings capacity of up to 70.8 Mt through phase 7.

Figure 15.3 through Figure 15.7 show the revised TSF 1 design for phases 4–7, and the TSF 2 design.

15.10.1 TSF Development and Summary of Current Operations

Construction of phase 1 of TSF 1 began in December 2016 and was completed in November 2018 with commissioning of the sulfide plant. Tailings were deposited initially from the emergency spigot and then typically from two to three spigots around the perimeter of the 1,190 m crest of the phase 1 embankment. The tailings initially have exhibited a solids content on the order of 24%. During the first two years of operations 4–5 m of water has been present over the top of the tailings surface. Reclaim water was managed by pumps on a rail-mounted sidehill reclaim system. The second raise, or phase 2 of TSF 1, was completed in April 2020 and construction of phase 3 is ongoing. The management of reclaim water has improved in the past year and currently the tailings surface is nominally 3 m below the top of the tailings water. A bathymetry survey was completed on 11 September 2020 and indicated a tailings average dry density of 0.68 t/m3.

The reclaim water management system was converted to a conventional pontoon system accessible for maintenance from ramps constructed within the northern portion of the impoundment. Based on additional tailings testing completed in early-2020, the solids content of the tailings has improved to approximately 28% as a direct result of improved throughput stability at the sulfide plant, improvements to type of flocculants used and process control in the tailings thickener. As part of this tailings testing in early-2020, Golder evaluated the effect of the addition of the flotation plant to the sulfide circuit. The testwork indicated an increase in the solids content to 34% and improvement in the final settled density based on an increase in the rate of tailings consolidation.

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15.10.2 Site Classification

The facilities are classified in accordance with the Canadian Dam Association (CDA) guidance (2013 Edition) as ‘High’ for the operational and post-closure phases. The ‘High’ classification is the third lowest in terms of risk with the dam classes being from least risk to greatest risk: Low, Significant, High, Very High, and Extreme.

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Figure 15.3 Phase 4 – Top of Embankment and Impoundment Grade

Anagold, 2020

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Figure 15.4 Phase 5 – Top of Embankment and Impoundment Grade

Anagold, 2020

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Figure 15.5 Phase 6 – Top of Embankment and Impoundment Grade

Anagold, 2020

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Figure 15.6 Phase 7 – Top of Embankment and Impoundment Grade

Anagold, 2020

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Figure 15.7 TSF 2 Impoundment Grading

Anagold, 2020

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15.10.3 Monitoring and Inspection

An Operational, Maintenance, and Surveillance (OMS) Plan was prepared by Golder with input and support from Anagold. The OMS Plan was prepared in accordance with the Turkish mining regulations (MoEU 2017) with additional guidance published by the Mining Association of Canada (MAC 2019). The OMS Plan is a ‘living document’ that is updated on an annual basis. In addition to providing the basic guidance for the management of process fluids, the OMS Plan does the following:

•Summarises the roles and responsibilities of Anagold personnel.

•Presents a description of the facility and pertinent design details.

•Provides maintenance and surveillance parameters and procedures.

•Outlines abnormal operating conditions.

•Details emergency preparedness and response protocols.

•Presents a conceptual closure plan.

The OMS Manual provides a documented framework for action, as well as a sound basis for measuring performance and demonstrating due diligence. It is intended to be a dynamic document that is reviewed and revised by site personnel and the Engineer of Record (EoR) on an annual basis and as operating conditions require. The OMS Manual includes a requirement for the annual dam safety inspection prepared by the EoR which includes a series of inspections at site that is documented in an annual Dam Safety Inspection Report. The first annual inspection for TSF 1 was conducted in Q4’19. The results of the inspection and data review indicated that the Çöpler TSF 1 is in good condition and operating in general accordance with the intended design of the facility. A review of the instrumentation indicated normal data trends and no unanticipated abnormal readings or ‘triggering events’ observed. Of the action items included in the report, none were considered serious in nature or otherwise a concern to the safety of the Çöpler TSF.

The TSF is inspected daily for signs of stress or damage. Daily and monthly operating data is collected on site and provided in a monthly report. The report estimates the settled solids volume in the TSF based on estimated bulk densities and provides for a comparison of actual tailings and water pool elevations compared to estimates made by Golder using data from the mine and tailings production plans and from the consolidation model that predicts settlement of the tailings. The difference between the actual tailings elevation and predicted elevations have shown close agreement generally less than 1 m.

In addition, members of the Anagold’s HSSER team also inspect the TSF monthly. The TSF is subject to fortnightly external official audits by the Erzincan Provincial Environmental Directorate. The authorised hydraulic structures inspection company, Hidro Dizayn, is always on site during construction, on behalf of the MoEUCC. The TSF design and engineering consultant is also on site during construction to ensure quality and conformance to design.

Anagold has established an Independent Tailings Review Board (ITRB), as per leading international best practices, to review tailings facilities as part of the review and oversight process. The ITRB reports directly to the senior management at a corporate level.

15.10.4 TSF Design

The TSF at Çöpler is a downstream, mass filled, dam. The technical specifications for the construction of the TSF conform with both Turkish national requirements and accepted good practice standards for tailings facilities, including; World Bank Standards, Canadian Dam Association Safety guidelines, Mining Association of Canada (MAC) Guide and the International Commission on Large Dams (ICOLD) to the Management of Tailings Facilities.

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Both the TSF 1 and TSF 2 designs consist of fully lined impoundments, including a compacted earth and rockfill embankment. The TSF 1 and TSF 2 designs include the following primary components:

•A compacted earth and rockfill embankment with a zoned upstream granular filter protection system. Both facilities will have 1 m of freeboard under their crest elevations and are designed to contain the PMP storm event. The downstream face of the ultimate embankments will be constructed at a composite slope of 1.7H:1V. The upstream face of the embankment will be constructed at a slightly shallower slope with slopes of 2.0H:1V to facilitate placement of the filter layers and liner system and a resultant composite slope on the order of 2.6H:1V after considering the operational benches. The filter layers and low-permeability soil layers are designed to be 1.5 m thick, as measured perpendicular to the slope. Measured horizontally, the layers are designed at 3.3 m wide each.

•TSF 1 is a downstream raise construction which will consist of seven phases (six raises)

•TSF 2, if constructed, is a downstream raise construction and is currently designed to be constructed in one phase.

•A composite liner system consisting of a 2 mm thick, double-sided, textured high-density polyethylene (HDPE) geomembrane and geosynthetic clay liner (GCL) over a low-permeability soil (i.e., clay) liner system that provides an equivalent protection to that provided by 5 m of a geologic barrier with k <10-9 m/s. A GCL is also substituted with low-permeability clay on select slopes steeper than 3H:1V as allowed by Turkish regulations.

•An impoundment gravity flow underdrain system for collection and monitoring of naturally occurring seeps and springs.

•An impoundment overdrain system for the collection and management of tailings seepage water through natural consolidation and drainage of excess process water.

•Perimeter roads and benches within and around the impoundment area for access and tailings distribution / reclaim water pipes.

•Tailings delivery and distribution system.

•Reclaim Systems.

15.10.5 Seismic Deformation Evaluation

The current deformation model provides the deformations under seismic loading conditions for a TSF 1 with 1,264 m crest elevation, which corresponds to phase 6 in the current design. Based on the average predicted deformations and the expected levels of liner strain, the TSF 1 phase 6 embankment is expected to remain stable when subjected to the design strong motion events. Simple deformation analysis by Bray and Travasarou (2007) was performed to assess the magnitude of earthquake induced movements on the phase 7 TSF 1 Embankment.

No deformation analysis was performed for TSF 2 considering it is a smaller dam and has a lower embankment height than TSF 1 and because of the similarities in design and foundation conditions. TSF 2 deformations are expected to be smaller than TSF 1 and in the acceptable deformation range as per the design criteria.

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15.10.6 Tailings Consolidation and Capacity

Golder updated the tailings consolidation modelling to include the TSF 1 and TSF 2 joint operations and to account for the tailings characteristics obtained from 2020 laboratory tests on POX and Flotation tailings. The updated consolidation model also included the current mine plan. In the model, TSF 1 was first filled to elevation of 1,219 m (to the limits of phase 3 with a crest elevation of 1,220 m allowing for 1 m freeboard) and then tailings deposition was switched to TSF 2 and tailings in the TSF 1 was let to rest until TSF 2 is filled for a period of approximately 3.4 years. The rest period in TSF 1 increases the tailings density from 0.85 t/m3 to 1.08 t/m3 due to the natural consolidation and results with an average settlement on the order of 7 m which results in a capacity gain of 3.2 Mm3 in TSF 1. The model results show that with the current mine plan and tailings characteristics TSF 1 and TSF 2 would have approximately 76 Mt and 14 Mt tailings capacity, respectively, over approximately 19.2 years of TSF 1 filling time.

The tailings tonnage estimate requires the sulfide plant feed to be adjusted to allow for the limestone added during processing for pH control. The limestone reacts with the acid to form gypsum. The applicable factor is 1.146. Commissioning of the Flotation Plant commenced in January 2022, once operational will also directly contribute to the tailings placed.

Based on the updated consolidation analysis and assumptions on the mine plan, tailings characteristics, and operational plans as stated herein, approximately 90.6 Mt of tailings can be stored in TSF 1 and TSF 2 combined. The average dry tailings density expected at end of filling is 1.17 t/m3 and 0.89 t/m3 in TSF 1 and TSF 2, respectively.

15.10.7 TSF Schedule Assumptions

The key assumptions related to the ongoing construction and expansion of TSF 1 as follows:

•Phase 4:

•There is a parcel of private land located east of Gully B that has not yet been purchased. If the private land cannot be purchased, contingent measures are in place to allow phase 4 to be constructed.

•Construction of the new Sabirli Road is required for Phase 4 to be completed. Construction commenced in Q3’21 and at the time of this report is on schedule for timely completion

•Phases 5 to 7:

•The design of the access roads and utility corridor for phases 5 to 7 considered construction of the haul road developed as part of the TSF 2 design which requires nominally 4 Mm3 of rock fill. This design has been shown starting with phase 5. If TSF 2 were not developed, the access road and utility corridor could be further optimised depending on the extent of development. If only phase 5 were to be developed, the access could be provided by a much smaller ramp. If phase 6 and/or phase 7 are developed, then a route like that shown would be required.

•Schedule:

•The current mine plan and schedule provides capacity within phase 3 through Q1’23, which generally required that construction of phase 4, would start in 2021. Phase 4 embankment works were progressed approximately 66% during 2021 and embankment and impoundment works are scheduled to continue through 2022. As an ongoing option, the TSF 2 design and approval activities were also substantially progressed during 2021

Sabirli road completion is scheduled for completion in time to achieve phase 4 TSF construction works.

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15.10.8 Further Work

There are opportunities that may offer significant reduction in capital costs with consideration of the following:

•Alternative TSF Considerations – The dam capacity to fill ratio for TSF 2 was approximately 1:25, which is significantly lower than TSF 1 due to the narrow and small valley where it is located. Several other options were identified in the CDMP21TRS Siting Study that would provide for reduced capital costs. Sites identified as TSF 4 and TSF 7 were determined to have dam capacity to fill ratios of 1:3.2 and 1:1.9, respectively based on conceptual designs only. Of the other sites considered in the CDMP21TRS Siting Study, TSF 4 was ranked second behind TSF 7 based on several environmental and social considerations namely due to its proximity and location with the Bağıştaş area, however, TSF 4, provides a significantly greater potential storage capacity with less fill required. TSF 7 would be highly visible to the Sabirli community but on the opposite side of Sabirli creek.

•Waste Rock Encapsulation – There are opportunities to consider encapsulation of potentially acid-generating (PAG) waste rock within portions of the downstream embankment within the limestone. A study commenced in Q3’21 to evaluate this potential

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16 MARKET STUDIES

16.1 Marketing and Meta Prices

The markets for gold and silver doré are readily accessed and available to gold producers. Currently, 100% of the gold and silver is delivered to the Istanbul Gold Refinery. Copper precipitate is currently produced from the sulfidisation, acidification, recovery, and thickening (SART) plant and sold into local markets in Turkey. The sulfide plant does not currently include a copper circuit. Provisions have been made in the plant design to include the copper circuit in the future if market conditions warrant.

Metal prices for the CDMP21TRS economic analysis were estimated after analysis of consensus industry metal price forecasts and compared to those used in other published studies. The metal prices selected have taken into account the current project life. The metal prices used for the economic analysis, shown in Table 16.1, are considered to be representative of industry forecasts.

Table 16.1 CDMP21TRS Economic Analysis Metal Price Assumptions

Metal Price	Units	2022	2023	2024	2025	Long-Term
Gold	$/oz	1,800	1,740	1,710	1,670	1,600
Silver	$/oz	24.00	23.00	22.00	21.00	21.00
Copper	$/lb	4.00	3.80	3.80	3.80	3.40

No external consultants or market studies were directly relied on to assist with the sales terms and commodity price projections used in the CDMP21TRS. The QP for this Section 16 agrees with the assumptions and projections presented.

16.2 Contracts

Anagold contracts the mining operations to a Turkish mining contractor. The contract contains provisions for escalation / de-escalation of fuel prices, foreign exchange rates, haul grade and distance and Turkish inflation. The terms and prices for the mining contract are within industry standards for mining contracts.

16.3 QP Opinion

Macroeconomic trends, taxes, royalties, data, and assumptions, interest rates, marketing information and plans are outside the expertise of the QP and are within the control of the registrant (see Section 25).

The CDMP21TRS QP considers it reasonable to rely on SSR because SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

Additionally, the Project has been in operation for a number of years, and following a review of the current supplied information, the opinion of the QP is the current plans and input parameters appear adequate for use as inputs to the CDMP21TRS.

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17 ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS

The following section describes the existing Project operations.

17.1 Environmental Studies and Material Impacts

The Çöpler mining and processing operations involve open pit mining from multiple pits, construction of multiple waste dumps to accommodate mined materials, processing of oxide ores and placement on a heap leach pad, and processing of sulfide ores with placement of tailings in a tails storage facility (TSF). These activities and facilities are carried out on treasury, pasture, and forestry lands, including some private lands.

In addition to the direct impacts on the involved lands, the operations impact on the surrounding lands and the local communities. Physical impacts may include changes to local surface and groundwater (including potential pollution), air quality impacts particularly from dust, and increased noise and vibration from mining and processing operations.

Operation of the Çöpler mining and processing facilities, and subsequent mining at Çakmaktepe, has been investigated and authorised by means of a series of Environmental Impact Assessments (EIAs), with positive decisions obtained from the Turkish Ministry of Environment, Urbanisation, and Climate Change (MoEUCC). These EIAs include specific actions designed to address all material impacts of the mining and processing operations. Anagold has remained in compliance with all aspects of the EIA and operating permits throughout the history of the Project.

The original 2008 EIA obtained on 16 April 2008 included three main open pits (manganese, marble contact, and main zones), five waste rock dumps (WRDs), a heap leach pad, a processing plant, and a TSF. The 2008 project description involved only the oxide resources.

The Çöpler project started its open pit and heap leach operation in 2010 and first gold was poured in December 2010. Additional EIA investigations have been submitted and approved, as required, to support ongoing mining and processing operations, including:

•EIA to allow operation of a mobile crushing plant approved 10 April 2012.

•EIA to allow waste dump capacity expansion, oxide capacity expansion to 23,500 tpd and a sulfidisation, acidification, recovery, and thickening (SART) plant approved 17 May 2012.

•EIA to allow the sulfide plant and heap leach area expansion approved 24 December 2014.

•EIA to allow the Çakmaktepe satellite pits expansion approved 26 January 2017.

•EIA to allow a Çakmaktepe capacity increase approved 9 August 2018.

•EIA to allow a second capacity expansion, including heap leach pads 5 and 6. TSF expansion and operation of a flotation plant approved 7 October 2021.

In addition, pending EIA processes include:

•EIA to allow Çakmaktepe second capacity increase to include initial mining from Ardich with EIA description file. The EIA project description file was submitted in October 2020 and a Public Hearing was held in November 2020. All public institutions gave positive feedback regarding the report and the approval process is ongoing with the MoEUCC.

•A review and evaluation meeting was held on the 15 December 2021 for Çakmaktepe EIA, and the approval process is ongoing.

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After the EIA positive decisions, additional permits and licences were required to be issued by government agencies consistent with the Turkish governing laws and regulations. These include land access permits (pasture and forestry); environmental permits and licences; workplace opening and operating permits; and licences and certificates. The status of project permits, and operating licences is documented in Section 3 of this report.

In the period following the receipt of the 2008 EIA permit, Anagold has conducted further technical studies to supplement the Turkish EIA studies and to establish plans and procedures to manage potential project impacts and meet IFC requirements. Significant operational management plans established as a result of these prior and ongoing studies include:

•Non-mining Wastes Management Plan

•Mining Waste Management Plan

•Water Resources Management Plan

•Biodiversity Management Plan

•Soil Management Plan

•Air Quality and Emissions Management Plan

•Mine Closure and Rehabilitation Plan

•Environmental Management System Framework

•Environmental Noise and Vibration Management Plan

•Hazardous Substances Management Plan

•Mine Closure Framework

•Resource Efficiency and Pollution Prevention Management Plan

•Cyanide Management Plan

17.2 Physical Features

The project site is in a transition region between Central and Eastern Anatolian climates. The region has a continental climate, where summers are hot and dry, and winters are cold and relatively humid. Owing to the mountain ranges bordering Erzincan Province on all sides, the region has a milder climate than the neighbouring provinces.

The long-term annual average precipitation for the Project site is 367 mm, including snow in the winter months. The annual average wind speed is 2.6 m/s. Maximum wind speeds are observed in spring. The prevailing wind direction is south.

The project site is in a rural area with no significant commercial or industrial air pollution sources. Scattered slag piles and ore extraction sites remain from the former manganese mining operations.

The ambient air quality monitoring programme on site indicated that SO2 and NO2 levels, and particulate matter (PM10) and dust deposition levels in ambient air are well below the limit values defined in Turkish Air Quality Standards. Heavy metal concentrations in dust were well below the limit values defined by European Commission (EC), World Health Organisation (WHO), and Turkish standards.

The railway and the İliç-Kemaliye Road passing near the Euphrates River are the mobile sources of noise in the area. The Euphrates-Karasu River is the largest surface water body near the Project; it borders the northern edge of the Project area. Peak flow rates are observed in April and May following the snow melt and rainfalls. All other streams in the vicinity of the Project area are intermittent, flowing between March–June.

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The surface water quality within the site was investigated at various water sampling locations throughout the site. Water quality is classified from class I (very good quality) to class IV (highly polluted, poor-quality water). Sampling has indicated class IV water quality for Sabırlı and Çöpler Creeks, and Karabudak Stream. Similarly, the Euphrates-Karasu River is classified as a class IV water resource. For all streams, metal concentrations, including aluminium, iron, copper, and arsenic are high, especially in the drainage from Sabırlı and Çöpler creek catchments. Elevated metal concentrations in these catchments are attributed to natural metallic enrichment from the surrounding geology.

17.2.1 Land Use

The prevalent land use and cadastral information for the Project and its environs is presented in Figure 17.1. The land use patterns are based on maps produced by the General Directorate of Rural Services. As observed in Figure 17.1, most of the Project area consists of pastureland, treasury, and forest. The Land Use Capability Classes (LUCC) for the Project area and environs is given in Figure 17.2.

Under the LUCC system, there are three main categories and eight classes (ranging between I and VIII).

•The first category covers classes I through IV and describes lands, which are suitable for cultivation and animal husbandry. This category has few limitations, except for class IV, which requires very careful management because of its greater limitations.

•The second category covers classes V through VII, which are unsuitable for cultivation, but which can support perennial plants when intensive conservation and development practices are applied. Under controlled conditions, this land may also support grazing and forestry. The soil type included in class VII has severe limitations, preventing the growth of cultivated plants due to characteristics such as the formation of steep slopes (which are exposed to medium to severe erosion) and shallow soil layers, possessing stony, salty, and sodic texture. As such their utilisation for agricultural purposes is very limited.

•The third category contains only the class VIII, which is suitable only for wildlife, sports, and tourism-related activities.

As shown in Figure 17.2, the Project area has VI, VII, and VIII classes of LUCC. The land use types in the Project area and its vicinity are:

•Degraded forest lands and coppice

•Barren forest lands

•Agricultural lands

•Settlements

The project area and surroundings are generally of low-land use capability and not suitable for agricultural activities. Although the agricultural activities are limited in the area, there are several small gardens which belong to the local villagers.

The forests in the area are under stress due to high grazing and illegal land use practices; pasture lands are used for the purpose of grazing, but it is illegal to use forestry lands for grazing. In general, the local soil has poor fertility due to its nature and elevation such that it only supports limited species of vegetation.

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17.2.2 Biological Features

Floral species from the Irano-Turanian and Mediterranean phytogeographic regions are dominantly observed at the site. Most of the flora species are identified in the dry meadow habitats in the Project area. Ruderal habitat (such as roadsides etc.) and rocky areas follow dry meadow habitats with respect to the floristic species diversity.

Flora and fauna surveys were conducted in the framework of the 2005–2007 EBS by specialists from Hacettepe University. Biodiversity of the site has been updated by the specialists from Gazi University and Hacettepe University via three seasonal surveys during 2011–2012. A Biodiversity Action Plan (BAP) was prepared, and a BAP Report has been provided as an appendix of the Environmental and Social Impact Assessment (ESIA) Report for the Sulfide Expansion Project. The flora species were classified according to their thread status with respect to Turkish Red Data Book of Plants and the International Union for Conservation of Nature (IUCN) and European Red List (ERL) Categories and Criteria.

There are four main vegetation types in the area namely: Quercus petraea subsp. pinnatiloba; Quercus libani and Quercus brantii forests; Irano-Anatolian steppe vegetation; and wooded steppes and rock habitat, while the rest of the site is designated for main mining activities. The faunal composition of the site is considered weak.

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Figure 17.1 Current Land Use Types and Cadastral Map for the Çöpler Project

Anagold, 2021

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Figure 17.2 Land Use Capability Classes (LUCC)

Anagold, 2021

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17.3 Social and Community Plans

The EIA studies are conducted according to the format stipulated by the Turkish EIA Regulation. The scope of the Turkish EIA studies differs from the scope of international ESIA studies (as established by the International Finance Corporation’s (IFC)’s Environmental and Social Performance Standards), especially in terms of social impacts and public disclosure processes. While the social impact assessment and public disclosure processes are also parts of the Turkish EIA studies, they are treated less rigorously than in IFC standards.

Anagold has conducted further investigations to supplement the Turkish EIA studies, initially to support the original project establishment and, then subsequently, to monitor the social and community attitudes and the impacts of ongoing mining operations on the adjacent communities. The fundamental data to assess social impact is derived from direct survey of the local community members in villages impacted by the mining operation. Significant (primary) surveys have included:

•Initial survey of 51 households in three villages (Sabirli, Bagistas and Dostal) presented collectively as part of the 2009 Çöpler Gold Project Social Impact Assessment (SIA) by KORA.

•Survey of 153 households in six villages (Çöpler, Bagistas, Bahcecik, Dostal, Yakuplu and Sabirli) presented individually performed by Middle East Technical University (January 2013).

•Survey of six villages performed by UDA Consulting (December 2014).

•SIA by SRK (2015).

•Survey by TANDANS Company (2017).

•Çöpler Mine Phase 2 SIA Peer Review Report by Intersocial Company.

•Çakmaktepe 2nd Expansion Project SIA Works by SRK (Ongoing).

•Survey by TANDANS Company (Ongoing).

Anagold has considered the outcomes from the community surveys and SIA assessments as a key input to establish and monitor the social action plans associated with the Project. These are also the basis to develop a strategic and planned approach to community investment and development programmes. Some significant social and community plans and policies developed as a result of these investigations address the following:

•Community health and safety

•Local employment

•Local procurement

•Community development fund (SKF)

•Donations

•Stakeholder engagement and community relations

•Grievance management

•Environmental and social sustainability

•Training management

•Cultural Heritage

•Land access and resettlement

•Communications

The performance and effectiveness of social and community plans are monitored, reviewed, and updated, as required, to meet changing community needs and expectations.

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17.4 Mine Closure

Mine rehabilitation and closure obligations are prepared and updated annually for the Çöpler project. Scheduling and costing of the closure tasks are made in accordance with the Anagold mine plan.

Cost estimates rely on data from mine operations including labour and equipment rates, material costs, groundwater well inventories, and electronic topography data.

Closure costs are estimated using the Standardised Reclamation Cost Estimator (SRCE). The SRCE is an industry standard tool developed to facilitate accuracy, completeness, and consistency in the calculation of costs for mine site reclamation.

SRCE utilises lengths, areas, volumes, flow rates, quantities, etc., provided or estimated by the user (based on the reclamation or closure actions). Some actions require crews and fleets with productivities either provided by the SRCE default settings or those provided by Anagold to estimate the time it takes to perform the work. Where available, these times are then multiplied by labour and equipment rates provided by Anagold.

The Heap Leach Draindown Estimator (HLDE) model is another industry standard tool used for estimating heap leach pad draindown curves for reclamation bonding purposes. The HLDE inputs are derived from site-specific data.

17.4.1 Closure Cost Estimate Assumptions – Waste Rock Dumps

All slopes on the WRDs will be regraded to 2.5H:1V to prepare them for covering, scarification, and revegetation. The sequence of costs in the schedule corresponds to the assumption that reclamation will occur as soon as each WRD reaches final configuration.

Anagold plans to encapsulate all potentially acid-generating (PAG) waste rock within the WRDs as part of mining operations, leaving no PAG material on the surface or outer portions of the WRDs at closure. Therefore, although some PAG cells are currently exposed, costs for construction of a buffer layer encapsulating PAG waste rock are accounted under operational costs and no additional costs for mitigation of current configurations are included in the ARO estimates.

Per the EIA Report, waste rock management will be carried out to allow for the construction of a buffer layer to prevent degradation of seepage and these costs are accounted under operational costs. The seepage collection ponds active during the operations period will be reclaimed during closure. Seepage from the WRDs will not be monitored during closure and post-closure.

17.4.2 Closure Cost Estimate Assumptions – Pits

Berms will be constructed around the perimeter of the pit to discourage public access. There are no other physical reclamation measures assumed for the pit walls.

Rapid refilling of the pits with water is the preferred method for the western part of the pit. Costs for pit refilling by pumping flow of 66 litres per second (L/s) for four years are included in the ARO estimates.

Some PAG rock will remain exposed in the pit walls after formation of a pit lake; therefore, some reclamation work will be necessary to address the requirement (legal obligation) to cover remaining PAG materials exposed in the pit after mining ceases.

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It is assumed that areas within the pit where PAG materials are exposed will be covered with 1 m of non-PAG (or non-acid-generating – NAG) material. The PAG materials exposed within the pit walls are assumed to be located on gentle or nearly flat slopes. Additional measures (e.g., reduction of pit wall slopes in exposed PAG areas to facilitate cover placement) are not taken into consideration at this time. No PAG cover will be required below the final pit lake elevation.

17.4.3 Heap Leach Pad

All slopes on the heap leach pads will be regraded to 2.5H:1V or flatter to establish a geotechnically stable closure configuration. Following regrading, the areas will be covered, scarified, and revegetated. The ARO estimates reflect the requirement per the EIA report that identifies 2–3 m of cover placement on the heap leach pad followed by growth medium placement after the reduction of heap and pond fluid inventory.

Although not a requirement in the EIA plan, there is a provision for extending half of the heap leach pad perimeter liner to contain heap material regraded beyond the existing liner during reclamation.

East and west buttresses are considered part of the heap leach pad area. The physical reclamation of this area by growth media placement and revegetation is included as a WRD.

The 2014 EIA discusses rinsing of the heap with fresh water with no subsequent fluid management. Rinsing of heap leach pads has been shown to be typically unnecessary and potentially detrimental to long-term chemical stability of gold heap leach.

Per the approach of the HLDE model mentioned above, heap drain-down will be initially managed for inventory reduction via recirculation and active evaporation, followed by active evaporation only. Active evaporation will continue until drain-down flows are reduced to a rate amenable to management with passive evaporation.

Following active solution management, when the heap drain-down flow rate decreases to a level where it can be managed exclusively within available emergency and process pond via passive evaporation, the two ponds will be converted to evapotranspiration (ET) cells. To convert process ponds to ET-cells, the ponds will require relining followed by backfilling with select material and revegetation.

Conversion costs are calculated based on experience from multiple Nevada sites.

In scheduling costs, the cost of construction of ET-cells is included at a time when drain-down rates reach a level that will allow fluid to be managed through the evapotranspirative capacity of ET-cells.

17.4.4 Tailings Storage Facility

Anagold submitted an EIA in 2014 that included TSF 1 and TSF 2. The current designs for TSF 1 and TSF 2 are within the 2014 EIA boundaries, except for a small portion of TSF 1 phase 7. TSF 1 phase 3 has been constructed and approved for use in October 2021 by the MoEUCC. The current mine plan only requires construction of TSF 1. Long-term management costs are included in the estimate and proportioned for the size of the TSF construction.

Reclamation of the life-of-mine (LOM) TSF includes the following actions:

•Reclamation of the TSF surface by placing a traffic layer and growth media followed by revegetation.

•Reclamation of the final TSF embankment.

•Fluid management including managing drainage from the TSF and removal of water ponding on the TSF surface due to consolidation of the tailings.

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The estimate includes costs for placement of a traffic layer over the tailings material in addition to the growth media layer. The starter embankment is built at 1.5H:1V with the final embankment at 2.0H:1V. The costs of placing 1 m cover over the embankment are also included.

Costs are included for tailings fluid management crews, pumping for recirculation and forced evaporation, as well as removal of the supernatant in the period soon after the TSF operations end.

17.4.5 Other

SRCE estimates costs to demolish buildings using productivities in conjunction with building volumes, wall areas, and slab volumes. Decontamination costs are included in the estimate for a decontamination crew to pressure-wash the plant site over a nominal number of weeks.

Production wells are assumed to be closed at the end of operation of the sulfide plant and monitoring wells are assumed to be abandoned at the end of the post-closure monitoring period.

17.4.6 Monitoring

The water quality and flow monitoring schedule during the operation, closure and post-closure monitoring period includes numbers of samples, frequencies, and durations for each closure phase. The monitoring locations include the groundwater monitoring wells around the heaps, WRDs, TSF and springs as well as pit lake water quality once the rapid filling begins.

17.4.7 Closure Planning

Closure planning costs are typical industry costs for development of closure plans and studies, reporting and preparation of closure designs and engineering.

17.4.8 Construction Management

Construction management costs include one supervisor during active reclamation. Costs are included for road maintenance, which will be carried out with a water truck and grader during active reclamation.

17.4.9 Human Resources

Closure personnel include a closure general manager, environmental manager, environmental technician, security, and surveyor for whom terminal benefits are included. Under the LOM schedule, the closure general manager would be present during the years of active reclamation and closure. Camp costs are included under general and administration costs.

For solution management, the cost of the heap drain-down management crew is assumed to be shared with those of the TSF. The annual Asset Retirement Obligation reports for EOY’20 and EOY’21 have been completed.

17.4.10 Closure Schedule

The EOY’20 closure was scheduled separately for the oxide and sulfide projects according to the mine plan and is consistent with the long-term management obligations expected for the TSF.

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Heap drain-down management starts at the end of heap leaching operations in the mine plan. Ore will be sent to the leach pad until the end-of-2030, although at a reduced rate after 2020. Management and reclamation on the heap will take place while other components of the Çöpler sulfide project continue to operate, with the active closure period starting after the end of deposition in the TSF.

17.4.11 Further Work

There may be an opportunity to utilise the heap drain-down solution in the sulfide circuit rather than disposing of it by forced evaporation, potentially reducing costs. This will require changes to the design of the evapotranspiration cells included in the current estimate.

Further studies and design work are required for the mitigation of PAG materials exposed in the pits to verify whether the proposed 1 m of non-PAG cover is practical and effective to implement.

The growth media inventory and expected amount to be recovered over the course of the Project should be compared to the sum of the growth media requirements of the Project facilities. Further work is required to determine the most sustainable revegetation covers to be employed.

17.5 Sustainability

Anagold aims to provide sustainability governance that not only meet or exceed the requirements of Turkish legislation, but also align with the expectations of ICMM (International Council of Mining & Metals) guidance and International Finance Corporation (IFC) Performance Standards, and the World Gold Council. The Anagold approach to policy development is to identify the most stringent standards and integrate them into project policy.

Çöpler project policies are supplemented by site-specific environmental and safety standards, management plans and procedures that are specifically tailored to the unique environmental and social challenges and permitting regulations of the site. These plans are certified to the requirements of international standards including ISO14001: 2015 and ISO45001.

Anagold maintains annual sustainability reporting for the Project, the report is produced to be in accordance with GRI Standards. The 2020 Sustainability Report has been completed and is publicly available. The 2021 Sustainability Report is currently under development.

Anagold has a dedicated Environmental, Health, Safety and Sustainability (EHS&S) Committee. The EHS&S Committee oversees, monitors, and reviews practice and performance in areas of safety, health, stakeholder relationships, environmental management, and other sustainability issues.

Sustainability is also a key responsibility for group level executives and site teams. The approach to sustainability is underpinned by the principle of collective responsibility and a belief that every employee must contribute to our sustainability performance – particularly on issues of health and safety and reporting of incidents.

17.5.1 Stakeholder Engagement

At the Çöpler project, Anagold has a wide-ranging stakeholder engagement programme which sets out the ways in which Anagold engages with stakeholders and ensures regular communication with stakeholder groups.

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During 2021, stakeholder consultations included meetings with shareholders, analysts, local communities, local and national authorities, contractors, government representatives, NGOs, universities, political parties, and trade union officials. Some of the key topics discussed included the Mine Expansion Project, Social Development Fund, exploration activities, cyanide and environmental awareness, local procurement, local contracting opportunities, training, and job creation.

The grievance mechanism is an important part of the Anagold local stakeholder engagement programme and the overall governance of sustainability. The community grievance mechanism has been developed to meet the requirements of both Turkish regulations and the IFC Performance Standards. The mechanism is designed to be widely accessible and there are access points available throughout each of the affected communities. There is also a dedicated access point for suppliers.

17.5.2 Health and Safety

Health and Safety Policy is guided by two key goals. First, to eliminate fatalities and serious injuries from our operations, and second, to continually reduce the number of minor injuries occurring on site. To fulfill these goals on the ground we implement:

•Robust systems and plans

•Risk assessment and controls

•Employee engagement

•Training

Anagold measures safety performance by tracking a range of leading and lagging safety indicators, the safety statistics reported also include exploration activities. All significant incidents are investigated and, based on findings, corrective action plans are developed to prevent recurrence.

17.5.3 Training and Development

The approach to the development of people is to strategically and continuously invest in staff training to ensure the business and operational needs both now and in the future are met. The development opportunities provided include technical skill development, leadership and business literacy skills, procedures and standards, and career development for staff. Çöpler has a specialised training centre with a capacity of 150 trainees.

Anagold carries out training and capability development programmes for our neighbouring community. Training is directed to future roles with the Project, while other training is focused on general skills development to enable people to seek gainful employment in other industries and locations throughout Turkey. This will help to broaden the economy and skills base in the Iliç District.

17.5.4 Industrial Relations

The workforce has no restrictions on union representation. Approximately 60% of the workforce at the Çöpler project are union members and have collective agreements in place. There have been no instances of industrial action.

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17.5.5 Diversity and Inclusion

Anagold does not set diversity or gender quotas for the workforce. Personnel are appointed based on merit and have specific objectives in place to ensure that the candidate pools for any position available throughout the company are made up of a range of qualified and diverse candidates. Women are paid equal with men in similar positions. The Anagold Diversity Policy commits the Project to provide:

•An environment in which all employees are treated with fairness and respect; and

•Equal access to opportunities, regardless of race, gender, sexual orientation and/or religious beliefs.

The approach to recruitment is to first look to local communities with appropriate skills. If unsuccessful, this is followed by recruiting from the wider region, followed by nationally, before finally looking internationally. The Anagold commitment to employing and developing local and national workers is reflected by the targets set for the Çöpler project:

•90% of unskilled workers to be drawn from communities impacted and affected by Anagold operations.

•80% of semi-skilled worker to be drawn from impacted and affected communities.

•80% of skilled workers to be Turkish citizens.

Suppliers are also encouraged to employ local workers whenever possible.

Local supply chains are preferred. Where supplier skills are lacking Anagold work with the suppliers to build capacity by providing training and mentoring.

17.5.6 Sustainable Community Development

To promote economic development in the communities neighbouring the Çöpler mine a Social Development Fund (SDF) was established in 2018. The SDF provides a structure under which Anagold will work in partnership with communities neighbouring the Çöpler mine, applicable Government agencies, third-party development partners and other relevant stakeholders, with the objectives of:

•Ensuring Anagold’s SDF funding of community programmes and projects is managed and distributed in a fair, transparent, and equitable manner.

•Building capacity within the local communities to participate in the benefits afforded by the mine and related regional economic and social development more actively.

•Moving away from donations type community relations expenditure by developing sustainable projects and programmes which address agreed social and community development priorities in the areas of agriculture, health, education, non-mine related income generation, and empowerment of underrepresented and disadvantaged groups.

•Where appropriate, reviving and promoting traditional customs and practices.

•Promoting independence from Anagold operations and assisting the communities to prepare for life beyond mining.

•Where appropriate, community relations expenditure by developing sustainable projects and programmes which address agreed social and community development priorities and/or benefit of public such as infrastructure, renovation, sponsorship, and construction.

Anagold will work with the community and other development partners in a manner that reflects the core values and principles of the SDF which include:

•Fairness and Equality – Impartial administration of the SDF, with all sectors of the SDF communities treated equally.

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•Transparency – Clear, publicly available processes for how the SDF is managed, and timely and fulsome reporting of decisions that are made, including financial reporting everyone has access to the same information.

•Cooperation and Partnership – Anagold working with the Community to focus on agreed development priorities. The SDF will not initiate programmes that are not requested by the community and in which the community do not have active and meaningful participation.

•Mutual respect – Everyone has a right to be heard and their opinion considered.

•Sustainability – Focusing on what counts over the long term and preparing for life beyond mine closure.

•At all times being fully compliant with relevant Turkish and International laws and conventions, and Anagold corporate policies and commitments.

While recipients of the SDF expenditure are the communities neighbouring the Çöpler mine, Anagold will retain ownership and governance control over all aspects of Anagold’s financial and in-kind contributions to the SDF. Anagold’s contribution to the SDF includes direct financial support, managerial/administrative support, and limited technical support.

Direct financial support has been approved by Anagold’s partners (SSR and Lidya) for ongoing annual funding to the SDF of $2 per ounce of gold produced from the Çöpler orebody. The SDF will replace a substantial proportion of Anagold’s existing discretionary community expenditure and direct funding towards development proprieties which are agreed with the community. The continuation of Anagold’s support to the SDF is at Anagold’s discretion, and will be influenced by, among other things, the success of the SDF and the community’s participation in ensuring the objectives of the SDF are achieved.

Managerial and administration support will be provided to the recipients of the SDF and Anagold’s policies, procedures, and management plans. Anagold will also cover the costs associated with stakeholder communication and consultation during the roll-out of the SDF, including support for the first three years in establishing a helpdesk facility for SDF applicants to receive assistance in preparing their applications.

While support to the SDF applicants on how to apply and administer their applications and projects will be available through a dedicated SDF helpdesk, where appropriate, and where relevant skills exist within the Company (and timing permits), Anagold will also support the SDF applicants with limited ad-hoc technical support as projects are being developed, and during the implementation phase. However, where a project requires specific and ongoing technical support, project applicants must ensure this is identified and resourced appropriately using third-party technical resources.

Anagold’s intensions for the SDF initiative are based on goodwill and respect for its neighbouring communities, however, Anagold acknowledges that other individuals, organisations, and government agencies may be more skilled and adept at identifying and implementing social and community development programmes and projects. As such it is Anagold’s desire that the SDF be implemented in such a way that third parties are attracted to participate in supporting community based SDF initiatives. In this way, the SDF can realise a greater funding base as well as attract leading skills in social and community development programme implementation. Third-party partners can include organisations providing development support or financial support including Government agencies, NGOs, or other credible development organisations. The SDF will not be used to fund third-party projects outside the approved SDF catchment area. Where a third-party partnership is part of an SDF application, the working relationships between Anagold, project applicants, and third-party partners must be clearly detailed in the Project application. Details of these relationships will form part of the application review process and be thoroughly scrutinised with respect to Anagold’s FCPA policy.

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While Anagold’s annual contribution to the SDF is substantial, not every project will receive funding. The SDF will be established to focus on participatory needs-based development priorities which support the abovementioned purpose. It is proposed that development priorities will be re-assessed every three-years.

17.5.7 Environmental Management

Anagold’s commitment to responsible environmental management is set out in the Environmental Policy, which complies with in-country legislation, the IFC Performance Standards, and the Equator Principles. The Çöpler Environmental Management System (EMS) is certified to the international ISO 14001: 2015 standard. The latest ISO 14001: 2015 external audit was completed successfully in December 2019.

17.5.8 Water Risk

The Çöpler project is in a high desert region in Eastern Turkey near the culturally significant Euphrates River. All water used at Çöpler is governed by strict permitting rules regarding abstraction and discharge under Turkish regulations. The approach to water management is to use water as efficiently as possible and to only draw as much needed and allowed within permitted limits. All water abstract is groundwater. Water used on site is recycled and re-used in the process plant. Water is not discharged to the environment.

17.5.9 Energy and Climate Change

All the electricity the Çöpler project uses is drawn from the Turkish national grid. Approximately 41% of Turkey’s national grid capacity comes from hydropower stations. The treatment of sulfide ore requires a more energy and CO2 intensive process than the oxide ore process that was previously the only ore treated at the Çöpler project. Anagold plan to use 2019, 269 GWh, as the baseline year for electricity use and efficiency, and to set targets based on 2019. The greenhouse gas emissions are published in the Anagold sustainability report.

17.5.10 Tailings Dam Management

Tailings produced by the Çöpler project are classified as class II non-hazardous. All tailings are sent to a carefully engineered TSF. Anagold has procedures in place to ensure that all parts of the TSF life cycle from construction to closure align with international best practice standards.

The TSF at the Çöpler project is a downstream mass filled dam. It became fully operational during the final quarter of 2018 with the start-up of the sulfide plant. The technical specifications for the construction of the Çöpler project TSF conforms with both Turkish national requirements and accepted good practice standards for tailings facilities, including:

•World Bank Standards

•Canadian Dam Association Safety Guidelines

•ICOLD (International Commission on Large Dams) Bulletins

•Turkish Hydraulic Works’ Technical Codes

•Mining Association of Canada (MAC) Guide to the Management of Tailings Facilities.

The Çöpler project TSF has been designed to withstand significant earthquakes up to a magnitude of 7.5 on the Richter scale. Modelling showed that even in the most severe seismic event, the wall of the TSF will heave with minimal risk of altering facility location or strength. There are no communities living directly downstream of the Çöpler project TSF.

The TSF uses a combination of technology, regular inspections and external oversight and audits to monitor the Çöpler project TSF (see Section 18.10.3).

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In addition to stability designs and monitoring, Anagold also has three groundwater monitoring wells in place both above and below the Çöpler project TSF, to monitor for signs of groundwater contamination. It was designed to meet the best in class requirements for class I (hazardous) waste, even though all tailings are classified as class-II (non-hazardous).

17.5.11 Water Management

The process of removing ore from the ground and extracting gold creates significant non-hazardous and some hazardous waste, which must be appropriately dealt with over the long- and short-term. Ensuring all waste is responsibly dealt with is crucial to protecting the health of the local environment and neighbouring communities.

To ensure that all waste, whether hazardous or non-hazardous, is reduced and dealt with in a safe and responsible manner, the Çöpler project has a detailed and comprehensive waste management plan. This is underpinned by the goal to reduce the amount of waste generated and to maximise the proportion of waste sent for recycling.

The bulk of the waste created at the Çöpler project is waste rock. All the waste rock created by the Çöpler project is carefully disposed of in engineered waste rock dumps. The design and management of all waste rock dumps is overseen by geotechnical engineers to ensure they have safe slope angles, maximum structural stability and management of any potentially acid forming materials are conducted appropriately by mine operations and thus meet the requirements of Turkish national regulations, industrial best practices and the IFC Performance Standards.

17.5.12 Cyanide Management

The use of cyanide is a critical part of the gold mining process. However, if not handled correctly, cyanide can have significant impacts on both environmental and human health. The use of cyanide at the Çöpler project is governed both by the requirements of Turkish national laws and regulations and aligned with industrial best practice. All employees and contractors who handle, transport, or dispose of cyanide are required to undertake specialised training in cyanide handling.

17.5.13 Biodiversity

The size, scale and location of mining operations means they can have a negative impact on local biodiversity. Failure to manage these risks and minimise the impacts on biodiversity could affect the social licence to operate and reputation. The Anagold aim is to restore sites (both operational and exploratory) and repair any damage done to the extent practicable. To do this, detailed records of the full range of biodiversity present as part of feasibility studies of any project or expansion. These studies form the basis for a Biodiversity Action Plan (BAP). The BAP sets out how impacted ecosystems are to be restored to their original state (or as close as possible) at the time of closure. Both the Çöpler project, its associated TSF and prospects have Biodiversity Action Plans in place. Anagold also conducts biodiversity monitoring studies each quarter with experts from Gazi and Hacettepe Universities.

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17.5.14 Air Quality

There is a potential for dust to be generated across many parts of the operation, including blasting, crushing, and milling, and the movement of large vehicles on haul roads. Dust management is a key focus across all facets of the operation. Air quality and the presence of dust is an important factor for local communities and workers. Ensuring management air quality for workers and communities is an important part of environmental management. Anagold has put in place a dust management plan at the Çöpler project to minimise the levels of dust in the air and ensure they fall within Turkish and IFC guideline limits. There are several monitoring stations across site and in the local communities. These stations record levels of airborne particulate matter and dust fall out. The results from the monitoring stations are reported to the relevant national authorities, and to local communities.

17.6 QP Opinion

Legal matters such as statutory and regulatory interpretations affecting the mine plan and environmental matters are outside the expertise of the QP and are within the control of the registrant (see Section 25).

Following a review of the information supplied, the opinion of the QPs is that it is reasonable to rely on the information provided by SSR as outlined above for use in the CDMP21TRS because a significant environmental and social analysis has been conducted for the project over an extended period, the project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

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18 CAPITAL AND OPERATING COSTS

Capital and operating cost estimates have been developed based on the current project costs, the mine and process designs, and discussions with potential suppliers and contractors. The estimated capital costs are to a feasibility level of accuracy and include a contingency of 10%.

18.1 Capital Costs

The following section describes the costs from the 2021 Çöpler District Master Plan, which was used as the starting point for the Initial Assessment.

As the Project has been in operation for a number of years, the level of project definition for the capital cost estimate is very high. The QP considers the capital estimate to be in the accuracy range of +/-15%.

Growth capital costs in the CDMP21TRS Reserve Case includes costs for:

•Ardich establishment and mine development

•Heap leach phases 5 and 6

•Road relocation, studies, and project management

•Explosives magazine relocation

Sustaining capital in the CDMP21TRS Reserve Case includes costs for:

•Tailings storage facility (TSF)

•Project team

•Technical services

•Administration

•Assay laboratory

•Mining

•IT

•Sulfide processing

•Oxide processing

•Environment

•Mineral / lands rights

•Health and safety

•Security

•Supply chain

•Reclamation

Capital costs assumptions to the end of 2021 and for the life-of-mine (LOM) are shown in Table 18.1.

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Table 18.1 Capital and Reclamation Costs

Description	Unit	Total LOM
Oxide
Growth	$M	69
Sustaining	$M	29
Sulfide
Growth	$M	–
Sustaining	$M	413
Reclamation and Other
Reclamation	$M	114
Working and Other	$M	–37
Total	$M	588

18.2 Operating Costs

Operating costs were estimated based on current site cost performance and contract costs including actual operational costs for labour, consumables, contracts and the Anagold budget assumptions. As the Project has been in operation for a number of years, the level of project definition for the operating cost estimates is very high. Given the available project performance data and the high project definition, no contingency was included in the cost estimate. The QP considers the operating cost estimate to be in the accuracy range of +/-15%. The projected LOM unit operating cost estimate is summarised in Table 18.2 and the average costs are shown in Table 18.3.

Table 18.2 Average Operating Costs Unit Rates

Activity	Unit	LOM Average Unit Cost
Mining	$/t mined	1.62
Processing – Heap Leach	$/t HL processed	14.45
Processing – Sulfide	$/t sulfide processed	35.91
Site Support and Office	$/t ore processed	5.21

Table 18.3 Summary of LOM Average Operating Costs

Cost	Total LOM<br>($M)	5-Year Average<br>per year<br>($/t ore)	LOM Average<br>per year<br>($/t ore)
Mining	766	14.98	10.15
Process	2,225	27.79	29.49
Site Support and G&A	473	7.14	6.27
Operating Costs	3,346	49.91	45.91

Mining costs include waste stripping costs

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18.3 Mining Cost Summary

Mining operating costs include:

•Drill and blast

•Load and haul

•Labour

•Dewatering

•Other indirects

Mining capital costs include:

•Fixed equipment

•Mobile equipment

•Office and supply

•Mine rehabilitation

•Studies

18.4 Processing and Infrastructure Cost Summary

The following has been included in the costs for processing:

•Oxide processing

•Sulfide processing

•Waste management

•TSF

•Utilities and services

•Reagents

•Plant infrastructure

•Plant mobile equipment

The following has been included in the capital costs for infrastructure cost estimates:

•Bulk services

•Site preparation

•Buildings and structures (new and refurbished)

•Communications

•IT hardware and software

•Security and access control

•Site costs

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•Mobile equipment

•Services contracts

•Community support

The following has been included in the operating cost estimates:

•Plant consumables

•Crusher consumables

•Screens

•Grinding media

•Filters

•Packaging plant bags

•Plant reagents

•Plant mobile equipment

•Plant maintenance

•Power

•Labour

•Production and dispatch

•Plant and infrastructure day work services

•Plant technical services

•Shift maintenance

•Laboratory service level agreement

•TSF water treatment

18.5 General and Administration Cost Summary

•Office and general expenses

•Site support costs

•Off-site Anagold offices

•Internal and external consultants

•Maintenance and inspection contracts

•Equipment and sundry

•Fuels and utilities

•Rentals and leases

•Insurance and insurance taxes

•IT hardware and software

•Personnel transport

•Communications

•Licences and land fees

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•Labour

•Accommodation and messing

•Medical support

•Flights

•Light vehicles

•Environmental, community development and engagement

•Banking and audit fees

•Legal

Table 18.4 Other Capital Costs

Item	Units	Amount
Closure	$M	114
Heap Leach Sustaining	$/t	0.15

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19 ECONOMIC ANALYSIS

19.1 Economic Assumptions

The economic analysis was prepared on a 100% project basis using the Reserve Case production schedule, operating, and capital assumptions on an annual basis. The assumptions for taxes and royalties were provided by SSR.

19.1.1 Metal Prices

Metal prices for the economic analysis were estimated after analysis of consensus industry metal price forecasts and compared to those used in other published studies. The metal prices used for the economic analysis, shown in Table 19.1, are considered to be representative of industry forecasts.

Table 19.1 CDMP21TRS Economic Analysis Metal Price Assumptions

Metal Price	Units	2022	2023	2024	2025	Long-Term
Gold	$/oz	1,800	1,740	1,710	1,670	1,600
Silver	$/oz	24.00	23.00	22.00	21.00	21.00
Copper	$/lb	4.00	3.80	3.80	3.80	3.40

19.1.2 Taxation

The Turkish government implemented a temporary rate increase from 20% to 22% for the periods of 2018–2020. From 2022 onwards, the effective tax rate is expected to return to 20%.

Investment incentive certificates (IIC) are available for investments that promote economic development. IIC’s can be classified as strategic in specific circumstances, thereby providing additional incentives. Anagold received a strategic IIC for the sulfide process plant. An IIC generates credits that offset corporate income taxes generated by the investment. The amount of investment credits generated from the IIC is based on eligible capital expenditures. These investment credits reduce the corporate tax rate to a minimum of 2% in a given tax period until the last quarter of 2023, thereafter it is assumed subsequent non-strategic IIC’s will be available, and the minimum rate will be 4%. Incentive tax credits can be carried forward to future tax periods indefinitely until exhaustion. Incentive tax credits and other tax pools are determined in the local currency, Turkish Lira, and subject to devaluation and revaluation as fluctuations against the US dollar occur. The cash flow model is prepared on a constant Turkish Lira basis.

Import duties are not included in the capital cost estimate for mining related imported equipment because they are exempted in the IIC’s.

19.1.3 Royalties

Under Turkish Mining Law, the royalty rate for precious metals is variable and tied to metal prices. The Çöpler project is subject to a mineral production royalty which is based on a sliding scale to gold price and is payable to the Turkish government.

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Table 19.2 details the current prescribed royalty rates applicable to heap leach production (revised September 2020). The royalties are calculated on total revenue with deductions allowed for processing and haulage costs of ore. As the Çöpler project produces silver and copper as by-products of the process of treating gold ore, revenue from by-products is included in the total revenue used for royalty calculations. Royalty rates are reduced by 40% for ore processed in country, as an incentive to process ore locally. The Çöpler project POX production is eligible for a 40% reduction to the royalty rate.

Table 19.2 Gold Royalty Rates

Metal Price(/oz Gold)	Prescribed Royalty <br>Rate<br><br>(%)	Royalty After 40% In-Country Processing Incentive<br>(%)
From
0	1.25	0.50
800	2.50	1.00
900	3.75	1.50
1,000	5.00	2.00
1,100	6.25	2.50
1,200	7.50	3.00
1,300	8.75	3.50
1,400	10.00	4.00
1,500	11.25	4.50
1,600	12.50	5.00
1,700	13.75	5.50
1,800	15.00	6.00
1,900	16.25	6.50
2,000	17.50	7.00
2,100	18.75	7.50

All values are in US Dollars.

The Çöpler project effective life-of-mine (LOM) royalty rate based on the financial model metal price assumptions and applicable deductions is approximately 4.9%.

Other than the royalty payments, there are no other known back-in rights, payments, or other agreements and encumbrances to which the property is subject.

19.1.4 QP Opinion on Inputs

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19.2 Reserve Case Economic Analysis Results

The economic analysis was prepared on a 100% project basis using the Reserve Case production schedule, operating, and capital assumptions on an annual basis. The Reserve Case production includes 22.5 Mt at 1.69 g/t Au oxide ore processed by heap leaching and 52.9 Mt at 2.33 g/t Au processed in the sulfide plant on a 100% project basis. Total production is 75.4 Mt at 2.14 g/t Au. Total gold production is 4.4 Moz. Mining at the Çöpler pit is completed in 2029 and at Ardich in 2034. Oxide heap leach stacking is completed in in 2034, while sulfide processing will continue from stockpiles until 2042.

The Reserve Case results include:

•After-tax NPV at a 5% real discount rate is $1.73 billion

•Mine life of 21 years

The after-tax cash flow is shown in Figure 19.1. The NPV results for before and after-tax over a range of discount rates is shown in Table 19.4. The sulfide and oxide production profiles are shown in Table 19.4 and gold production in Figure 19.3. Cash costs are shown in Table 19.5.

Figure 19.1 CDMP21TRS Reserve Case After-Tax Cash Flow

OreWin, 2022

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Table 19.3 CDMP21TRS Reserve Case Results Summary

Item	Unit	Reserve Case
Oxide Processed
Heap Leach Quantity	kt	22,557
Au Feed Grade	g/t	1.69
Sulfide Processed
Quantity Milled	kt	52,892
Au Feed Grade	g/t	2.33
Total Processed
Processed	kt	75,448
Au Feed Grade	g/t	2.14
Total Gold Produced
Oxide – Gold	koz	765
Sulfide – Gold	koz	3,604
Total – Gold	koz	4,369
Oxide – Gold Recovery	%	61
Sulfide – Gold Recovery	%	91
5-Year Annual Average
Average Gold Produced	kozpa	278
Free Cash Flow	$Mpa	158
Total Cash Costs (CC)	$/oz gold	880
All-in Sustaining Costs (AISC)	$/oz gold	1,071
Key Financial Results
Life-of-Mine (LOM) CC	$/oz gold	803
LOM AISC	$/oz gold	966
Site Operating Costs	$/t treated	45.91
After-Tax NPV5%	$M	1,732
Mine Life	years	21

5-Year Annual Average is for the period 1 January 2021 through 31 December 2025

Table 19.4 CDMP21TRS Reserve Case Before and After-Tax NPV

Discount Rate	Before-Tax NPV<br>($M)	After-Tax NPV<br>($M)
Undiscounted	2,729	2,555
5%	1,824	1,732
10%	1,322	1,268
12%	1,185	1,140

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Figure 19.2 CDMP21TRS Reserve Case Processing

OreWin, 2022

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Figure 19.3 CDMP21TRS Reserve Case Gold Production and Recovery

OreWin, 2022

Table 19.5 CDMP21TRS Reserve Case Cash Costs

Description		Reserve Case
Mining and Rehandle		766
Process, Freight, and Refining		2,031
Site Support		393
Royalties		353
Total Production Costs		3,543
Total Cash Costs (CC)		803
Sustaining Capital		442
Fixed Lease Payments		192
Site G&A		81
All-in Sustaining Costs (AISC)		4,257
	/oz gold

All values are in US Dollars.

Process, Freight, and Refining includes by-product credits and excludes fixed lease costs.

Royalties are calculated in the period incurred and applied to cash flow in the subsequent year.

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A financial model was prepared on a 100% project basis using the Reserve Case production schedule and operating and capital assumptions on an annual basis. The assumptions for taxes and royalties were provided by SSR. The corporate tax rate in Turkey is 23% in 2022 but will revert to 20% from 2023. The royalty rate for precious metals under Turkish Mining Law is variable and tied to metal prices. As Çöpler ores are processed on-site, the applicable royalty rate for POX processing is subject to a further 40% reduction for certain qualifying operating costs. The average royalty calculated as a proportion of gross revenue in the Reserve Case is approximately 4.9%.

Metal prices were estimated after analysis of consensus industry metal price forecasts and metal prices used in other comparable studies. The prices used for the economic analysis are shown in Table 19.6.

Table 19.6 CDMP21TRS Reserve Case Metal Price Assumptions

Metal Price	Units	2022	2023	2024	2025	Long-Term
Gold	$/oz	1,800	1,740	1,710	1,670	1,600
Silver	$/oz	24.00	23.00	22.00	21.00	21.00
Copper	$/lb	4.00	3.80	3.80	3.80	3.40

The after-tax NPV sensitivity to metal price variation is shown in Table 19.7 for gold prices from $1,000–$2,000/oz. Cost sensitivity is shown in Table 19.8.

Table 19.7 CDMP21TRS Reserve Case Gold Price Sensitivity

After-Tax NPV ($M)	Long-Term Gold Price (/oz)
Discount Rate	1,000	1,350	1,400	1,600	1,750	1,800	2,000
Undiscounted	967	1,974	2,104	2,555	2,890	2,985	3,398
5%	769	1,370	1,447	1,732	1,939	1,998	2,252
10%	645	1,029	1,079	1,268	1,405	1,444	1,611
12%	608	935	977	1,140	1,257	1,291	1,435
15%	563	822	856	987	1,082	1,110	1,226
18%	525	735	762	870	948	970	1,066
20%	504	687	711	806	875	894	979

All values are in US Dollars.

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Table 19.8 CDMP21TRS Reserve Case Cost Sensitivity

Change from Base NPV5% (M)
Variable	Base Value	–20%	–10%	0%	10%	20%
Capital Cost	575	1,790	1,761	1,732	1,702	1,671
Mining Cost	1.62	1,833	1,782	1,732	1,681	1,630
Processing Cost	29.49	1,957	1,845	1,732	1,618	1,503
Site Operating Cost	108	1,785	1,758	1,732	1,705	1,679
Gold Royalty	81	1,769	1,751	1,732	1,713	1,694

All values are in US Dollars.

19.2.1 Project Cash Flow

The after-tax cash flow and average LOM AISC unit cost is shown in Table 19.5.

The annual revenue, operating cost and capital costs and net cash flow is tabulated in Table 19.9.

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Table 19.9 CDMP21TRS Reserve Case Project Cash Flow

Item
		2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	2035	2036	2037	2038	2039	2040	2041	2042	2043	2044+
	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M
Gross Revenue
Realisation Costs
Freight and Refining
Royalties
Total – Realisation Costs
Operating Costs
Mining
Processing – Heap Leach
Processing – Sulfide Plant
Site Support
G&A
Total – Operating Costs
Operating Surplus
Total – Capital Costs
Net Cash Flow Before Tax
Tax
Net Cash Flow After Tax

All values are in US Dollars.

Royalties are paid in the period after they are accrued

2044+ covers the period from 2044–2053

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20 ADJACENT PROPERTIES

There are no adjacent properties that are applicable to the CDMP21TRS.

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21 OTHER RELEVANT DATA AND INFORMATION

There is no other relevant data or information that is applicable to the CDMP21TRS.

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22 INTERPRETATION AND CONCLUSIONS

Following a review of the information supplied, the opinion of the QPs is that it is reasonable to rely on the information provided by SSR as outlined above for use in the CDMP21TRS because a significant environmental and social analysis has been conducted for the Project over an extended period, the Project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

Mineral Resources were originally discussed and disclosed in the Çöpler District Mineral Resource 2021 Technical Report Summary (2021MR). The 2021MR is an independent Technical Report Summary prepared to provide a preliminary technical and economic study of the economic potential of the Çöpler District mineralisation to support the disclosure of Mineral Resources. Some content from the 2021MR report has been used in the CDMP21TRS. Sections relating to the Mineral Resource have not been changed from the 2021MR, as that work remains the most current study work available.

Mineral Resources and Mineral Reserves in the CDMP21TRS are reported in accordance with subpart 1300 of US Regulation S-K Mining Property Disclosure Rules (S-K 1300).

The CDMP21TRS has identified additional Mineral Resources and additional Mineral Reserves when compared to prior studies.

The 2021MR Initial Assessment Case demonstrated that there is significant economic potential that may be derived from the copper in the Çöpler Mineral Resource. Given this economic potential it is then concluded that it is valid to report the Mineral Resources using the Mineral Resource metal prices and pit shell.

Further study and analysis will be required to advance the understanding of this potential.

The CDMP21TRS QPs have not identified any relevant technical and/or economic factors that require resolution with regards to the Mineral Resource estimates. Significant factors that could materially affect the Mineral Resources and Mineral Reserve are:

•Environmental, Permitting Social and Community – the Çöpler project is subject to the laws and regulations of Turkey, the mine has several local communities that are nearby. In order to operate the mine, Anagold must maintain appropriate relations with all the authorities and stakeholders. Social, community and government relations are managed by Anagold and include programmes and engagement with the local communities and both local and national governments. Anagold has remained in compliance with all aspects of the Environmental Impact Assessments (EIA) and operating permits throughout the history of the Project.

•Seismic impacts – the Çöpler project is in an area with a history of significant seismic activity that could negatively impact mining operations.

•Geotechnical impacts – slope recommendations have significant impacts on the Mineral Reserve and the continued study will allow the Mineral reserves to be maximised.

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23 RECOMMENDATIONS

Key recommendations from the CDMP21TRS are:

•Continue to update and evaluate the Çöpler District Master Plan as the existing Mineral Resources and Mineral Reserves are updated and as new prospects are advanced.

•Undertake infill drilling at Çöpler and update the copper Mineral Resource estimate.

•Prepare further studies of the copper recovery options.

•Geotechnical review and study of the re-evaluation of the pit re-designs.

•Optimisation of the sulfide flotation circuit, POX, and process operation.

•Metallurgical testwork on future oxide, sulfide, and copper ore sources.

•Optimisation of the oxide heap leach circuit.

•Optimisation of the mining rates to increase gold production.

•Stockpile reconciliation and management studies.

•Review and adapt the ore control and stockpiling strategies to optimise recovery and throughput and maximise gold production.

•Reconcile monthly blend and gold production with predictive modelling.

•Continue drilling at Ardich.

•Geotechnical studies of Ardich.

•Reconciliation studies of Çöpler.

•Update Çöpler and Ardich resource models and estimates.

•Further study of the 2021MR Initial Assessment to advance and next stage of study:

•Geotechnical studies

•Environmental Impact Assessments (EIA) and permitting

•Blasting studies

•Metallurgical studies.

23.1 Mineral Resources

Specific recommendations related to the Mineral Resource are:

•Mineral Resource models should be updated on a campaign basis following the completion of planned drilling programmes. Where significant new data has been obtained (either exploration data, or production data), an annual model update roster should be adequate, but only required where warranted by the introduction of new data that has potential to result in a material change in the model (such as by significant modifications to the geological interpretation, or by substantial expansion of the dimensions of the mineralisation).

•The Çöpler model has not been updated since 2016. It is recommended that a new model be developed to incorporate the new exploration data obtained since that time, and to check interpretations relative to grade control data to help hone the interpretation.

•Continue drilling at Ardich.

•An update to the Ardich model is warranted given the quantum of new data that has been obtained since the most-recent update, and the status of the deposit as shown in the Initial Assessment Case.

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•Both Çöpler and Ardich are geologically complex deposits with multiple metals that must be tracked along with oxidation type and lithological domains, further complicated by extensive structural disruption. Work on verifying and adjusting resource model domains and parameters should be continued to help facilitate a greater understanding of the deposits, hopefully resulting in improved resource estimates.

•Since the mineralisation locally follows the lithological contacts and structural features, using a search ellipse that follows these trends (dynamic anisotropy) should be evaluated.

•An audit of the databases used to house exploration and grade control data should be undertaken on a reasonably regular basis (e.g., annually). This should include review of all related procedures, monitoring observance to the procedures, and spot checks of the database itself to identify errors and omissions.

•A comprehensive and consistent suite of assays should be collected routinely in exploration drilling. This should be formalised as a requirement across all exploration drilling. Estimation into the resource models should involve all components that may be of future interest.

•The routine collection of in-pit mapping data is encouraged as this information provides invaluable experiential knowledge to inform interpretations based on exploration data.

•Detailed scheduling and design of the sulfide ore stockpiles should be completed. Results from ongoing metallurgical testwork will assist in determining the optimal stockpiling strategy and in reconciliation success.

•Further refinement of the modelled carbonate and sulfide sulfur (SS) grades in the resource model should be completed.

•Further mapping and definition of the local and regional fault structures, alteration types, and other domains should be completed to reduce or realise geotechnical risk in the areas where these structures intersect the pit.

23.2 Mineral Reserves

Specific recommendations related to the Mineral Reserve are:

•Re-design of Çöpler pits at updated metal prices.

•Geotechnical review and study of the re-evaluation of the pit re-designs.

•Optimisation of float circuit pressure oxidation (POX) and process operation including metallurgical testwork on Ardich and Çöpler.

•Review and monitor the stockpiling procedures and criteria to optimise the feed to the plant.

•Optimisation of the mining rates to increase gold production.

•Stockpile reconciliation and management studies.

•Geotechnical studies of Ardich.

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24 REFERENCES

Abzalov, M.Z. Quality Control of Assay Data: A Review of Procedures for Measuring and Monitoring Precision and Accuracy, Exploration and Mining Geology, 17(3–4): 1–14, 2008.

Altman, K, Liskowich, M, Mukhopadhyay, D K, Shoemaker, S J, Çöpler Sulfide Expansion Project Prefeasibility Study, 27 March 2011.

Altman, K, Bascombe, L, Benbow, R, Mach, L, Shoemaker, SJ, Çöpler Resource Update, Erzincan Province Turkey, 30 March 2012.

Altman, K, Bair, D, Bascombe, L, Benbow, R, Mach, L, Swanson, B, Çöpler Resource Update, Erzincan Province Turkey, 28 March 2013.

Amec Foster Wheeler Çöpler Sulfide Expansion Project Feasibility Update 2015.

Anatolia (2009) Çöpler Project, East Central Turkey Preliminary Mine Reclamation & Closure Plan, 2009, Anatolia Minerals Development, Limited.

Barr (2012), Pit Wall Stability Analysis, Çöpler Mine, August 2012

Bloom, L., Analytical Services and QA/QC, for Society of Exploration Geologists, April 2002. Project Documents.

Easton, C L, Pennstrom, W J, Malhotra, D, Moores, R C, Marek, J M, Çöpler Gold Project East Central Turkey Preliminary Assessment Sulfide Ore Processing, 4 February 2008.

Golder (2013a), Çöpler Mine Sulfide Expansion Project, Flood Management Plan, May 2013 Golder Associates.

Golder (2013b), Çöpler Mine Sulfide Expansion Project, Groundwater Modeling Report, September 2013 Golder Associates.

Golder (2013c) Çöpler Sulfide Project Tailings Storage Facility Siting Study, 17 December 2013, Golder Associates.

Golder (2014a), Çöpler Sulfide Project – Stability Evaluation of Planned Waste Dump Facilities, Technical Memorandum, 28 February 2014, Golder Associates.

Golder (2014b), Geotechnical Report, Sulfide Plant Facilities – Updated Report Çöpler Sulfide Project, 10 March 2014, Golder Associates.

Golder (2014c), Çöpler Mine – Pit Slope Design Review, April 2014, Golder Associates.

Golder (2014d) Çöpler Sulfide Project – Tailings Storage Facility Analysis and Design, 28 July 2014, Golder Associates.

Golder (2015a), Çöpler Sulfide Project – Tailings Storage Facility, Summary of Design and Expansion to 46.6Mt Capacity, Technical Memorandum, March 2015.

Golder (2015b), Çöpler Sulfide Expansion Project – Stability Evaluation of Planned Waste Dump Facilities, Technical Memorandum, 14 May 2015, Golder Associates.

Golder (2015c), Geotechnical Report, Sulfide Plant Facilities – Detailed Design Recommendations, Çöpler Sulfide Project, 8 October 2015, Golder Associates.

Golder (2016a), Tailings Storage Facility, Detailed Design Criteria, Revision 3, February 2016.

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Golder (2016b), Çöpler Sulfide Project – Tailings Storage Facility Summary of Design and Expansion to 45.9Mt Capacity, Technical Memorandum, April 2016.

Golder (2021), Data Review and Geotechnical Model, Çöpler Geotechnical Design Review 2021, October 2021.

Golder (2021), Çöpler Pit Slope Design Review, November 2021.

Golder (2021), 2021 Ardich Project Slope Stability Study, Geotechnical Support for the Pre Feasibility Study, December 2021.

Golder (2021), Pit Slope Stability Evaluation, Çöpler Open Pit Mine, November 2019.

Hacettepe University, Gazi University (Hacettepe and Gazi Universities, 2014 (Interim), İliç (Erzincan) Çöpler Complex Mine Capacity Increase Project – Report on Biological Diversity, 2014 (Interim).

Independent Mining Consultants, Inc., Çöpler Project Resource Estimate Technical Report, 19 October 2005.

International Union for Conservation of Nature (IUCN), 1994: Red List: http://www.iucnredlist.org/static/categories_criteria_3_1.

Jacobs (2012) Çöpler Sulfide Project Feasibility Study, Site Conditions, 30 May 2012, Jacobs.

Jacobs (2014a), Çöpler Sulfide Expansion Project Definitive Feasibility Report, 15 June 2014, Jacobs.

Jacobs (2014b), Crushing and Grinding Systems for Handling Clayey Ore Trade-Off Study, 21 January 2014, Jacobs.

Marek, J M, Pennstrom, W J, Reynolds, T, Technical Report Çöpler Gold Project Feasibility Study, 30 May 2006 (Samuel Engineering, Inc.).

Marek, J M, Moores, R C, Pennstrom, W J, Reynolds, T, Technical Report Çöpler Gold Project, 2 March 2007 as amended 30 April 2007 (Independent Mining Consultants, Inc.).

Marek, J M, Benbow, R D, Pennstrom, W J, Technical Report Çöpler Gold Project East Central Turkey, 5 December 2008 (Amended and Restated; supersedes 11.07.2008 version).

Marsden, J. O., Çöpler Project – Heap Leach Model Review, 24 October 2014, Metallurgium, Phoenix, AZ.

Marsden, J. O., Çöpler Heap Leach Project Gold Recovery Assumptions – Rev 1, 26 March 2015, Metallurgium, Phoenix, AZ.

Marsden, J. O., Çöpler Project Heap Leach Model Development and Gold Recovery Assessment – Final Rev 2, 27 March 2015, Metallurgium, Phoenix, AZ.

OreWin Pty. Ltd., 2020. Çöpler District Master Plan 2020, 27 November 2020. (CDMP20TR)

OreWin Pty. Ltd., 2021. Çöpler District Mineral Resource 2021 Technical Report Summary 29 September 2022. (2021MR)

Outotec (2015a), Thickening Test Report S1482TE Çöpler, 16 September 2015, Perth, Australia.

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Outotec (2015b), Thickening Test Report S1482TE_B Çöpler, 16 September 2015, Perth, Australia.

Outotec (2015c), Thickening Test Report S1482TF Çöpler Appendix, 16 September 2015, Perth, Australia.

Outotec (2015d), Thickening Test Report S1482TF Çöpler (repeats), 20 November 2015, Perth, Australia.

Outotec (2015e), Thickening Test Report S1482TE_B Çöpler (Repeats), 20 November 2015, Perth Australia.

Parrish, I.S. 1997. Geologist's Gordian knot: to cut or not to cut. Mining Engineering, vol. 49. pp 45–49.

Pyper, R., Description of Process Gold Production Model and Assumptions, 4 February 2015, Kappes, Cassiday & Associates, Australia Pty Ltd., Perth, Western Australia.

Samuel Engineering (2011) Çöpler Sulfide Expansion Project Prefeasibility Study, 27 March 2011, Samuel Engineering.

SGS Lakefield Oretest (2015), Anagold Çöpler Sulfide Pilot Plant and Batch Testing Program, Pressure Oxidation and Cyanidation Campaign 5 Main Report, Job No: CP100, 30 October 2015, Perth, Australia.

SRK (2008) Çöpler Complex (Manganese, Gold, Silver, Copper) Mining Project EIA Report, 2008, SRK Consulting.

SRK (2012a) Assessment of Çöpler Sulfide Tailings According to Waste Acceptance Criteria, August 17, 2012 (Memorandum) SRK Consulting.

SRK (2012b), Çöpler Mine Sulfide Expansion Feasibility Study – Environment and Permitting, November 2012, SRK Consulting (Turkey).

SRK (2012c), Çöpler Gold Mine-Sulfide Project Waste Geochemical Assessment, September 2012, SRK Consulting (Turkey).

SRK (2014), Çöpler Complex Mine Capacity Expansion Project, Final EIA Report. October, 2014, SRK Consulting (Turkey).

SSR (2020). Announcement: SSR Mining Announces Exploration Results on the In-pit Copper-Gold Porphyry C2 Target at Çöpler, 25 November 2002.

Watts, Griffis and McQuat Limited, Update of the Geology and Mineral Resources of the Çöpler Prospect, 1 May 2003.

Yetkin, E., various dates. Multiple independent QA/QC reports referred to specifically throughout Section 9.

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25 RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT

The CDMP21TRS QPs have relied on the following information provided by SSR in preparing the findings and conclusions in this Technical Report Summary regarding the following aspects of modifying factors:

•Macroeconomic trends, taxes, royalties, data, and assumptions, and interest rates.

•This has been used in Section 19 as described in this section. The QPs have relied exclusively on SSR for this information.

•Marketing information and plans within the control of the registrant.

•This has been used in Sections 16 and 19 as described in those sections. The QPs have relied exclusively on SSR for this information.

•Legal matters outside the expertise of the qualified person, such as statutory and regulatory interpretations affecting the mine plan.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

•Environmental matters outside the expertise of the qualified person.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

•Accommodations the registrant commits or plans to provide to local individuals or groups in connection with its mine plans.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

•Governmental factors outside the expertise of the qualified person.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

Following a review of the information supplied, the opinion of the QPs is, that it is reasonable to rely on the information provided by SSR as outlined above for use in the CDMP21TRS, because a significant environmental and social analysis has been conducted for the Project over an extended period, the Project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

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Document

Exhibit 96.2

Explanatory Note

SSR Mining Inc. (the “Company”) previously filed the Marigold 2021 Technical Report Summary (the “Marigold21TRS”), with an effective date of December 31, 2021 and issued on February 23, 2022, as Exhibit 96.2 to its Annual Report on Form 10-K for the year ended December 31, 2021, as amended. The Marigold21TRS has been amended to reflect certain revisions in compliance with Subpart 1300 of Regulation S-K, which revisions consist of adding confirmatory statements and other modifications that SSR does not consider material. The amended Marigold21TRS has been reissued as of September 29, 2022 and is presented with an effective date of December 31, 2021. The information in this amended Marigold21TRS has not been updated to reflect events, information or developments occurring after the effective date.

This page does not constitute a part of the amended Marigold21TRS.

Title Page

Project Name:	Marigold
Title:	Marigold 2021 Technical Report Summary
Location:	Humboldt County, Nevada, USA
Effective Date of Technical Report Summary:		31 December 2021
Effective Date of Mineral Resources:		31 December 2021
Effective Date of Mineral Reserves:		31 December 2021

Qualified Persons (QPs):

•Bernard Peters, BEng (Mining), FAusIMM (201743), employed by OreWin Pty Ltd as Technical Director - Mining, was responsible for the overall preparation of the Marigold21TRS and, the Mineral Reserve estimates, Sections 1 to 5; Section 10; Sections 12 to 25.

OreWin Pty Ltd ACN 165 722 574

140 South Terrace Adelaide 5000

P +61 8 8210 5600 E orewin@orewin.com W orewin.comi

Signature Page

Project Name:	Marigold
Title:	Marigold 2021 Technical Report Summary
Location:	Humboldt County, Nevada, USA
Date of Signing:	29 September 2022
Effective Date of Technical Report Summary:		31 December 2021

/s/ Sharron Sylvester

Sharron Sylvester, Director, OreWin Pty Ltd

/s/ Bernard Peters

Bernard Peters, Technical Director – Mining, OreWin Pty Ltd, BEng (Mining), FAusIMM (201743)

/s/ Sharron Sylvester

Sharron Sylvester, Technical Director – Geology, OreWin Pty Ltd, BSc (Geol), RPGeo AIG (10125)

21013Marigold21TRS220927Rev0.docx ii

TABLE OF CONTENTS

1EXECUTIVE SUMMARY	1
1.1Introduction	1
1.2Property Description and Location	1
1.3Land Tenure and Ownership	1
1.4Geology andMineralization	2
1.4.1Regional Geology	2
1.4.2Local and Property Geology	2
1.4.3Property Structure	3
1.4.4Mineralization	3
1.4.5Alteration	4
1.5Exploration	4
1.5.1Exploration – Marigold	4
1.5.2Exploration – Trenton Canyon and Buffalo Valley	5
1.6Mineral Resources	5
1.7Mineral Reserves	6
1.8Mining Operations	7
1.9Mineral Processing	8
1.10Infrastructure	9
1.11Environmental, Permitting and Social Responsibility	10
1.12Market Considerations	10
1.13Capital and Operating Cost Estimates	11
1.15Economic Analysis	12
1.15.1Life-of-Mine Production	14
1.15.2Cost Statistics	14
1.16Interpretation and Conclusions	15

21013Marigold21TRS220927Rev0.docx iii

1.17Recommendations	15
1.17.1Processing	16
1.17.2Metallurgy / Analytical	16
1.17.3Mineral Resources	16
1.17.4Mine Planning	16
1.17.5Exploration Drilling	17
1.17.6Mine Operations	17
1.17.7Maintenance Operations	17
2INTRODUCTION	18
2.1Terms of Reference	18
2.2Qualified Persons	19
2.3Qualified Persons Property Inspection	19
2.4Units and Currency	19
2.5Abbreviations and Acronyms	19
2.6Effective Dates	21
3PROPERTY DESCRIPTION	22
3.1Property Location	22
3.2Corporate Structure	24
3.3Land Tenure and Ownership	25
3.3.1Owned Real Property	25
3.3.2Owned Unpatented Mining Claims	26
3.3.3Leasehold Rights	27
3.4Royalties and Encumbrances	31
3.5Environmental Liabilities	32
3.6Operating Permits	32
3.7Permits, Mineral, and Surface Rights	32

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3.8Other Significant Factors and Risks	32
4ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	33
4.1Access	33
4.2Climate and Physiography	33
4.3Infrastructure	33
5HISTORY	35
5.1Historical Exploration Work	35
5.2Historical Production Work	38
6GEOLOGICAL SETTING,MINERALIZATION, AND DEPOSIT	42
6.1Regional Geological Setting	42
6.2Local Geology	46
6.2.1Stratigraphy	47
6.2.2Igneous Rocks	48
6.2.3Regional Structure	49
6.3Property Geology	49
6.3.1Property Stratigraphy	49
6.3.2Property Structure	52
6.3.3mineralization	55
6.3.4Alteration	56
6.4Deposit Geology	59
6.4.1Mackay Pit	59
6.5Deposit Type	61
7EXPLORATION	64
7.1Gravity Surveys	64
7.1.1Gravity Survey Pre-2015	64

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7.1.2Gravity Survey Post-2015	64
7.1.3Gravity Stations	64
7.1.4Terrain Corrections	65
7.1.5Interpretation	66
7.2Exploration Drilling	67
7.2.1Exploration Drilling at Marigold (Pre-2014)	67
7.2.2Exploration Drilling at Marigold (2014–2017)	68
7.2.3Exploration Drilling at Marigold (2018–2021)	68
7.2.4Marigold Sulfide Drilling Program	68
7.2.5Exploration Drilling at Valmy (1968–2006)	69
7.2.6Exploration – Trenton Canyon and Buffalo Valley	69
8SAMPLE PREPARATION, ANALYSES, AND SECURITY	75
8.1Sample Preparation and Analysis	75
8.2Sample Security	78
8.2.1Sample Security until 2013	78
8.2.2Sample Security Valmy Property	78
8.2.3Sample Security 2014–2021	78
8.3Quality Assurance/Quality Control (QA/QC) Procedures	78
8.3.1QA/QC Procedures Pre-2014	78
8.3.2QA/QC Procedures Valmy Property	80
8.3.3QA/QC Procedures 2014–2017	81
8.3.4QA/QC Procedures 2018–2021	84
8.4Conclusions and Recommendations	87
9DATA VERIFICATION	88
9.1Marigold Assay Database	89
9.2QP Opinion	90

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10MINERAL PROCESSING AND METALLURGICAL TESTING	91
10.1Metallurgical Testwork	91
10.2Process Optimization Metallurgical Testwork	92
10.3Gold Recovery Modelling	93
11MINERAL RESOURCE ESTIMATES	95
11.1Introduction	95
11.2Drillhole Database	95
11.3Domains	95
11.4Geological Interpretation	98
11.5Exploratory Data Analysis	99
11.5.1Outlier Restriction	100
11.6Material Density	101
11.7Variograms	102
11.8Cell Model and Grade Estimation	104
11.9Model Validation	107
11.10Resource Classification	111
11.10.1Ore Reconciliation	112
11.11Mineral Resource Statement	112
11.12Comparison with Previous Estimates	113
11.13QP Opinion	114
11.14Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects	114
12MINERAL RESERVE ESTIMATES	115
12.1Mineral Reserves Estimate	115
12.2Cut-off Grade	116
12.3Royalties, Net Proceeds and Excise Tax	117

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12.4Dilution	117
12.5Mining Recovery	117
12.6Comparison with Previous Estimates	118
13MINING METHODS	119
13.1Geotechnical, Hydrological, Pit, and Other Design Parameters	119
13.1.1Open Pit Geotechnical Reports Review	120
13.1.2Pit Optimizations and Designs	121
13.2Pit Phases and Timing	122
13.3Production Rates, Mine Life, Dimensions and Dilution Factors	122
13.4Stripping Requirements	125
13.5Required Mining Fleet and Machinery	126
13.6Ore Control Drilling and Method	127
13.7Drilling and Blasting	128
13.8Loading Operations	128
13.9Hauling Operations	128
13.10Mine Support	129
13.11Mine Maintenance	129
13.12Mine General and Administration	129
13.13Mine Safety	129
13.14Mine Dewatering	141
14PROCESSING AND RECOVERY METHODS	145
14.1Introduction	145
14.1.1Ore Stacking on Leach Pad	145
14.1.2Leaching Solution to the Pad	145
14.1.3Pregnant Solution	146
14.1.4Carbon Processing	147

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14.1.5Refining	147
14.1.6Ventilation	147
14.1.7Planned Processing Upgrade Projects	147
14.1.8Reagents	147
14.2Gold Recovery	148
14.2.1Recovery from Heap Leaching	148
15INFRASTRUCTURE	150
15.1Site Access, Power and Water	150
15.1.1Site Access	150
15.1.2Power	150
15.1.3Operations Water Supply	150
15.2Buildings and Facilities	152
15.2.1Buildings and Facilities in Main Plant and Offices Area	152
15.2.2Additional Buildings and Facilities on Site	152
15.2.3Additional Facilities on Section 20	152
15.3Explosives Magazine	153
15.4Tailings Storage Facility and Water Diversion	153
15.5Leach Pads and Solution Ponds	153
15.6Waste Dumps	153
16MARKET STUDIES	156
16.1Marketing and Metal Prices	156
16.2Contracts	156
16.3QP Opinion	156
17ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS	157
17.1Environmental Studies	157

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17.2Environmental Permits	158
17.3Environmental Impacts	160
17.4Environmental Monitoring Program	160
17.5Reclamation and Closure	160
17.6Community Relations and Social Responsibilities	161
17.7QP Opinion	161
18CAPITAL AND OPERATING COSTS	162
18.1Introduction	162
18.2Capital Costs	162
18.3Operating Costs by Category	163
18.3.1Mine Operating Costs	163
18.3.2G&A	164
19ECONOMIC ANALYSIS	165
19.1Introduction	165
19.1.1QP Opinion on Inputs	165
19.2Mine and Leaching Production Statistics	166
19.3Sales and Refinery Process	167
19.4Revenue	167
19.5Operating Costs	168
19.6Royalties	168
19.7Cash Costs and AISC	168
19.7.1Taxation	169
19.8Excluded Costs	170
19.9Sensitivity Analysis	170
20ADJACENT PROPERTIES	173
21OTHER RELEVANT DATA AND INFORMATION	175

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22INTERPRETATION AND CONCLUSIONS	176
23RECOMMENDATIONS	178
23.1.1Processing	178
23.1.2Metallurgy / Analytical	178
23.1.3Mineral Resources	178
23.1.4Mine Planning	179
23.1.5Exploration Drilling	179
23.1.6Mine Operations	179
23.1.7Maintenance Operations	179
24REFERENCES	180
25RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT	184

TABLES

Table 1.1Summary of Marigold Mineral Resource Estimate Exclusive of Mineral Reserve (as at 31 December 2021) Based on $1,750/oz Gold Price	6
Table 1.2Summary of Metallurgical Recoveries of Marigold Mineral Resource Estimate Exclusive of Mineral Reserve (as at 31 December 2021) Based on $1,750/oz Gold Price	6
Table 1.3Summary of Marigold Mineral Reserve Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price	7
Table 1.4Summary of Metallurgical Recoveries of Marigold Mineral Reserve Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price	7
Table 1.5Historical Heap Leach Production and Recovery	8
Table 1.6Metal Prices	11
Table 1.7Summary of Capital and Reclamation Costs	11
Table 1.8Summary of Operating Costs	12
Table 1.9Marigold Key Economic Indicators	13
Table 1.10LOM Operating and Production Statistics Mining and Leaching Production	14

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Table 1.11LOM Average Costs per Payable Ounce of Gold Sold	15
Table 2.1Abbreviations and Acronyms	19
Table 2.2Units of Measurement	21
Table 3.1Marigold Surface Lands	25
Table 3.2MMC-Owned Unpatented Mining Claims	26
Table 3.3Decker Lease Unpatented Mining Claims	28
Table 3.4Vek & Andrus Lease Unpatented Mining and Millsite Claims	29
Table 5.1Marigold Historical Production: Tonnes, Grade, Contained, and Recovered Gold Ounces as of 31 December 2021	39
Table 5.2Marigold Mine Production April 2014 to 31 December 2021	39
Table 5.3Summary of Exploration Work Carried out to End of November 2021	40
Table 7.1Summary of Drilling History	73
Table 8.1Analytical Methods for Gold for the Marigold Assay Resource Database	76
Table 8.2Comparison of the NN Mean Gold Grades	81
Table 8.3List of Certified Standards Used Between 2014 and 2017	83
Table 8.4List of CRMs used between 2018 and 2021	86
Table 11.1Basic Au g/t Statistics of 7.6 m Bench Composites within the Mineralized Envelopes by Domain	100
Table 11.2Outlier Restriction Values and Distance for Various Domains	101
Table 11.3Summary of Density for Different Material	101
Table 11.4Correlogram Parameters Used to Estimate Different Domains	103
Table 11.5Cell Model Limits	104
Table 11.6Model Attributes	104
Table 11.7Probability Percentages for Cells Au>0.14 g/t	105
Table 11.8Estimation Parameters for Mineralized Stockpiles	106
Table 11.9Resource Classification Parameters	111

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Table 11.10Ore Reconciliation for the Period Between 2018 and 2021	112
Table 11.11Summary of Marigold Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price	113
Table 11.12Details of Marigold Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price	113
Table 12.1Summary of Marigold Mineral Reserves Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price	116
Table 12.2Details of Marigold Mineral Reserves Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price	116
Table 12.3Key Economic Parameters for Mineral Reserves Estimate	117
Table 13.1Overall Slope Angles by Azimuth	121
Table 13.2Mining Phase Design Summary	122
Table 13.3Annual Production Schedule Tonnes Mined	125
Table 13.4Marigold Mining Fleet Equipment List	127
Table 13.5RIB Design Criteria	143
Table 15.1Pump Assets	150
Table 16.1Metal Prices	156
Table 17.1Baseline Studies	157
Table 17.2Marigold Mine Permits 31 December 2021	159
Table 18.1Summary of Capital and Reclamation Costs	162
Table 18.2Summary of Operating Costs	163
Table 19.1Marigold Key Economic Indicators	166
Table 19.2Mining and Leaching Production	167
Table 19.3Forecast Metal Prices	168
Table 19.4Unit Operating Costs	168
Table 19.5Operating Costs per Payable ounce of Gold Sold	169
Table 19.6Cash and AISC Unit Costs	169
Table 19.7Marigold21TRS Gold Price Sensitivity	171

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Table 19.8Marigold21TRS Gold Price Sensitivity	171
Table 19.9Cash Flow	172
Table 20.1Past Production and Mineral Resources for Adjacent Properties	173

FIGURES

Figure 3.1Marigold Mine Location	23
Figure 3.2Marigold Corporate Structure	24
Figure 5.1View to the East–South-East over the Cyanide Leach Tanks from the Marigold Mine prior to World War II	35
Figure 5.2Location of Marigold Areas	37
Figure 6.1Location of the Marigold Mine in North-Central Nevada within the Basin and Range Physiographic Province	42
Figure 6.2Model of Shelf-Slope to Basin in late Cambrian-early Ordovician of Nevada, with Carbonate Rocks to East and Siliciclastic and Volcanic Rocks to West	43
Figure 6.3Schematic Model of Devonian - Mississippian Compression on the Western Margin of North America	44
Figure 6.4Major Igneous, Tectonic, and Mineralizing Events in Northern Nevada	46
Figure 6.5Location of Marigold and the Battle Mountain Mining District on the Battle Mountain-Eureka Mineral Trend	47
Figure 6.6Schematic Tectono-stratigraphic Section of the Rock Units at Marigold	50
Figure 6.7Plan View Map Showing Distribution of Paleozoic Units at Marigold.	51
Figure 6.8The Top Surface of the Valmy Formation with the Current Property Boundary	53
Figure 6.9Cross Section 11,200N Highlighting Inferred Permian Growth Fault and Associated Antithetic Normal Faults with a Steep West Dip	54
Figure 6.10Normal Displacement of Alluvium and Tuff Immediately South of the Basalt Pit	55
Figure 6.11Plan View of the Marigold Mine Area showing the Spatial Distribution of 1.0 g/t Au Grade Shells Over an 8 km Northerly Trend	57

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Figure 6.12Gold in Arsenian Pyrite Overgrowths on Pyrite Grains in Unoxidized Rock	58
Figure 6.13Native Gold Occurs with Iron Oxide in Weathered Rocks	58
Figure 6.14Cross Section 13,200N Highlighting Distribution of Gold in Antler Sequence and Valmy Formation Rocks	59
Figure 6.15Cross Section 16,000N Highlighting the HideOut Deposit Hosted by Antler Sequence Rocks	60
Figure 6.16Plan Reference of Cross Sections in Figure 6.9, Figure 6.14, and Figure 6.15	61
Figure 6.17Model Illustrating Inferred Processes Related to Formation of Carlin-Type Gold Deposits (CTGD) and Distal Disseminated Silver–Gold Deposits	63
Figure 7.1Marigold Mine Gravity Survey Stations in 2014	65
Figure 7.2Gravity Stations	66
Figure 7.3Marigold Mine Gravity Survey Compilation, Complete Bouguer Anomaly Oblique Image	67
Figure 7.4Plan View of Drilling Carried out on Trenton Canyon and Buffalo Valley	70
Figure 7.5Chart of Drilling Completed by SSR since 2014	71
Figure 7.6Plan View of Drilling Carried out by SSR Since 2014	72
Figure 7.7Plan View of All Drilling to End of November 2021	74
Figure 8.1Scatter Plot Between FA Gold Values with AA Finish and Gravimetric Finish	79
Figure 8.2Q-Q Plot between FA Gold Values with AA Finish and Gravimetric Finish	79
Figure 8.3Cross Section with SSR Drillholes (drillhole number prefix MRA) and Historical Drillholes Along Section 8000N	80
Figure 8.4Cumulative Normal Distribution Comparing Composites from SSR Drilling and Historical Drilling	81
Figure 8.5Blank Results	82
Figure 8.6Cumulative Frequency Distribution Comparing Original and Duplicate (field) Assay Results	84
Figure 8.7Blank Results	85

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Figure 8.8Cumulative Frequency Distribution Comparing Original and Duplicate (Field) Assay Results	87
Figure 9.1Cross Section Showing the Increase in Tonnage Estimated as Mineralized	90
Figure 10.1Column Test Results – All Marigold Areas	92
Figure 10.2Bottle Roll vs. Column Recovery – All Marigold Areas	92
Figure 10.3Exploration Database (2017) AuCN vs. AuFA – All Data	94
Figure 11.1Location of the Seven Major Domains over Depleted Topography as at 31 December 2017	96
Figure 11.2Oblique Plan View Showing the Major Structures	97
Figure 11.3Typical East–West Cross Section Along 10,300 N	98
Figure 11.4Typical Bench Plan (level=5000)	99
Figure 11.5Typical East–West Cross Section Along 10,400N Looking North, with Estimated Cell Grades Au g/t	107
Figure 11.6Typical Plan 4950 Elevation, with Estimated Whole Cell Grades Au g/t	108
Figure 11.7Swath Plot Along Eastings	109
Figure 11.8Swath Plot Along Northings	110
Figure 11.9Swath Plot Along Elevation	111
Figure 13.1Marigold Ultimate Pit (end of year 2032)	124
Figure 13.2Mine Annual Production Schedule	126
Figure 13.3End of Production Year 2022	130
Figure 13.4End of Production Year 2023	131
Figure 13.5End of Production Year 2024	132
Figure 13.6End of Production Year 2025	133
Figure 13.7End of Production Year 2026	134
Figure 13.8End of Production Year 2027	135
Figure 13.9End of Production Year 2028	136

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Figure 13.10End of Production Year 2029	137
Figure 13.11End of Production Year 2030	138
Figure 13.12End of Production Year 2031	139
Figure 13.13End of Production Year 2032	140
Figure 13.14Existing and Proposed Dewatering Wells	142
Figure 14.1Simplified Marigold Processing Flowsheet	146
Figure 14.2Average Annual Reagent Consumption	148
Figure 14.3Marigold Heap Leach Pad Gold Recovery Curve from March 1990 through December 2021	149
Figure 15.1LOM Site Map Showing Final Pit Limits, Waste Dumps, and Leach Pad	151
Figure 15.2Well Sites	154
Figure 15.3Main Infrastructure Area	155
Figure 15.4Plant, Shops and Offices	155
Figure 20.1Plan Map Showing Marigold Property Outline and mineralization Relative to Adjacent or Nearby Mines or Published Deposits	174

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1 EXECUTIVE SUMMARY

1.1 Introduction

The Marigold 2021 Technical Report Summary (Marigold21TRS) is an independent Technical Report Summary that has been in prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300) for SSR Mining Inc. (SSR), on the Marigold mine (Marigold or the Property).

The Property is located in Humboldt County, Nevada, U.S. and is directly owned by Marigold Mining Company (MMC), a wholly-owned (100%) subsidiary of SSR.

The Marigold21TRS Qualified Persons (QPs) have reviewed the supplied data and information and it appears accurate and complete and accept this information for use in the Marigold21TRS. Information and data supplied by SSR that were outside the areas of expertise of the QPs and was relied upon when forming the findings and conclusions of this report are detailed in Section 25. Any individual or entity referenced as having completed work relevant to the Marigold21TRS, but not identified therein as a QP, does not constitute a QP.

The Marigold21TRS should be construed in light of the methods, procedures, and techniques used to prepare the Marigold21TRS. Sections or parts of the Marigold21TRS should not be read or removed from their original context.

1.2 Property Description and Location

Marigold is located in south-eastern Humboldt County, accessible by public road off Interstate Highway 80 corridor in the northern foothills of the Battle Mountain Range, Nevada, US.

Activities at the Property are centred at approximately 40°45′ north latitude and 117°8′ west longitude.

The Property is situated approximately 5 km south–south-west of the town of Valmy, Nevada at Exit 216 off Interstate Highway 80. Other nearby municipalities include Winnemucca and Battle Mountain, Nevada, which lie approximately 58 km to the north-west and 24 km to the south-east of the Property, respectively.

1.3 Land Tenure and Ownership

The authorized Marigold Plan of Operations (PoO) area for Marigold currently encompasses approximately 10,703 ha with approximately 3,296 ha within the PoO permitted for mining-related disturbance. Land and mineral ownership within the PoO are within the corridor initially governed by the Pacific Railroad Act of 1862, and, as such, these areas generally have a “checkerboard” ownership pattern. Mineral claims in Nevada are managed federally by the Bureau of Land Management (BLM).

SSR holds a 100% interest in the Property through its wholly-owned subsidiary, MMC. Surface and mineral rights at the Property comprise the following: real property owned by MMC; unpatented mining claims owned by MMC; and leasehold rights held by MMC with respect to unpatented mining claims, mill site claims, and certain surface lands.

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Some of the leases require MMC to make certain net smelter return (NSR) royalty payments to the lessors and comply with other obligations, including completing certain work commitments or paying taxes levied on the underlying properties. The NSR royalty payments are based on the specific gold-extraction areas and are payable when the corresponding gold ounces are extracted, produced, and sold. The NSR royalty payments vary between 0% and 10.0% of the value of gold production, net of off-site refining costs, which equates to an annual average ranging from 3.7% to 10.0% and a weighted average of 7.8% over the life-of-mine (LOM).

1.4 Geology and Mineralization

The Property is located on the northern margin of the Battle Mountain-Eureka trend of mineralization, in the Battle Mountain Mining district, in north central Nevada, U.S.

1.4.1 Regional Geology

The western part of the North American continent has undergone a complex history of extensional and compressional tectonics from the Proterozoic through to the Quaternary. Predominantly Paleozoic rifting and basin subsidence led to the formation of thick (hundreds of metres) passive margin sedimentary sequences and repeated inter-plate collisions caused accretion of arc related volcanics and ocean floor rocks, which were pushed together with the basin sediments to form fold and thrust belts. Subsequent extension related to subduction and back arc basin rifting resulted in the development of Basin and Range topography. Crustal thinning caused by the extension allowed the rise of magma close to the surface, which produced extensive and voluminous magmatism from the middle Eocene to late Miocene. Crustal extension with bimodal (mafic and felsic) volcanism occurred in the region from the late Miocene to the present day.

Marigold is located in north-central Nevada within the Basin and Range physiographic province, bounded by the Sierra Nevada to the west and the Colorado Plateau to the east.

1.4.2 Local and Property Geology

Sedimentary Rocks

Four packages of Paleozoic sedimentary and metasedimentary rocks are present at Marigold. In ascending tectono-stratigraphic order, they include: the Cambro-Ordovician Preble-Comus Formation; the Ordovician Valmy Formation of the Roberts Mountain allochthon; the Pennsylvanian-Permian Antler overlap sequence; and the Mississippian-Permian Havallah sequence of the Golconda allochthon.

Comus-Preble Formation

The Comus-Preble Formation consists of fine-grained siliciclastic turbidite sequences, mudstone, siltstone, limey mudstone, limestone, debris flows, and mafic volcanic flows. Based on data compiled from downhole televiewer logs, abrupt lithologic change from overlying rocks correlates with a transition from tight, east-vergent, overturned folds to open folds.

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Valmy Formation

The Valmy Formation consists of quartzite, argillite, and lesser chert and metabasalt, all of which are complexly folded and faulted in the Marigold mine area. The total thickness of the Valmy Formation is approximately 450 m at Marigold, although true thickness of the section is likely less than 200 m. Where the contact is not eroded or structurally displaced, the top of the Valmy Formation is unconformably overlain by rocks of Pennsylvanian age.

Antler Sequence

The Antler overlap sequence is composed of Pennsylvanian to Permian-aged rocks assigned to three formations: the basal Battle Formation; the Antler Peak Limestone Formation; and the Edna Mountain Formation. These Formations represent a transgressive sequence of fluvial-to-shallow marine rocks that include conglomerate, sandstone, limestone, siltstone, and debris flows. Antler sequence rocks are relatively undeformed, except for offset and rotation along Basin and Range normal faults and potentially low-amplitude, long-wavelength (kilometres to tens of kilometres) F4 folding likely related to Mesozoic deformation. The Antler sequence is in thrust contact with the overlying and partially contemporaneous Havallah sequence.

Havallah Sequence

The uppermost package of Paleozoic rocks exposed at Marigold is the Mississippian-Permian Havallah sequence. The Havallah sequence is an assemblage dominated by siltstone, metabasalt, chert, sandstone, conglomerate, and carbonate rocks. These marine sedimentary rocks were deposited in a fault-bounded deep-water trough (Ketner, 2008) and subsequently obducted over the Antler sequence along the Golconda thrust (Roberts, 1964).

1.4.3 Property Structure

The main structural corridor and apparent primary controlling feature for the localization of the deposits at Marigold is a 1.5 km wide by >10 km long half graben rotated no more than 045° to the west and bound by east dipping early Permian growth faults and younger (post-Triassic) east dipping faults. This half graben structure is cut by north-west to north-east striking pre-mineralization structures with relatively minor offset and a series of south-west striking post-mineralization extensional normal faults parallel to the Oyarbide Fault.

1.4.4 Mineralization

The gold deposits at Marigold cumulatively define a north trending alignment of gold mineralized rock more than 8 km long.

Gold mineralizing fluids were primarily controlled by fault structure and lithology, with tertiary influence by fold geometry. The deposition of gold was restricted to fault zones and quartzite-chert dominant horizons within the Valmy Formation and high permeability units within the Antler sequence. Gold mineralization was also influenced by fold geometry in the Valmy Formation.

Rocks within the Marigold mine area are oxidized to a maximum depth of approximately 450 m. The redox boundary is not consistent throughout the Property and is substantially influenced by lithology. Shale, argillite, and siltstone units are frequently unoxidized adjacent to pervasively oxidized quartzite horizons. Gold occurs natively in fractures in association with iron oxide.

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1.4.5 Alteration

Alteration of rocks includes silicification along high-angle mineralizing structures and decalcification of carbonate horizons. Argillic alteration of quartz monzonite intrusive bodies occurs in fault zones and areas of high hydrothermal fluid flow. The intensity of alteration decreases towards the core of the intrusions.

1.5 Exploration

Currently, exploration work is performed by SSR staff. SSR self-funds all work to develop exploration targets.

1.5.1 Exploration – Marigold

1.5.1.1 Historical Exploration – Marigold

The first recorded gold production from Marigold was in 1938 from an underground mine. Approximately 9,000 t of ore averaging approximately 6.85 g/t Au was processed before production was halted by World War II. Several unsuccessful attempts were made to re-open and operate the mine before exploration activities re-commenced in 1968.

From 1968 through 1985, several companies took an interest in, and conducted exploration programs across, the Marigold area. The exploration activities during this time led to encouraging results and the acquisition of rights over additional parcels of land.

In 1986, a joint venture was formed between SFP Minerals (a subsidiary of Santa Fe Pacific Railroad) and the Cordex Group, which consolidated some of the land holdings over the Marigold area. In March 1988, a production decision was made on the 8S deposit, and by September 1988 stripping had begun on the 8S pit (McGibbon, 2004).

In August 1989, the first gold doré bar was poured at the Marigold mill.

In March 1992, Rayrock Mines (operating company for Cordex) purchased a two thirds ownership interest in the Property, and with the remainder held by Homestake Mining Company (Homestake).

In 1994, mining of the 8S deposit was completed, and the Marigold mill was no longer used to process ore. At this point, Marigold became a run-of-mine (ROM) heap leach operation.

Some five years later, under the ownership of Glamis Gold Ltd. (Glamis Gold), the Basalt, Antler, and Target II deposits were discovered in Section 31 at the south end of the Property and subsequently mined.

In 2007, discovery holes were drilled in the Red Dot deposit. and by mid-2009, a total of two million ounces of gold had been recovered from Marigold.

1.5.1.2 Exploration and Drilling Activities Since 2014 – Marigold

After SSR’s purchase of Marigold was completed in 2014, the exploration activities of previous owners were reviewed.

Between 2014 and 2016, SSR completed gravity surveys from 3,164 stations with the main objective of delineating possible fluid conduits or feeder structures for the Marigold mineralization.

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Meanwhile, in October 2015, the three millionth ounce was poured at Marigold.

SSRs exploration programs targeted the discovery of near-surface gold mineralization proximal to Marigold’s open pits and had the result of upgrading existing Inferred Mineral Resources to Indicated Mineral Resources. SSR drilled a total of 713 drillholes for 178,272 m from 2014 to 2017.

From 2018 through to the end of 2021, a further 995 holes have been drilled. This era of drilling included:

•950 reverse circulation (RC) holes,

•45 diamond core holes.

The 2018–2021 drilling adds a further 343,232 m of drilling to the Project database, bringing the total drilling in the history of the Marigold project to 9,435 drillholes for 1,988,280 m. This includes recent drilling on the Trenton Canyon and Buffalo Valley prospects.

Marigold has now been in continuous operation for more than 30 years and poured the four millionth ounce of gold in 2020.

1.5.2 Exploration – Trenton Canyon and Buffalo Valley

The Trenton Canyon project is located approximately 4 km south of New Millennium at Marigold and is one of three historically producing mines on a 100%-owned 8,900 ha parcel acquired from Newmont in 2019. The Buffalo Valley project is located approximately 10 km south-west of New Millennium.

Gold mineralization at Trenton Canyon is structurally controlled with significantly less dissemination than at Marigold, with the net result being higher gold grades in a smaller volume of mineralized rock.

Exploration work on the Trenton Canyon and Buffalo Valley properties consists of drilling, geophysical surveying, remote sensing, geochemical surveying, and mapping.

SSR has completed 13 exploration diamond core holes on Trenton Canyon totalling 10,131 m, and 249 RC drillholes for 73,165 m. As of December 2021, one diamond core hole has been completed at Buffalo Valley to a depth of 597.5 m.

1.6 Mineral Resources

The Mineral Resources for Marigold were estimated based on an optimized pit shell at a payable gold grade of 0.065 g/t (gold assay factored for recovery, royalty, and net proceeds) using an assumed gold price of $1,750/oz. The Mineral Resources are reported exclusive of Mineral Reserves in Table 1.1 and Table 1.2.

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Table 1.1 Summary of Marigold Mineral Resource Estimate Exclusive of Mineral Reserve (as at 31 December 2021) Based on $1,750/oz Gold Price

		Mineral Resource
	Measured				Indicated				Measured + Indicated				Inferred
	Tonnage<br>(Mt)		Au Grade<br>(g/t)		Tonnage<br>(Mt)		Au Grade<br>(g/t)		Tonnage<br>(Mt)		Au Grade<br>(g/t)		Tonnage<br>(Mt)		Au Grade<br>(g/t)
Marigold		–		–		115.3		0.43		115.3		0.43		21.8		0.36
Total		–		–		115.3		0.43		115.3		0.43		21.8		0.36

1. The Mineral Resource estimate was prepared in accordance with S-K 1300.

2. The Mineral Resource estimate is based on an optimized pit shell at a cut-off grade of 0.065 g/t payable gold (gold assay factored for recovery, royalty, and net proceeds), with a gold price assumption of $1,750/oz.

3. The Mineral Resources estimate is reported below the as-mined surface as at 31 December 2021 and is exclusive of Mineral Reserves.

4. The point of reference for Mineral Resources is the entry to the carbon columns in the processing facility.

5. Metallurgical recoveries used are, on average, 67% for gold.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

Table 1.2 Summary of Metallurgical Recoveries of Marigold Mineral Resource Estimate Exclusive of Mineral Reserve (as at 31 December 2021) Based on $1,750/oz Gold Price

Mineral Resource<br><br>Classification	Tonnage<br><br>(Mt)	Au Grade<br><br>(g/t)	Contained Gold <br>(koz)	Cut-off Grade<br><br>(Au g/t)	Metallurgical Recovery<br>(%)
Measured	–	–	–	–	–
Indicated	115.3	0.43	1,611	0.065	66%
Measured + Indicated	115.3	0.43	1,611	0.065	66%
Inferred	21.8	0.36	250	0.065	75%

1. The Mineral Resource estimate was prepared in accordance with S-K 1300.

3. The Mineral Resources estimate is reported below the as-mined surface as at 31 December 2021 and is exclusive of Mineral Reserves.

4. The point of reference for Mineral Resources is the entry to the carbon columns in the processing facility..

5. Metallurgical recoveries used are, on average, 67% for gold.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

1.7 Mineral Reserves

The Mineral Reserve estimate, shown in Table 1.3 and Table 1.4 reported in accordance with S-K 1300. The Mineral Reserves estimate is based on all available data for Marigold.

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Table 1.3 Summary of Marigold Mineral Reserve Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price

Mineral Reserve		Mineral Reserve
	Proven				Probable				Total
	Tonnage<br><br>(Mt)		Au Grade<br><br>(g/t)		Tonnage<br><br>(Mt)		Au Grade<br><br>(g/t)		Tonnage<br><br>(Mt)		Au Grade<br><br>(g/t)		Contained Gold <br>(koz)
In Situ		–		–		203.8		0.48		203.8		0.48		3,173
Leach Pad		–		–		–		–		–		–		237
Total		–		–		203.8		0.48		203.8		0.48		3,410

1. The Mineral Reserve estimate was prepared in accordance with S-K 1300.

2. The Mineral Reserve estimate is based on metal price assumptions of $1,350 gold.

3. The Mineral Reserve estimate is reported at a cut-off grade of 0.065 g/t Au.

4. Economic analysis for the Mineral Reserve has been prepared using long-term metal prices of $1,600/oz.

5. No mining dilution is applied to the grade of the Mineral Reserves. Dilution intrinsic to the Mineral Reserves estimate is considered sufficient to represent the mining selectivity considered.

6. The point of reference for Mineral Reserves is the entry to the carbon columns in the processing facility.

7. SSR has 100% ownership of the Project.

8. Metals shown in this table are the contained metals in ore mined and processed.

9. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

Table 1.4 Summary of Metallurgical Recoveries of Marigold Mineral Reserve Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price

Mineral Reserves	Tonnage<br><br>(Mt)	Au Grade<br><br>(g/t)	Contained Gold<br>(koz)	Cut-off Grade<br><br>(Au g/t)	Metallurgical Recovery<br>(%)
Proven In Situ	–	–	–	–	–
Probable In Situ	203.8	0.48	3,173	0.065	74.69
Probable Leach Pad	–	–	237	–	–
Total	203.8	0.48	3,410	0.065	74.69

1. The Mineral Reserve estimate was prepared in accordance with S-K 1300.

2. The Mineral Reserve estimate is based on metal price assumptions of $1,350 gold.

3. The Mineral Reserve estimate is reported at a cut-off grade of 0.065 g/t Au.

4. Economic analysis for the Mineral Reserve has been prepared using long-term metal prices of $1,600/oz.

5. No mining dilution is applied to the grade of the Mineral Reserves. Dilution intrinsic to the Mineral Reserves estimate is considered sufficient to represent the mining selectivity considered.

6. The point of reference for Mineral Reserves is the entry to the carbon columns in the processing facility.

7. SSR has 100% ownership of the Project.

8. Metals shown in this table are the contained metals in ore mined and processed.

9. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

1.8 Mining Operations

Marigold uses standard open pit mining methods at a LOM sustained mining rate of approximately 250,000 tpd. The mine conducts conventional drilling and blasting activities with a free face trim row blast to ensure stable wall rock conditions. Electronic detonators are used to control the timing of the blasthole detonation.

Mining occurs on 15.2 m benches for pre-stripping waste and selected ore areas when mining with the P&H electric shovel. One blasthole sample is taken for ore control Blasting is done with an ammonium nitrate and fuel oil (ANFO) blend and a sensitized ANFO emulsion. The ore control mark-out procedure includes blast movement analysis for 90% of ore production blasts.

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Loading operations are currently performed using one electric shovel and three hydraulic shovels. Waste and ore haulage is performed with a fleet of 300 t class haul trucks.

Equipment maintenance is performed on site for all equipment. There are no contract mining operations on site.

The Marigold geotechnical management plan (GMP) includes highwall monitoring using three radar systems which provide full coverage for the (largest) Mackay pit, or can be deployed in smaller pits, if required. Routine monitoring of waste dumps, leach pads and inactive pits using INSAR data is performed by a third party on a monthly basis.

1.9 Mineral Processing

The Marigold processing plant and processing facilities combine ROM heap leaching, carbon adsorption, carbon desorption and electro-winning circuits to produce a final precious metal (doré) product.

All processing of ore, which is oxide in nature, is completed via run-of-mine (ROM) heap leaching. ROM ore is delivered to the leach pad by haulage truck and stacked in 6.1 m to 12.2 m lifts. At any given time, approximately 0.5 Mm2 of pad area is being leached.

Barren leach solution (cyanide-bearing solution, very low in Au grade) is applied selectively to different areas of the heap leach pad.

The pregnant solution (gold-bearing) is then collected from the leach pad in pregnant solution pond(s) before it is pumped to carbon column trains where gold is adsorbed from solution onto activated carbon. Carbon loaded with gold is taken from the carbon columns and transported to the process facility where gold is stripped from the carbon by solution. The gold-bearing solution is passed through electro-winning cells where metals are plated out. The plated material is retorted for mercury removal and drying prior to smelting for final precious metal recovery.

From March 1990 through December 2021, gold recovery from the heap leach pad was 71.1%. Historical production figures for the Marigold heap leach pad are shown in Table 1.5. This recovery was achieved with 90–120 day primary leach. The current total gold recovery of more than 70% from ROM ore compares favourably to similar mining operations, and given current and past gold prices, suggests that a crushing circuit is not required.

Table 1.5 Historical Heap Leach Production and Recovery

Ore<br>(Mt)	Gold Loaded<br>(koz)	Au Grade<br>(AuFA g/t)	Gold Recovered (koz)	Gold Recovery (%)
324	5,490	0.53	3,904	71.1

Marigold uses an assay method that measures cyanide-soluble gold. This technique generates a value that represents the head grade of the ore in terms of the amount of gold in a finely ground sample that can be dissolved by a sodium cyanide solution. The gold content of the final solution is measured using atomic absorption (AA).

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All Marigold blasthole samples are assayed for cyanide-soluble gold. Samples from each ore polygon delineated by ore control are selected for fire assay based on the grade distribution for the polygon tonnage and targeting one sample per every 2,000 short tons of ore. Therefore, some samples have two assay values: an AuCN (cyanide soluble) value; and an AuFA (fire assayed) value. The ratio of AuCN:AuFA provides the theoretical maximum gold recovery that can be achieved.

Testwork has demonstrated that, generally, all ore at Marigold behaves similarly. The ratio of AuCN / AuFA is an important characteristic. A best fit linear regression from approximately 155,000 pairs of fire assays (field AUFA in the database) and cyanide soluble assays (field AUAA in the database shows the AuCN / AuFA ratio is 0.8037:1 (~80% cyanide soluble gold), based on the most recent assessment in 2017.

The LOM actual leach pad recovery is 74% (including in-process gold inventory through December 2021).

An adjustment factor can be calculated using the chemical maximum AuCN / AuFA recovery and the actual pad recovery:

Actual: 74% / Chemical: 80% = 0.92

Therefore, the estimated recovery from the ROM heap leach can be expressed as:

Heap Leach Recovery = AuCN / AuFA x 0.92

1.10 Infrastructure

Marigold is accessible via Interstate Highway 80 in northern Nevada and is approximately 5 km south–south-west of Valmy in Humboldt County. The site access road supports two lanes of traffic and consists of hard packed clay and gravel.

The infrastructure facilities at Marigold include ancillary buildings, offices and support buildings, access roads into the plant site, power distribution, source of fresh water and water distribution, fuel supply, storage and distribution, waste management and communications.

The power supply for Marigold is provided by NV Energy Inc. via a 120 kV transmission line to site. Site power draw is 5 MW. After exiting the main substation, power is distributed through a 25 kV distribution grid.

Water for Marigold is supplied from three existing groundwater wells located near the access road to the Property. Marigold owns groundwater rights and collectively allows up to 3.134 Mm3 of water consumption annually, the majority of which is used as makeup water for process operations. On average, total freshwater makeup is 2.4 m3/min. Approximately 5.3 m3/min of fresh water is required during peak periods in the summer months. The water is primarily consumed by retention in the heap leach pad, evaporation, processing operations and dust suppression.

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1.11 Environmental, Permitting and Social Responsibility

Significant portions of the Property exist on public lands administered by the BLM. Therefore, the majority of environmental studies related to mining activities are conducted under BLM authority as part of the National Environmental Policy Act (NEPA) regulations, which require various degrees of environmental impact analyses dictated by the scope of the proposed action. Marigold has undergone several significant NEPA actions in the normal course of operational planning; the most recent is an amendment to the existing PoO to permit the future mining of all pits to their planned maximum depths.

The environmental baseline studies to support the Environmental Impact Statement (EIS) process were initiated in 2013. These baseline studies completed in preparation for the Plan of Operations – Mackay Optimization Project Amendment included, but were not limited to, socioeconomics, air quality impacts, cultural and archaeological resources, groundwater model, pit lake model, screen level ecological risk assessment (SLERA), waste rock / material characterization, water characterization, sage grouse habitat evaluation, evaluations for flora and fauna, and feasibility evaluation and pilot testing for rapid infiltration basins.

The final EIS record of decision approving the amended plan of operations was received 30 October, 2019. Scope of the amended plan of operations included:

•Increasing surface disturbance by 833 ha on private and public lands.

•Consolidation of multiple pits into three larger pits with associated expansion of pits, waste rock storage areas and leach pads.

•Mining below the historical water table requiring installation of facilities to extract and dispose of excess groundwater.

The approval allowed for infiltration of excess dewatering water by way of rapid infiltration basins (RIBs) on private (leasehold) land north of the mine. A proposed relocation of the RIBs to an alternative location on adjacent BLM land is in the process of being approved. RIB testing, approval and construction, and the associated water pollution control permit issuance is expected by early 2023. In the interim, mine dewatering and infiltration is proceeding according to the LOM plan by means of temporary surface discharge permits allowing water diversion into local watercourses.

Subsequent to the EIS, a minor modification was submitted and approved through the BLM and the NDEP to increase the total approved disturbance to 3,296 ha; which was related to converting some land for heap leach cell 19B construction, modifying waste rock storage facilities, and converting some land to infill.

SSR has a reasonable expectation that all necessary operating permits will be granted within the required timeframes to meet the LOM plan.

1.12 Market Considerations

The metal prices used in the Marigold21TRS are based on an internal assessment of recent market prices, long-term forward curve prices, and consensus among analysts regarding price estimates. For the economic analysis in the Marigold21TRS, the metal prices shown in Table 1.6 were used.

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Table 1.6 Metal Prices

Metal	Unit	Average	2022	2023	2024	2025	2026	>2026
Gold Price	$/oz	1,647	1,800	1,740	1,710	1,670	1,600	1,600
Silver Price	$/oz	21.56	24.00	23.00	22.00	21.00	21.00	21.00

Marigold currently produces gold/silver doré bars. The doré refining terms are typical and consistent with standard industry practices and reflect similar contract conditions for doré refining worldwide.

The doré is securely transported by road freight to a refinery where it is refined into gold bullion. The bullion is sold by SSR to banks that specialise in the purchase and sale of gold bullion.

1.13 Capital and Operating Cost Estimates

Costs related to the development of reserves are based on a combination of historical site costs for fixed costs and a zero-base cost method for calculating variable costs. The variable costs are based on tonnage mined, tonnage processed, or hours worked for mining, maintenance, process, and administration costs. The total planned spend is divided by tonnes mined for mining and maintenance unit costs, and ore tonnes stacked for process and administration unit costs.

LOM project capital costs (excluding closure costs) are summarized in Table 1.7

Table 1.7 Summary of Capital and Reclamation Costs

Capital Costs	Total<br>($M)
Exploration and Development	9.1
Sustaining Capital	348.9
Mine Development	10.3
Total Capital Costs	368.3
Reclamation	71.8
Total Capital and Reclamation	440.1

Sustaining capital costs include:

•Replacement of mining equipment as it reaches its economic life during the remaining 11 years of mining. The majority relates to replacing haul trucks and excavators but is also covers drills and mine support equipment. Equipment replacement represents approximately 25% of future sustaining capital costs.

•Major equipment rebuilds and component replacement. In order to maintain equipment availability for the extended equipment lives, major equipment is programmed for rebuilds at set points during its economic life. Approximately 50% of future sustaining capital is capitalized parts and maintenance costs associated with these rebuilds. Major components with a life of more than one year are capitalized.

•Costs associated with ongoing expansion of the leach pad and associated process infrastructure represents about 8% of future capital.

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•Dewatering and permitting costs total about 17% of future sustaining capital, with the majority associated with dewatering infrastructure (wells, pipelines, rapid infiltration basins) that are required to lower the water table in advance of planned mine development.

1.14 Operating Costs by Category

The LOM operating costs estimate is $10.05/t of processed ore. Operating costs per tonne for the LOM and next five years of operations are shown in Table 1.8.

Table 1.8 Summary of Operating Costs

Operating Costs		Total LOM (M)	/t Ore
	Years 1–5
Mining		1,469	7.53
Processing		373	1.57
Site Support		214	0.92
Total Operating Cost		2,056	10.01

All values are in US Dollars.

Totals may vary due to rounding

1.15 Economic Analysis

Cash flow projections commenced on 1 January 2022 and are estimated over the remaining LOM based on estimates of sales revenue, site production costs, capital expenditures, and other cash flows, including taxes and reclamation expenditures, all presented on a real cash flow basis.

Cash inflows from sales assume all production within a period is sold, with minimal working capital movements, using the gold price in Table 1.6.

The estimates for site production costs, sustaining capital and reclamation expenditures have been developed specifically for Marigold and are presented in the relevant sections of the Marigold21TRS.

Based on SSR’s projections as set forth in the Marigold21TRS, Marigold will incur cash costs of $1,009 per payable ounce of gold sold and AISC of $1,154 per payable ounce of gold sold over the LOM. The after-tax NPV using a 5% discount rate and mid-year cash flows approach is $860M over the LOM.

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Table 1.9 Marigold Key Economic Indicators

Item	Unit	Amount
Oxide Processed
Heap Leach Quantity Placed	Mt	205
Gold Feed Grade	g/t	0.48
Total Gold Produced
Total Payable Gold	koz	2,536
Gold Recovery	%	79.71
5-Year Annual Average
Total Payable Gold Produced	koz/yr	215
Free Cash Flow	$M/yr	72
Total Cash Costs (CC)	$/oz payable Au	1,042
All In Sustaining Costs (AISC)	$/oz payable Au	1,278
Key Financial Results
Total Cash Costs (CC)	$/oz payable Au	1,009
All In Sustaining Costs (AISC)	$/oz payable Au	1,154
Site Operating Costs	$/t treated	10.07
After-Tax NPV	$M	860
Discount Rate	%	5
Mine (processing) Life	years	17

1. Recovery includes impact of starting pad inventory

2. Includes operating cost plus $0.02/t refining cost

3. Differences between the Mineral Reserve and LOM quantities used in the economic analysis are due to differences between planned and actual 31 December 2021 face positions

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1.15.1 Life-of-Mine Production

Mined material is either placed on the waste dumps or directly onto the leach pad over the course of 11 years of active mining.

A summary of projected mine and gold production over the LOM is shown in Table 1.10, resulting in total production of 2.5 Moz payable gold.

Table 1.10 LOM Operating and Production Statistics Mining and Leaching Production

Year	Total Mined <br><br>(kt)	Ore Mined <br><br>(kt)	Au Grade<br><br><br> (g/t)	Waste Mined <br><br>(kt)	Gold Recovery <br><br>(%)	Recover-able Gold Stacked (koz)	Gold Produced <br><br>(koz)
2022	102,616	21,818	0.53	80,798	76.4%	282.2	230.0
2023	90,646	22,010	0.38	68,637	76.2%	203.6	260.1
2024	92,828	21,410	0.42	71,418	76.4%	218.3	200.4
2025	93,988	15,713	0.50	78,275	74.9%	190.1	201.9
2026	90,633	16,538	0.40	74,095	76.1%	161.3	182.3
2027	83,225	20,857	0.40	62,369	74.0%	198.5	138.3
2028	90,358	20,207	0.64	70,151	73.8%	305.5	302.7
2029	78,934	26,911	0.43	52,023	71.5%	265.5	278.6
2030	93,861	18,188	0.47	75,673	72.4%	198.2	220.3
2031	87,403	13,103	0.68	74,299	75.6%	215.6	209.8
2032	28,105	7,792	0.72	20,313	75.6%	136.0	162.1
2033	–	–	–	–	–	–	77.7
2034	–	–	–	–	–	–	29.9
2035	–	–	–	–	–	–	15.0
2036	–	–	–	–	–	–	10.0
2037	–	–	–	–	–	–	7.0
2038	–	–	–	–	–	–	10.0
Total	932,597	204,547	0.48	728,050	74.7%	2,374.9	2,535.9

1. Gold produced from 2033 onwards is derived from the residual recoverable gold remaining in the leach pad when mining is completed and is recovered through continued leaching from 2033 to 2038.

2. Recoverable Gold Stacked on Pads refers to gold content of ore stacked on the pads in that period that is recoverable by the leaching process. Gold Produced refers to the amount of gold recovered from the heap in that period and processed to product for sale. The difference between the values in these columns is due to the lag effect of the leach cycle on gold dissolution in the heap and ounces already in the pads as of 1 January 2022.

3. Overall leaching recovery excludes impact of previously placed recoverable ounces

4. The mismatch between Mineral Reserves and the LOM production quantities is due to the difference between LOM plan and actual face positions at the end of 2021.

5. Totals may vary due to rounding.

1.15.2 Cost Statistics

Over the mine life, cash costs are estimated to average $1,009 per payable ounce of gold sold, and AISC is estimated to average $1,154 per payable ounce of gold sold.

Table 1.11 summarizes the estimated components of the cash costs and AISC per payable ounce of gold sold over the LOM.

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Table 1.11 LOM Average Costs per Payable Ounce of Gold Sold

Operating Costs	Value<br>($/payable oz of gold sold)
Mine Operations	579
Processing	147
General and Administration	84
Inventory Adjustment	54
Royalties and Refining (net of silver credits)	144
Total Cash Costs (CC)	1,009
Sustaining Capital	138
Other Capital	8
Total AISC	1,154

1. Inventory adjustment represents carrying values of starting leach pad and doré inventory at 1 January 2022, which are released into cash costs over the LOM through to 2038 as the associated gold ounces are sold.

2. Payable ounces of gold sold over the LOM total 2,535.9 koz.

3. Cash costs and AISC per payable ounce of gold sold are non-GAAP financial measures.

4. Totals may vary due to rounding.

1.16 Interpretation and Conclusions

The conversion of Mineral Resources to Mineral Reserves used industry best practices to determine operating costs, capital costs, and recovery performance. Therefore, the estimates are considered to be representative of actual and future operational conditions.

Possible areas of uncertainty that could materially impact the estimate of Mineral Reserves at Marigold include the commodity price assumptions, capital and operating cost estimates, estimation methodology, and the geotechnical slope designs for the pit walls. These reasonably foreseeable impacts of the uncertainties in the cost, operations and estimation assumptions are discussed in Section 22.

SSR has initiated exploration and Mineral Resources and Mineral Reserves development activities to enhance Marigold’s operating margins and extend the mine life. Further studies will examine the sulfide-hosted gold and could include further drilling evaluation and metallurgical testwork.

1.17 Recommendations

SSR should continue its commitment to safe gold production and continuous progress within the guidelines of its environmental and social license to operate drive improvements at Marigold. The following recommendations include work that has already been identified by SSR and in some cases is in progress.

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1.17.1 Processing

Pursue an upgrade to the barren pumping system in order to maintain a solution to ore ratio in excess of 1.5 as the leach pad increases height and new expansions are further from the barren ponds. An upgrade to the pumping system will aid in reducing the current work in-progress (WIP) inventory and decrease the likelihood of building inventory in the future. Perform a trade -off study between carbon-in-column (CIC) efficiency loss at high flow vs addition CIC trains with increased efficiency over the life of mine plan. The additional column trains to create a one-pass recovery system are a significant improvement to the system however there is still opportunity to optimize the flow rates through the columns and pull ounces forward through increased efficiency.

1.17.2 Metallurgy / Analytical

Investigate sample processing automation throughout the assay lab to decrease potential for bias and increase representivity. Continue work on fully implementing the ICP-OES to reduce detection limits for gold on leach pad, plant, and blasthole samples. Continue to conduct metallurgical test work with the goal of understanding all future leach ore at Marigold and how test results compare to resource model predictions. Perform more detailed test work on the sulfide ore types to better understand the value of this material at Marigold in the future. Utilize the new LECO machine for sulfur and carbon speciation both in current Marigold ore but also in conjunction with drilling activities being performed for near pit expansions. These data can be utilized to optimized reagent addition as well as reduce operational risks associated with preg-robbing material.

1.17.3 Mineral Resources

Incorporate geological data (from in-pit mapping) and hard boundaries (from faults that offset mineralization) into the resource model. Costs associated with this project are minor. As the mine progresses into zones below the water table, undertake a review of the effects of the water table on grade distribution and potential loss of fines.

Re-assay all samples that report the cyanide soluble gold assay values as zero and have not been assayed by the FA method outside of the current LOM pit designs. This should be conducted in a phased-in manner and will help convert Mineral Resources to Mineral Reserves and increase the volume of Mineral Resources and Mineral Reserves.

Collect additional density samples from core holes and in pit, where required, to obtain an improved spatial distribution of density values. Attempt to obtain additional samples from the upper levels of the deposit at between 0–152.4 m deep. It is planned by SSR that one sample be collected for every 9.1 m (30 ft) downhole from surface. The density testwork could be completed at Marigold’s on-site laboratory and a proportion of these samples should be sent to a commercial laboratory for QA/QC purposes.

Upgrade the Mineral Resources classifications and infill drilling program. Systematically design infill drill programs to increase the confidence of the model estimates based on the LOM plan within sparsely drilled areas and before ultimate pit walls are finalized.

1.17.4 Mine Planning

Develop and evaluate a digital twin of the mine haulage network utilizing industrial mathematics to iterate on material destinations over the LOM and optimise haulage profiles. Code projections of dewatering progress to the mine planning model. Record weekly plan variances, explanations, and associated actions for trending.

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1.17.5 Exploration Drilling

Conduct RC exploration drilling to target the lateral extensions of structures known to contain mineralization. This drilling could target near-surface, higher grade oxide mineralization. The estimated cost for this project is between $3M and $5M spent over a period of 3–5 years.

1.17.6 Mine Operations

To improve utilization of existing dispatch tools onsite as well as implementation of industrial mathematics-based haulage simulation tools for strategically optimized efficiencies throughout the LOM. Training of dispatch personnel for operation of updated fleet management systems onsite to optimise load / haul fleet resourcing and positively improve site productivity should be undertaken. Site will also deploy simulation software for strategic haulage network planning. This haulage simulator can be used to identify opportunities for mine planners and operations personnel to optimise material destinations.

1.17.7 Maintenance Operations

The Marigold maintenance team is committed to remain focused on improved maintenance operations at the site with the aim of increasing equipment availabilities and reducing unit costs. Projects underway include disciplined work planning and execution and consumables wear optimization.

Following improvements experienced from previous years’ initiatives, Marigold’s maintenance teams remain focused on improving quality of planned work execution at the site. Key areas of focus include plan compliance improvements, Komatsu PC7000 shovel availability increases, and coordination with supply chain for improved parts availability. These projects are primary enablers to ensuring site production requirements may be met.

In addition to systematic improvements, site has undertaken multiple trials to improve upon life of wear parts and maintenance consumables in the operation. These improvements include Shovel GET wear analysis, engine air filter pre-cleaners (de-risks potential supply chain shortages), and truck bed liner wear packages. Scale of sustaining improvements are pending based upon successful trials within the operation.

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2 INTRODUCTION

The Marigold21TRS has been prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300).

Marigold mine is owned directly by SSR’s wholly-owned subsidiary, Marigold Mining Company (MMC). SSR has reported that the total cost of the gross mineral properties, plant and equipment as of 31 December 2021 was $489.1M.

2.1 Terms of Reference

The Marigold21TRS is an independent Technical Report Summary (TRS) on the Project, prepared for SSR by the Marigold21TRS Qualified Persons (QPs). The TRS is based on information and data supplied to the QPs by SSR and other parties where necessary. Any individual or entity referenced as having completed work relevant to the Marigold21TRS, but not identified therein as a QP, does not constitute a QP. Marigold21TRS QPs have reviewed the supplied data and information and it appears accurate and complete and accept this information for use in the Marigold21TRS. The primary source of data for the Marigold21TRS is the Marigold 2021 Project Update Report.

Information and data supplied by SSR that were outside the areas of expertise of the QPs and was relied upon when forming the findings and conclusions of this report are detailed in Section 25.

The QPs have used their experience and industry expertise to produce the estimates and approximations in the Marigold21TRS. It should be noted that all estimates and approximations contained in the Marigold21TRS will be prone to fluctuations with time and changing industry circumstances.

The purpose of the Marigold21TRS is to report the Mineral Resources and Mineral Reserves for the project. This report is a Feasibility Study (FS) that represents forward-looking information. The forward-looking information includes metal price assumptions, cash flow forecasts, projected capital and operating costs, metal recoveries, mine life and production rates, and other assumptions used in the FS. Readers are cautioned that actual results may vary from those presented. The factors and assumptions used to develop the forward-looking information, and the risks that could cause the actual results to differ materially are presented in the body of this report under each relevant section.

The conclusions and estimates stated in the Marigold21TRS are to the accuracy stated in the Marigold21TRS only and rely on assumptions stated in the Marigold21TRS. The results of further work may indicate that the conclusions, estimates and assumptions in the Marigold21TRS need to be revised or reviewed.

The Marigold21TRS is intended to be used by SSR, subject to the terms and conditions of its contract with OreWin. Recognizing that SSR has legal and regulatory obligations, OreWin has consented to the filing of the Marigold21TRS with US SEC. Except for the purposes legislated, any other use of this report by any third party is at that party's sole risk.

A list of the references used to prepare the Marigold21TRS is provided in Section 24.

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2.2 Qualified Persons

The following people served as the QPs as defined in subpart 1300 of US Regulation S-K Mining Property Disclosure Rules (S-K 1300):

2.3 Qualified Persons Property Inspection

OreWin personnel Sharron Sylvester Technical Director – Geology and QP, and Graeme Baker, Principal Mining Consultant visited the Project 18–19 February 2020. The site visit included briefings from mining, geology, and exploration personnel; site inspections of potential areas for mining, including underground; discussions with staff; and review of the existing infrastructure and facilities around the Project site.

Bernard Peters has not visited the site.

2.4 Units and Currency

This Report uses metric measurements except where otherwise noted. The currency used is US dollars ($) unless otherwise stated.

2.5 Abbreviations and Acronyms

A list of abbreviations and acronyms used in the Marigold21TRS is shown in Table 2.1. The units of measurement used are shown in Table 2.2.

Table 2.1 Abbreviations and Acronyms

Abbreviation/ Acronym	Term/Definition
%	percent
°C	degrees Celsius
µ	micron
3D	three dimensional
5N	5 North
8D	8 Deep
8N	8 North
8S	8 South
8Sx	8 South Extension
AA	atomic absorption
ADR	adsorption/desorption/recovery
Ag	silver (element)
AISC	all‑in sustaining costs
---	---
amsl	above mean sea level
ANFO	ammonium nitrate and fuel oil
ARD	absolute relative difference
ARO	asset retirement obligation
As	arsenic (element)
ASX	Australian Stock Exchange
Au	gold (element)
AuEq	gold-equivalent
BLM	Bureau of Land Management
BTU	British thermal unit
CaO	lime
CIL	carbon‑in‑leach

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cm	centimetre
CN	cyanide
CRD	carbonate‑replacement deposit
CSR	corporate social responsibility
CTGD	Carlin‑type gold deposit
Cu	copper (element)
d	day
DEM	digital elevation model
E	east
EA	Environmental Assessment
EDA	exploratory data analysis
EIS	Environmental Impact Statement
EOY	end of year
FA	fire assay
g	gram
G&A	General and Administration
g/t	grams per tonne
GAAP	Generally Accepted Accounting Principles
GET	ground engaging tools
GMP	Geotechnical Management Plan
GPS	Global Positioning System
ha	hectare
Hg	mercury (element)
hp	horsepower
ID3	inverse distance cubed
IFRS	International Financial Reporting Standards
IRA	inter‑ramp angle
IRR	internal rate of return
ISO	International Organization for Standardization
J	joules
kg	kilogram
km	kilometre
km2	square kilometre
kV	kilovolt
kW	kilowatt
L/s	litres per second
---	---
LDL	lower detection limit
LOM	life‑of‑mine
M	million
m	metre
m2	square metre
m3	cubic metre
Ma	million years
MCC	motor control centre
min	minute
mL	millilitre
mm	millimetre
MMC	Marigold Mining Company
Moz	million ounces
Mt	million tonnes (metric)
MW	megawatt
tpd	tonnes per day
mV/V	millivolts per volt
Mya	million years ago
N	north
NaCN	sodium cyanide
NDEP	Nevada Division of Environmental Protection
NEPA	National Environmental Policy Act
NN	nearest neighbour
NPV	net present value
NSR	net smelter return
oz	Troy ounce
Pb	lead (element)
pH	acidity
PM	Preventative Maintenance
PoO	plan of operations
ppm	parts per million
QA/QC	quality assurance/quality control
QP	qualified person
RC	reverse circulation
ROM	run-of-mine

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RTK	Real-Time Kinematic
S	south
Sb	antimony (element)
SG	specific gravity
SI	International System of Units
SLERA	screen-level ecological risk assessment
SRTM	shuttle radar topography mission
SSR	SSR Mining Inc.
st	short ton (imperial)
t	tonne (metric)
t/m3	tonnes per cubic metre
TSF	tailings storage facility
TSX	Toronto Stock Exchange
---	---
TZN	Terry Zone North
US SEC	U.S. Securities and Exchange Commission
USGS	United States Geological Survey
UTM	Universal Transverse Mercator
VLF-EM	very-low-frequency electromagnetic
W	west
wk	week
yr.	year
Zn	zinc (element)

Table 2.2 Units of Measurement

Type	Unit	Unit Abbreviation	Si Conversion
area	hectare	ha	10,000 m2
area	square kilometre	km2	100 ha
concentration	grams per tonne	g/t	1 part per million
length	foot	ft	0.3048 m
length	mile	mi	1,609.34 km
mass	pound	lb	0.453592 kg
mass	troy ounce	oz	31.103481 g
mass	metric tonne	t or tonne	1,000 kg
mass	short ton	st or ton	2,000 lbs
temperature	degrees Celsius	°C	°C=(°F - 32) x 5/9

2.6 Effective Dates

The report has several effective dates, as follows:

•Effective date of the Technical Report Summary: 31 December 2021

•Effective date of Mineral Resources: 31 December 2021

•Effective date of Mineral Reserves: 31 December 2021

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3 PROPERTY DESCRIPTION

3.1 Property Location

Marigold is located in south-eastern Humboldt County along the Interstate Highway 80 corridor in the northern foothills of the Battle Mountain Range, Nevada, U.S. Activities at the Property are centred at approximately 40°45′ north latitude and 117°8′ west longitude.

Figure 3.1 shows the Property outline relative to these towns and Interstate Highway 80.

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Figure 3.1 Marigold Mine Location

SSR, 2021

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3.2 Corporate Structure

Marigold is 100% owned by SSR. The corporate structure of the subsidiary companies owned by SSR is shown in Figure 3.2.

Figure 3.2 Marigold Corporate Structure

SSR, 2021

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3.3 Land Tenure and Ownership

The authorized plan of operations (PoO) area for Marigold currently encompasses approximately 10,703 ha with approximately 3.296 ha within the PoO permitted for mining-related disturbance. Land and mineral ownership within the PoO are within the corridor initially governed by the Railroad Act, and, as such, these areas generally have a “checkerboard” ownership pattern. Mineral claims in Nevada are managed federally by the Bureau of Land Management (BLM).

SSR Mining Inc. (SSR Mining) holds 100% interest in the Property through its wholly owned subsidiary, Marigold Mining Company (MMC). Surface and mineral rights at the Property comprise real property owned by MMC; unpatented mining claims owned by MMC; and leasehold rights held by MMC with respect to unpatented mining claims and mill site claims and surface lands.

3.3.1 Owned Real Property

MMC owns the following surface lands at Marigold shown in Table 3.1.

Table 3.1 Marigold Surface Lands

Parcel<br>Number	Area (ha)	Location
007-0401-25	65.28	SE1/4 Section 22, T.34N, R.43E
007-0461-09	259.00	Section 9, T.33N, R.43E
007-0461-14	259.00	Section 17, T.33N, R.43E
007-0404-10, 007-0404-11, 007-0404-12, 007-0404-13 (Lot 8, Parcel 1-4), 007-0404-05 (Lot 11), 007-0404-06 (Lot 12), 007-0404-09 (Lot 15), 007-0403-03 (Lot 3)	84.40	Section 33, T.34N, R.43E
007-0461-42 (Parcel A) and 007-0461-43 (Parcel B)	259.00	Section 21, T.33N, R.43E
007-0461-44 (Parcel C) and 007-0461-45 (Parcel D)	259.00	Section 29, T.33N, R.43E
007-0481-06	254.40	Section 1, T.32N, R.42E
007-0491-03	277.90	Section 5, T.32N, R.43E
07-0461-39	16.19	Section 16, T.33N, R.43E
07-0461-41	32.37	Section 30, T.33N, R.43E
07-0491-02	64.75	Section 6, T.32N, R.43E
07-0481-13	16.19	Section 12, T.32N, R.42E

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3.3.2 Owned Unpatented Mining Claims

MMC owns a total of 323 unpatented mining claims at Marigold, as shown in Table 3.2.

Table 3.2 MMC-Owned Unpatented Mining Claims

BLM Serial Numbers	Claims	Total Number of Claims
NMC371561 to NMC371573	APRI # 1 to APRI # 13	13
NMC519580	APRI # 14	1
NMC552229	APRI # 15	1
NMC361136 to NMC361161	VAL #237 to VAL #262	26
NMC600391 to NMC600402	VAL #1013 to VAL #1024	12
NMC371574 to NMC371609	TYLER # 1 to TYLER # 36	36
NMC454876 to NMC454911	REMARY #237 to REMARY #272	36
NMC552228	REMARY FRACTION	1
NMC359040 to NMC359057	MARY # 73 to MARY # 90	18
NMC400277 to NMC400288	HS #123 to HS #134	12
NMC400289	HS #134A	1
NMC358968 to NMC359003	MARY# 1 to MARY # 36	36
NMC371610	BONZ # 1	1
NMC371612	BONZ # 3	1
NMC371614	BONZ # 5	1
NMC371616	BONZ # 7	1
NMC371618 to NMC371627	BONZ # 9 to BONZ # 18	10
NMC371630 to NMC371639	BONZ # 21 to BONZ # 30	10
NMC451485 to NMC451488	BONZ # 33 to BONZ # 36	4
NMC487422	REBONZ # 2	1
NMC487423	REBONZ # 4	1
NMC487424	REBONZ # 6	1
NMC487425	REBONZ # 8	1
NMC487426 to NMC487427	REBONZ # 19 to REBONZ # 20	2
NMC487428	REBONZ # 31	1
NMC524363	REBONZ # 32	1
NMC1112641 to NMC1112686	GINGER #1 to GINGER #46	46
NMC362237 to NMC362272	LCL #1 to LCL #36	36
NMC684371 to NMC674382	EJM #1 to EJM #12	12
Total Number of Claims		323

BLM Serial number reflect the old BLM-LR2000 system. Claims require annual maintenance fee / renewal notification by 1 September each year. All claims expire on 1 September 2022 at 11:59:59 A.M.

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3.3.3 Leasehold Rights

MMC holds leasehold rights in each of the following leases:

•Mineral Lease Agreement made and entered into as of 20 June 1986, by and between Donald J. Decker and Suzanne R. Decker, as lessors, Nevada North Resources (USA) Inc., as lessee, and Nevada North Resources Inc. (as amended, the Decker Lease).

•Lease Agreement made and entered into as of 15 September 1985, by and between Vek Associates, as lessor, and Rayrock Mines, doing business as Cordex, as lessee (as amended, the Vek & Andrus Lease).

•Lease Agreement made and entered into as of 1 August 1988, by and between Euro-Nevada Mining Corp., Inc., as lessor, and Rayrock Mines, doing business as Cordex, as lessee (as amended, the Euro-Nevada Lease).

•Lease Agreement made and entered into as of 1 August 1988, by and between the Board of Regents of the University of Nevada System, as lessor, and Donald J. Decker, Suzanne Decker, Nevada North Resources (USA) Inc., and Rayrock Mines, doing business as Cordex, as lessee (the University of Nevada Lease).

•Minerals Lease dated and effective 17 June 1988, by and between SFP Minerals Corporation, as lessor, and Santa Fe Pacific Mining, Inc., as lessee (the SFP Lease).

•Minerals Lease dated and effective as of 19 February 1986, by and between Southern Pacific Land Company, as lessor, and SFP Minerals Corporation, as lessee (the Southern Pacific Land Company Lease).

•Minerals Sublease dated and effective 30 April 1986, by and between SFP Minerals Corporation, as sublessor, and Santa Fe Pacific Mining, Inc., as sublessee (as amended, the Southern Pacific Land Company Sublease and, together with the Decker Lease, the Vek & Andrus Lease, the Euro-Nevada Lease, the University of Nevada Lease, the SFP Lease and the Southern Pacific Land Company Lease, collectively, the Leases).

•Minerals Lease Agreement made and entered into as of 5 June 1987, by and between Donald J. Decker and Suzanne R. Decker, as lessors, Nevada North Resources (USA) Inc. and Welcome North Mines (U.S.) Inc., as lessees (the Franco-Nevada Lease).

•Minerals Lease Agreement made and entered into as of 20 December 1994, by and between Nevada North Resources (USA), Inc. by and between Nevada North Resources (USA), Inc., as lessors, and Santa Fe Pacific Gold Corporation, as lessee (the Nevada North Lease).

•Minerals Lease Agreement made and entered into as of 16 October 2012, by and between New Nevada Resources, LLC and Lease Agreement made and entered into as of 16 October 2012 by and between New Nevada Resources, LLC and New Nevada Lands, LLC, as lessors, and Newmont Mining Company, as lessee (the New Nevada 2012 Lease).

•Minerals Lease Agreement made and entered into as of December 3, 2014, by and between New Nevada Resources, LLC and Lease Agreement made and entered into as of 3 December 2014 by and between New Nevada Resources, LLC and New Nevada Lands, LLC, as lessors, and Newmont Mining Company, as lessee (the New Nevada 2014 Lease).

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Decker Lease Claims

Pursuant to the Decker Lease, MMC has leasehold rights to 170 unpatented mining claims, as shown Table 3.3 The initial term for the Decker Lease was through 25 May 1991 and, thereafter, as long as operations continue.

Table 3.3 Decker Lease Unpatented Mining Claims

BLM Serial Numbers	Claims	Total Number of Claims
NMC48409 to NMC48412	RED # 21 to RED #24	4
NMC48415 to NMC48426	RED # 27 to RED # 38	12
NMC56187 to NMC56198	RED # 39 to RED # 50	12
NMC56199 to NMC56216	RED # 52 to RED # 69	18
NMC271665 to NMC271688	RED #201 to RED #224	24
NMC271689 to NMC271716	RED #601 to RED #628	28
NMC365642 to NMC365677	KIT # 1 to KIT # 36	36
NMC678030 to NMC678047	RED 1801A to RED 1818A	18
NMC678055 to NMC678063	RED 1826A to RED 1834A	9
NMC552226 to NMC552227	RED # 23A to RED # 24A	2
NMC871541 to NMC871547	NURED 1819 to NURED 1825	7
Total Number of Claims		170

BLM Serial numbers reflect the legacy serial numbers from the BLM-LR2000 system. The new serial numbers are non-sequential and can be found in the BLM-MLRS system. Claims require annual maintenance fee / renewal notification by 1 September each year

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Vek & Andrus Lease Claims

Pursuant to the Vek & Andrus Lease, MMC has leasehold rights to 205 unpatented mining and millsite claims, as shown in Table 3.4. The initial term of the Vek & Andrus Lease was through 15 September 1995 and runs for terms of 10 years and, at the lessee’s sole option, may be renewed for up to eight successive 10-year periods, upon prior written notice.

Table 3.4 Vek & Andrus Lease Unpatented Mining and Millsite Claims

BLM Serial Numbers	Claims	Total Number of Claims
NMC271972 to NMC272007	COT # 1 to COT # 36	36
NMC275733	COT # 38	1
NMC275750 to NMC275753	COT # 55 to COT # 58	4
NMC275755	COT # 60	1
NMC275757	COT # 62	1
NMC275759 to NMC275767	COT # 64 to COT # 72	9
NMC342068 to NMC342071	COT # 73 to COT # 76	4
NMC297554 to NMC297571	VAL # 1 to VAL # 18	18
NMC347463 to NMC347475	VAL # 19 to VAL # 31	13
NMC297572 to NMC297607	VAL # 37 to VAL # 72	36
NMC361164 to NMC361172	COT FRAC # 1 to COT FRAC # 9	9
NMC371559 to NMC371560	COT # 75A to COT # 76A	2
NMC822614	RECOT 37	1
NMC822615 to NMC822619	RECOT 39 to RECOT 43	5
NMC822620	RECOT 45	1
NMC822621	RECOT 47	1
NMC822622 to NMC822626	RECOT 50 to RECOT 54	5
NMC822627	RECOT 59	1
NMC822628	RECOT 61	1
NMC822629	RECOT 63	1
NMC822630	RECOT 63B	1
NMC822560 to NMC822613	GMMCMS 1 to GMMCMS 54	54
Total Number of Claims		205

BLM Serial numbers reflect the legacy serial numbers from the BLM-LR2000 system. The new serial numbers are non-sequential and can be found in the BLM-MLRS system.

NMC822560 to NMC822613 are Mill Site Claims and require annual maintenance fee / renewal notification by 1 September each year.

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Euro-Nevada Lease Claims

Pursuant to the Euro-Nevada Lease, MMC has leasehold rights to 36 unpatented mining claims, as shown in Table 3.5. The original term for the Euro-Nevada Lease was five years, and, at the lessee’s option, the Euro-Nevada Lease may be renewed for up to 10 additional and successive five-year periods, upon giving the lessor prior written notice. The last Euro-Nevada five year renewal notification was provided 25 May 2018.

Table 3.5 Euro-Nevada Lease Unpatented Mining Claims

BLM Serial Numbers	Claims	Total Number
NMC373649 to NMC373684	SAR# 37 to SAR# 72	36
	Total Number of Claims	36

BLM Serial numbers reflect the legacy serial numbers from the BLM-LR2000 system. The new serial numbers are non-sequential and can be found in the BLM-MLRS system.

Claims require annual maintenance fee/renewal notification by 1 September each year BLM Serial number reflect the old BLM-LR2000 system

University of Nevada Lease Claims

Pursuant to the University of Nevada Lease, MMC has leasehold rights to property in Section 19, T.33N., R.43E., Humboldt County, Nevada, identified as Humboldt County Assessor’s parcel number 007 461 19. The initial term of the University of Nevada Lease was ten years, and the lessee may renew the lease for successive ten-year periods upon providing the lessor with prior written notice. A new agreement was executed on 1 August 2018 and extends through 31 July 2038.

SFP Lease Claims

Pursuant to the SFP Lease, MMC has leasehold rights to property in Sections 5, 9, 17, and 31, T.33N., R.43E., Humboldt County, Nevada. The initial term of the SFP Lease was for 20 years or for so long, thereafter, as mining is conducted on a continuous basis.

Southern Pacific Land Company Lease Claims

Pursuant to the Southern Pacific Land Company Lease, MMC has leasehold rights to property in Sections 13 and 25, T.34N., R.42E.; Sections 19, 29, 31, and 33, T.34N., R.43E.; and Section 7, T.33N., R.43E., Humboldt County, Nevada. The initial term of the Southern Pacific Land Company Lease was for 25 years and for so long, thereafter, as the lessee continues to exercise its rights on any portion of the property.

Southern Pacific Land Company Sublease Claims

Pursuant to the Southern Pacific Land Company Sublease, MMC has leasehold rights to certain property in Sections 19, 29, 31, and 33, T.34N., R.43E.; Section 7, T.33N., R.43E.; and Sections 1, 13, and 25, T.33N., R.42E., Humboldt County, Nevada. The initial term of the Southern Pacific Land Company Sublease was for 25 years and for so long, thereafter, as the lessee exercises any rights granted by such sublease.

Franco-Nevada Lease Claims

Pursuant to the Franco-Nevada Lease, MMC has leasehold rights to 82 unpatented mining claims, as set out in Table 3.6. The initial term for the Franco-Nevada Lease was from 5 June 1987 for a period of 50 years and for so long, thereafter, as the lessee exercises any rights granted by such lease.

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Table 3.6 Franco-Nevada Lease Unpatented Mining Claims

BLM Serial Numbers	Claims	Total Number of Claims
NMC379514 to NMC379585	N-1 to N-72	72
NMC623992 to NMC623995	N-109 to N-112	4
NMC676435	N-20A	1
NMC676436	N-22A	1
NMC676437 to NMC676440	N-28A to N-31A	4
Total Number of Claims		82

Claims require annual maintenance fee/renewal notification by 1 September each year

Nevada North Lease

Pursuant to the Nevada North Lease, MMC has leasehold rights to 12 unpatented mining claims, as set out in Table 3.7. The initial term for the Nevada North Lease was from 20 December 1994 for a period of 10 years and for so long, thereafter, as the lessee exercises any rights granted by such lease.

Table 3.7 Nevada North Lease Unpatented Mining Claims

BLM Serial Numbers	Claims	Total Number of Claims
NMC409224 to NMC409235	BC-1 to BC-12	12
Total Number of Claims		12

Claims require annual maintenance fee/renewal notification by 1 September each year

New Nevada 2012 Lease

Pursuant to the New Nevada 2012 Lease, MMC has leasehold rights to property in Section 33, T.33N., R.43E., Humboldt County, Nevada. The initial term for the New Nevada 2012 Lease was from 16 October 2012 for a period of 20 years and for so long, thereafter, as the lessee exercises any rights granted by such lease.

New Nevada 2014 Lease

Pursuant to the New Nevada 2014 Lease, MMC has leasehold rights to property in Sections 11, T.33N, R.44E; Section 23, T.33N, R.42E; and Section 35, T.33N, R.42E, Humboldt County, Nevada. The initial term for the New Nevada 2014 Lease was from 3 December 2014 for a period of 20 years and for so long, thereafter, as long as the lessee exercises any rights granted by such lease.

3.4 Royalties and Encumbrances

Each Lease requires MMC to make certain net smelter return (NSR) royalty payments to the lessors and comply with certain other obligations, including completing certain work commitments or paying taxes levied on the underlying properties. These NSR royalty payments are based on the specific gold-extraction areas and are payable when the corresponding gold ounces are extracted, produced and sold. The NSR royalty payments vary between 0% and 10.0% of the value of gold production net of off-site refining costs, which equates to an annual average ranging from 3.7% to 10.0% and a weighted average of 7.8% over the life-of-mine (LOM).

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3.5 Environmental Liabilities

At present, there are no known environmental liabilities to which the Property is subject. Further discussion on environmental matters with respect to the Property is provided in Section 17.

3.6 Operating Permits

Marigold holds active, valid permits for all facets of the current mining operation as required by county, state, and federal regulations. MMC performs duties on leased lands pursuant to all federal and state requirements, and all the Leases are maintained in good standing. MMC engages in concurrent reclamation practices and is bonded for all permitted features, as part of the Nevada permitting process.

Further discussion on permitting requirements with respect to the Property is provided in Section 17.

3.7 Permits, Mineral, and Surface Rights

Mining activities at Marigold are authorized by and conducted under both federal and state regulatory requirements, notably the General Mining Law of 1872, the National Environmental Policy Act of 1970, and the Federal Land Policy and Management Act of 1976. All requirements are administered by the BLM, along with applicable statutes and regulations within the Nevada Revised Statutes and Nevada Administrative Code, administered by the Nevada Division of Environmental Protection.

Further discussion regarding Marigold’s mineral and surface rights, including leasehold rights under the Leases, is provided in Section 3. Further discussion regarding permitting requirements with respect to the Property is provided in Section 17.

3.8 Other Significant Factors and Risks

SSR have advised that there are no other known significant risks that may affect access, title or the right or ability to perform mining-related work on the Property.

SSR have advised that there are no other known significant risks that may affect access, title or the right or ability to perform mining related work on the Property.

Legal matters such as statutory and regulatory interpretations affecting the mine plan and environmental matters are outside the expertise of the QPs (see Section 25).

The Marigold21TRS QPs considers it reasonable to rely on SSR because SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

Additionally, the project has been in operation for a number of years, and following a review of the current supplied information, the opinion of the QPs is the current plans appear adequate to address any issues related to environmental compliance, permitting, and local individuals or groups.

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4 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

4.1 Access

Access to the Property is via a 5 km public road (hard-packed clay and gravel) off the Valmy exit (Exit 216) on Interstate Highway 80.

4.2 Climate and Physiography

Elevations at Marigold range from approximately 1,372–1,890 mamsl. The climate is typical of the Great Basin region of the western U.S., with temperatures ranging from highs of 40°C in summer to lows of –7°C in winter. Annual precipitation is relatively low, ranging from 15–20 cm per year, with approximately 50% of precipitation occurring as snowfall during the months of December through March.

The climate presents no restrictions on the operating season, and Marigold operates year-round. Terrain varies from a relatively flat alluvial plain to sloped foothills at the base of the Battle Mountain Range. Vegetation mainly comprises sagebrush, rabbit brush, and a variety of grasses and forbs. Fauna is not abundant on the Property primarily due to the lack of surface water and limited forage. No threatened or endangered plant or animal species have been noted within the Property’s operating area.

4.3 Infrastructure

Marigold has been in continuous operation since 1989. There is significant infrastructure existing on site for delivering power and water to the various mine shops, leach pad, and process and ancillary facilities. The Property is located in a favourable area for natural resource development with significant resources in place to support the mining industry. The nearby towns of Winnemucca and Battle Mountain host the majority of the local workforce. Contractor support, transportation, and general suppliers are all readily available in these communities as well as in Elko, which is located approximately 142 km east of Marigold and serves as a major hub for mining operations in northern Nevada. Employees are transported to the Property primarily by contract buses and light-duty vehicles owned by MMC.

Approximately 5.3 m3/min of fresh water is required during peak periods in the summer months. The water is primarily consumed by retention in the heap leach pad, evaporation, processing operations and dust suppression. Marigold also owns 0.893 Mm3 annually of surface water storage rights associated with the Trout Creek Dam (J-666). In addition, in September 2019, Marigold was issued water rights permits associated with the activities described in the Plan of Operations – Mackay Optimization Project Amendment, including permits for the dewatering during mine operations and evaporative losses from a future pit lake that will develop in closure.

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The tailings storage facility (TSF) has been decommissioned and reclaimed. The only remaining activity concerning the TSF is ongoing monitoring.

Details regarding completed, in progress, and future waste dumps at Marigold can be found in Section 13. The leach pad is discussed in detail in Section 14. Further discussion on the Property’s infrastructure is provided in Section 15.

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5 HISTORY

5.1 Historical Exploration Work

The first recorded gold production from the Property near Valmy, Nevada occurred in 1938 when the Marigold Mining Company, owned by Frank Horton, developed and operated an underground mine which came to be known as Marigold. Figure 5.1 shows the Marigold mine prior to World War II.

The Horton family processed approximately 9,000 t of ore averaging about 6.85 g/t Au before World War II halted production. In 1943, Mr. Horton’s estate sold its interest in the Property and claims. Several unsuccessful attempts were made to open and operate the mine before exploration activities began again in 1968.

Figure 5.1 View to the East–South-East over the Cyanide Leach Tanks from the Marigold Mine prior to World War II

SSR, 2017

From 1968 through 1985, several companies conducted exploration programs in the Marigold area and completed a total of 126 exploratory drillholes. Records document the activities of Homestake (1968), St. Joe (1979), Decker Exploration (1979), Placer Amex (1979–1980), True North, Marigold Development Company (MDC) (1981–1983), Welcome North (1984), and Nevada North Resources (USA) Inc. (1985–1986). Other groups that conducted work in the area include Newmont, Kerr-McGee, SFP Minerals Corporation, Cordex/Rayrock Mines, and Vek/Andrus Associates (partnership between Vic Kral, Ralph Roberts, Bob Reeve, and Bill Andrus composed of Vek Associates and Andrus Resources Corporation).

From 1983 through 1984, MDC excavated a small open pit over the historical Marigold underground workings, producing 2,812 t containing 271 oz gold (McGibbon, 2004).

In 1985, Vek/Andrus Associates drilled three holes under the supervision of Ralph Roberts in the Section 8 area of the Property, just north-east of the old underground mine. Roberts invited Andy Wallace of Cordex to view the drilling results, and Wallace was encouraged by the deep level of oxidation, presence of favourable rock units, anomalous indicator elements, and anomalous gold values. The operating partner Cordex, an exploration syndicate composed of Dome Exploration (U.S.) Ltd., Lacana Gold Inc. (Lacana) and Rayrock Mines, leased the Vek/Andrus Associates claim block in September 1985 and began a drilling program in November 1985. Drillholes NM-3 and NM-4 intersected 21.3 m

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of 2.40 g/t Au and 25.9 m of 7.54 g/t Au, respectively. These were the discovery holes for the 8 South (8S) ore body (Roberts, 2002).

The Property is within the “checkerboard” railway lands, where the U.S. Government originally awarded the surface, water, and mineral rights for alternate sections (2.5 km2 of land) to the Santa Fe Pacific Railroad as an incentive to develop the transcontinental railway project in the 1860s. Santa Fe Pacific Railroad eventually became the parent company of SFP Minerals. Following further drilling in the 8S deposit in the spring of 1986, a joint venture was formed between SFP Minerals and the Cordex group, which consolidated some of the land holdings over the Marigold area.

In late-1986, the Cordex group leased other claims, including the historical Marigold mine, Top Zone, East Hill, and Red Rock area from various claim holders (Figure 5.2).

In March 1988, Rayrock Mines (operating company for Cordex) made a production decision on the 8S deposit, and, by September 1988, it began stripping on the 8S pit (McGibbon, 2004).

In August 1989, the first gold doré bar was poured at the Marigold mill.

In March 1992, Rayrock Mines purchased a two thirds ownership interest in the Property, and Homestake Mining Company (Homestake), which had taken Lacana’s interest through previous corporate mergers, held the remaining one third ownership interest in the Property.

In 1994, mining of the 8S deposit was completed, and the Marigold mill was no longer used to process ore. At this point, Marigold became a run-of-mine (ROM) heap leach operation.

In March 1999, Glamis Gold Ltd. (Glamis Gold) purchased all the assets of Rayrock Mines, resulting in Glamis Gold holding a two thirds ownership interest in Marigold, and Homestake continuing to hold a one third ownership interest. In the same year, the Basalt, Antler and Target II deposits were discovered at the south end of the Property in Section 31. These deposits were mined and partially backfilled with the unmined East Basalt deposit which is currently under development as an easterly extension of the original Basalt pit.

By January 2001, a total of one million ounces of gold had been recovered from the Property. In July 2001, Glamis Gold released a revised NI 43-101 Technical Report (Glamis Gold, 2001) to report the Mineral Resources and Mineral Reserves for Section 31 of the Property.

In 2006, Glamis Gold merged with Goldcorp Inc. (Goldcorp), resulting in a Goldcorp subsidiary holding a two thirds ownership interest in Marigold, as operator, and Homestake, which had been acquired by Barrick Gold Corporation (Barrick) in 2001, continued to hold the remaining one third ownership interest.

In 2007, discovery holes were drilled in the Red Dot deposit.

By mid-2009, two million ounces of gold had been recovered from Marigold.

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Figure 5.2 Location of Marigold Areas

SSR, 2021

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On 4 April 2014, SSR (formerly Silver Standard Resources Inc.) completed the acquisition of Marigold from subsidiaries of Goldcorp and Barrick. Subsequently, SSR filed an updated NI 43-101 Technical Report in November 2014 to support the October 2014 press release that announced the estimates of Mineral Resources and Mineral Reserves, and the LOM at Marigold.

In August 2015, Marigold mine acquired 2,844 ha of adjacent land from Newmont. This land included previously mined areas known as the Mud pit, NW pit, and the Valmy pits. Exploration drilling in the area had been completed by a combination of companies including Hecla Mining Company (Hecla), SFP Minerals, and Newmont.

In October 2015, the three millionth ounce was poured at Marigold. Marigold has now been in continuous operation for more than 30 years and poured the four millionth ounce of gold in 2020.

On 31 July 2018, SSR filed an updated NI 43-101 Technical Report (SSRTR18) to support press release 18-09 (18 June 2018) updating the life-of-mine plan and confirming near-term production growth and robust economics. In that report it was stated that, as at 31 December 2017, a total of 8,440 drillholes for 1,645,048 m of drilling had been completed on the Property.

From 2018 through to the end of November 2021, a further 975 holes have been drilled, including 932 RC holes and 43 diamond core holes (24 with RC pre-collars). This adds a further 337,910 m of drilling, bringing the total to 9,323 drillholes for 1,940,438 m.

A summary of the exploration work carried out on the Property is shown in Table 5.3.

Further discussion on historical drilling programs with respect to the Property is provided in Section 7.

5.2 Historical Production Work

Historically, gold recovery at Marigold was initially a milling circuit with a carbon-in-leach (CIL) process and then a ROM heap leach process where the ore is dumped on a lined leach pad and irrigated with a dilute cyanide solution. The tonnes, grade, and contained and recovered ounces from the start of commercial production in August 1989 to 31 December 2021 is provided in Table 5.1.

An overall average recovery for the milling circuit was 92%, and it is calculated to be at 73% with the ROM heap leach process until October of 2010 when the recovery equation was updated.

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Table 5.1 Marigold Historical Production: Tonnes, Grade, Contained, and Recovered Gold Ounces as of 31 December 2021

Process Type	Tonnes<br><br>(Mt)	Au Grade<br>(g/t)	Contained Gold (koz)	Recovered Gold (koz)
Leach Pad	323.9	0.527	5,489	3,904
Milled	4.6	3.13	483	458
Total	328.5	0.56	5,973	4,362

The Marigold mine production for April 2014 to 2021 is shown in Table 5.2.

Table 5.2 Marigold Mine Production April 2014 to 31 December 2021

Mine Production	Tonnes<br>(Mt)	Au<br>(g/t)	Contained Ounces<br>(koz)
April 2014—31 December 2021	177.7	0.41	2,350

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Table 5.3 Summary of Exploration Work Carried out to End of November 2021

Year			Company		Exploration Type	Details
1968–1985			Various exploration and mining groups		Drilling	7,037.2 m in 126 drillholes.
1985–1999			Cordex and Rayrock Mines		Drilling	335,500.7 m in 2,358 drillholes.
	Geophysics			1989 – CSAMT survey conducted by Quantec Geoscience using Zonge CSAMT System covering 33 EW and NW-SE lines, spaced 300.3 m and 499.9 m. A total of 59.2 km covered.
		1997/1999 – CSAMT survey conducted by Zonge Geoscience using Zonge CSAMT System covering 33 EW and NW–SE lines, spaced 300.3 m and 499.9 m. A total of 51.8 km covered.
		1998 – Gravity survey conducted by Zonge Geoscience using Scintrex Gravity Meter, Trimble GPS System survey conducted on 150 m square grid and data collected from a total of 1,252 stations.
		1999 – Induced Polarization conducted by Zonge Geoscience using Zonge IP system, Dipole-Dipole Array, A = 182.9 m, 1 line N20W. A total of 3.0 km covered.
1999–2006			Glamis Gold		Drilling	486,648.9 m in 2,506 drillholes.
	Geophysics			2004 – Airborne Magnetic conducted by Pearson, deRidder & Johnson, Inc. using Ultra Light System / 75.0 m EW flight lines, 300.3 m NS tie lines. A total of 323.5 km covered.
2006–2013			Goldcorp		Drilling	528,225.7 m in 1,870 drillholes.
	Geophysics			2009 – Magneto-telluric/Induced Polarization survey conducted by Quantec Geoscience, using Quantec Titan System. 11 lines in various orientations. A total of 46.4 km covered.
		2010 – Induced Polarization conducted by Zonge Geoscience using Zonge IP system, Dipole-Dipole Array, A= 150.0 m and 200.0 m, 27 lines EW, spaced 300.3 m –1,499.9 m. A total of 117.5 km covered.
		2009–2010 – Review of all geophysical survey data and compilation of Marigold geophysical data by J L Wright Geophysics.
2006–2013 cont.d			Goldcorp		MMI Survey	2007–2009 – Initial survey in 2007 covered Red Dot area, and, in 2008–2009, most of undisturbed land within Marigold was covered. A total of 11,493 samples were taken. Samples collected every 15.2 m along 117 EW lines separated by 30.5 m. In 2007, samples were analysed for Ag, As, Au, Ba, Cd, Co, Cu, Pb, Pd, Sm, Y, Zn, and Zr. In 2008, Pd was dropped. In 2009, Co, Sm, Y, and Zr were dropped and replaced with Mg, Sr, and Sb.

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1985–2006	Newmont (including Hecla and SFP Minerals)	Drilling	109,363 m in 867 drillholes. Data was acquired from Newmont with the acquisition of the 2,844 ha Valmy property in 2015.
2014	SSR	Geophysics	J L Wright Geophysics conducted a gravity survey. Magee Geophysical Services LLC conducted the field data collection. The gravity measurements were collected from 1,358 stations using two LaCoste and Romberg Model-G gravity meters at a grid spacing of 150 m x 150 m. (Magee, 2014)
2014–2017	SSR	Drilling	178,272 m in 713 drillholes.
2016	SSR	Geophysics	Gravity survey conducted by Magee Geophysical Services LLC. A total of 1,806 stations were acquired on a 150 m square grid and 150 m x 300 m staggered grid. Relative gravity measurements were made with LaCoste and Romberg Model-G gravity meters. Topographic surveys were performed with Trimble Real-Time Kinematic (RTK) and Fast-Static GPS. (Magee, 2016)
2018–Nov.2021	SSR	Drilling	343,233 m in 995 drillholes (259,339 m in 732 drillholes in Marigold; 83,894 m in 263 holes in Trenton Canyon and Buffalo Valley)
2020	SSR	Geophysics	Two reflection seismic lines covering 16.9 km. The lines were surveyed by Riolada Surveying LLC and Xtreme Drilling completed the shot holes. Bird Seismic acquired the data, and processing was completed by SubTerraSies and Wright Geophysics.
2021	SSR	Geophysics / Soil Samples	In 2021 a proprietary airborne hyperspectral dataset was acquired with district-scale coverage. This dataset includes mineral maps generated from short and long wave infrared sources. a soil sampling program was completed by North American Exploration on behalf of SSR Mining consisting of 3,284 soil samples covering 14.5 km2 of mountainous terrain predominantly east the previously mined pits at Trenton Canyon.

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6 GEOLOGICAL SETTING, MINERALIZATION, AND DEPOSIT

6.1 Regional Geological Setting

Marigold is located in north-central Nevada within the Basin and Range physiographic province bounded by Sierra Nevada to the west and the Colorado Plateau to the east (Figure 6.1).

Figure 6.1 Location of the Marigold Mine in North-Central Nevada within the Basin and Range Physiographic Province

Modified after Hamilton, 1987

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Paleozoic basement rocks of north-central to north-eastern Nevada generally comprise four distinct tectonostratigraphic assemblages: the eastern carbonate assemblage; the slope or transitional assemblage; the western siliceous and volcanic assemblage; and the overlap assemblage (Roberts, 1964). These rocks record a complex history of compressional and extensional tectonics affecting the western margin of North America from the early Paleozoic through present.

Late Proterozoic rifting associated with the breakup of Rodinia resulted in passive margin sedimentation on the miogeocline of the proto-Pacific margin of western North America (Cook and Taylor, 1977; Wallace et al., 2004; Cook, 2015). Subsidence and sedimentation continued along the passive margin from the Late Proterozoic through Devonian, a period of approximately 240 million years (Cook and Taylor, 1977; Cook, 2015). Carbonate platform rocks (eastern assemblage) 4,800 to 7,000 m thick developed on the eastern margin of the miogeocline. Debris flow, turbidite, and lime mudstone of the transitional assemblage accumulated on the slope further west, and siliceous and volcanic rocks belonging to the western assemblage were deposited in the basin plain (Figure 6.2) (Roberts, 1964; Cook and Corboy, 2004; Cook, 2015).

Figure 6.2 Model of Shelf-Slope to Basin in late Cambrian-early Ordovician of Nevada, with Carbonate Rocks to East and Siliciclastic and Volcanic Rocks to West

Cook and Corboy, 2004

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Evidence for an enigmatic late Devonian to Early Mississippian tectonic event, known as the Antler orogeny, is recorded by folding and thrusting of Ordovician western assemblage rocks and formation of the Antler highland (Roberts, 1964). In north-central Nevada, western assemblage rocks are tectonically emplaced over eastern assemblage rocks along the Roberts Mountain thrust, although the legitimacy of the thrust is disputed (Ketner, 2013). Uplift and erosion of the Antler highland in the Pennsylvanian shed clasts of western assemblage rocks into a foreland basin, forming basal units of the Pennsylvanian-Permian overlap assemblage (Figure 6.3).

Figure 6.3 Schematic Model of Devonian - Mississippian Compression on the Western Margin of North America

Cook and Corboy, 2004

Marine sedimentary rocks and submarine volcanic rocks accumulated in a basin west of the Antler orogenic belt from the Mississippian to the Permian. These rocks were transported eastward and structurally emplaced on top of western assemblage and overlap assemblage rocks along the Golconda thrust during the Permo-Triassic Sonoma orogeny (Roberts, 1964). The mechanism for compression resulting in the Sonoma orogeny is controversial, and modern work by Ketner (2008) has called into question the relationship between the Sonoma orogeny and the Golconda thrust.

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Compression during the Jurassic and Early Cretaceous resulted in subduction of oceanic plate material beneath continental crust of western North America, generating large volumes of intermediate to felsic melts along a magmatic arc and emplacement of plutons into the Sierra Nevada batholith. Continued compression resulted in accretion of oceanic arc terrane onto the continental margin, forming thrust belts and ophiolite sequences. Collectively, these Andean and Cordilleran style compression events are known as the Nevadan orogeny. The Nevadan orogeny resulted in substantial back-arc shortening and formation of the Luning-Fencemaker fold-thrust belt in Nevada (Wyld et al., 2003). A major mode of felsic plutonism also occurred in Nevada during the late Jurassic (~155–160 Ma) (du Bray, 2007).

Late Jurassic and Cretaceous compression formed an extensive fold and thrust belt further east in Utah and Wyoming during the Sevier orogeny. Flat-slab subduction of the Farallon plate underneath North America from the late Cretaceous to Eocene resulted in thick-skinned deformation and uplift of the Rocky Mountains from New Mexico to British Columbia during the Laramide orogeny. The second major mode of felsic plutonism occurred in Nevada during this time (~90–95 Ma) (du Bray, 2007), associated with porphyry-style base metal mineralization events.

As the Laramide orogeny waned into the Eocene, there was a major transition from compressional to extensional tectonic regimes in Nevada. Extensional tectonic stresses, evidenced by block faulting and titling, have dominated Nevada from the Late Eocene to the present. Three temporally distinct orientations of post-Cretaceous crustal extension have been identified: north-west–south-east in the Late Eocene to middle Miocene; west–south-west–east–north-east in the middle Miocene; and north-west–south-east in the late Miocene to present (Zoback et al., 1994). These extension events resulted in the development of basin and range physiography seen throughout central Nevada. The landform is characterized by a series of horsts and grabens that created narrow north–north-east oriented ranges separated by flat bottomed valleys. Extension and resultant crustal thinning are associated with the third major magmatic pulse in Nevada, during which time several porphyry copper–gold systems developed. In addition, the famous Carlin-type gold deposits (CTGD) of northern Nevada are thought to have formed during this time (~36–42 Ma) (Cline et al., 2005).

Magmatism of andesitic to rhyolitic affinity dominated from the Late Eocene to Early Miocene with the production of voluminous ash flowsheets, plutons, hypabyssal intrusives and calderas. Volcanic arc-related andesitic igneous activity continued in western Nevada from early to late Miocene. Further east in central and eastern Nevada, rift related bi-modal rhyolite and tholeiitic basalt were emplaced in the Mid Miocene and are related to epithermal silver–gold deposits in the region. A summary of significant geologic events of northern Nevada is presented in Figure 6.4.

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Figure 6.4 Major Igneous, Tectonic, and Mineralizing Events in Northern Nevada

Wallace et al., 2004

6.2 Local Geology

The Property is in the Battle Mountain mining district on the northern end of the Battle Mountain-Eureka trend, a conspicuous lineament of sedimentary-hosted gold deposits (Figure 6.5). The Battle Mountain district hosts numerous mineral occurrences, including porphyry copper–gold, porphyry copper–molybdenum, skarn, placer gold, distal disseminated silver-gold, and Carlin-type gold systems.

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Figure 6.5 Location of Marigold and the Battle Mountain Mining District on the Battle Mountain-Eureka Mineral Trend

Modified after Wallace et al., 2004

6.2.1 Stratigraphy

The Battle Mountain mining district is underlain by Paleozoic metasedimentary and metavolcanic rocks that are cut by Jurassic, Cretaceous, and Eocene intrusions. Post-mineralization tuff, volcanic rock, and detritus were deposited and preserved in structural and paleotopographic lows. The oldest rocks in the Battle Mountain mining district are para-autochthonous Cambro-Ordovician carbonate, clastic, and volcanic rocks in the footwall of the Roberts Mountain allochthon; assigned to the Comus-Preble Formation, (Cook, 2015). The Comus-Preble Formation comprises fine-grained siliciclastic turbidite sequences, mudstone, siltstone, limey mudstone, limestone, debris flows, and mafic volcanic flows.

Rocks of the Roberts Mountain allochthon were thrust eastward during the Devonian-Mississippian Antler orogeny. This event resulted in intense deformation, including folding and intra-formational thrusting of the metasedimentary units that comprise the Roberts Mountain allochthon. Rocks of the allochthonous clastic assemblage in the Battle Mountain district were previously separated into the Cambrian Scott Canyon Formation, Cambrian Harmony Formation, and the Ordovician Valmy Formation, complicating the understanding of Paleozoic tectonic processes affecting the district. Recent work by Ketner (2008; 2013) proposed the abandonment of the Scott Canyon Formation and reassignment of these rocks to the Valmy and Harmony Formations. Ketner (2008) demonstrated the Harmony Formation conformably overlies the Valmy Formation, eliminating the necessity for the Dewitt thrust mapped by Roberts (1964) and Theodore (1991).

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Unconformably overlying rocks of the clastic assemblage is the autochthonous Antler overlap sequence; a Pennsylvanian-Permian package of conglomerate, limestone, siltstone, and debris flow. Basal Antler sequence rocks were deposited as material eroded off the Antler highland into a foreland basin during the Antler orogeny. The base of the Antler sequence, the Battle Formation, is a coarse conglomerate up to approximately 220 m thick (Roberts, 1964) that contains clasts derived from the Roberts Mountain allochthon and underlying para-autochthonous rocks. The Battle Formation was deposited in a fluvial-to-shallow marine environment, with coarse, locally derived boulders at the base and interbedded limestone and siltstone units toward the top.

Disconformably overlying the Battle Formation is the Antler Peak Limestone Formation, a package of shallow marine carbonate rocks over 180 m thick at its type locality (Roberts, 1964). The Antler Peak Limestone Formation contains abundant brachiopod, coral, and crinoid fossils. The type of section for the Antler Peak Limestone Formation is in the Battle Mountain Range at Antler Peak.

The Permian Edna Mountain Formation disconformably overlies the Antler Peak Formation and consists of locally present basal debris flow and brown weathering phosphatic siltstone (McGibbon, 2005) at least 120 m thick. Unoxidized Edna Mountain Formation is black in color and difficult to differentiate from unoxidized siltstone of the Havallah sequence in drill cuttings and in the field.

Allochthonous rocks of the Mississippian-Permian Havallah sequence were tectonically emplaced over rocks of the Antler sequence, Valmy Formation, and Preble-Comus Formation during the Permo-Triassic Sonoma orogeny (Theodore, 2000; McGibbon, 2005). The Havallah sequence includes chert, siltstone, limestone, conglomerate, sandstone, and submarine volcanic rocks. The total thickness of the sequence is thought to exceed 2.8 km (Roberts, 1964).

6.2.2 Igneous Rocks

The oldest igneous rocks in the district are submarine pillow basalts within the Cambro-Ordovician Preble-Comus and Ordovician Valmy Formations.

Volcanic rocks within the Preble-Comus are only known from drill core and consist of submarine pillow basalt and volcaniclastic units derived from a continental source. These rocks are typically highly altered due to their age, submarine emplacement, present surface to near-surface position, and exposure to hydrothermal systems.

Metabasalt belonging to the Valmy Formation outcrops in the vicinity of Trout Creek south of the Oyarbide fault. On the east side of the district at Elder Creek, diorite dikes of Devonian age are inferred based on cross-cutting relationships. Mesozoic igneous rocks include a relatively unaltered Jurassic lamprophyric dike (Fithian, 2015) and an abundance of north-west striking Cretaceous granodiorite and quartz monzonite porphyry dikes and stocks.

Late Cretaceous granodiorite and quartz monzonite porphyry rocks are associated with molybdenum mineralizing systems at Buckingham, Trenton Canyon, and Buffalo Valley (Doebrich and Theodore, 1996).

Cenozoic igneous activity coincided with the onset of extensional tectonism throughout the Basin and Range province and normal reactivation of north and north-west striking faults in the Battle Mountain district (Doebrich and Theodore, 1996).

Late Eocene to Early Oligocene granodiorite to monzogranite intrusive stocks and dikes are associated with copper-gold mineralizing systems in the district, such as those at Converse and Copper Canyon. Intrusive dikes and sills are typically low relief slope forming units with very little outcrop in part due to argillic alteration where it has been exposed to hydrothermal fluids.

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Tertiary volcanic rocks in the district are post-mineralization. Oligocene to Miocene rhyolitic tuff and basaltic andesite flows are intercalated with Tertiary gravels and are locally ridge-forming units. The youngest volcanic rock, Pliocene (2.8–3.3 Ma) basalt, is present south-east of Copper Canyon (Doebrich and Theodore, 1996).

6.2.3 Regional Structure

Geophysical and isotopic evidence indicate that broad structural zones within the Battle Mountain-Eureka trend may be related to large-scale tectonic processes affecting the western margin of North America from the late Proterozoic through Mesozoic (Grauch et al., 2003). These features may be associated with deep crustal faults that originated as rift or transform faults during Proterozoic breakup of Rodinia, or as faults accommodating late Paleozoic compressional tectonic events (Grauch et al., 2003). Within the Battle Mountain-Eureka trend, deep crustal normal faults with a north-west, north, and north-east strike have influenced sedimentation, deformation, magmatism, extension, and mineralization (Grauch et al., 2003).

In the Battle Mountain mining district, the most prominent surface fault expressions are thrust faults related to Paleozoic-Mesozoic compressional tectonism, and normal faults related to Cenozoic extensional tectonic regimes. There is evidence of a more cryptic late Paleozoic transtensional fault system throughout the district, which is potentially late to post-Antler orogeny. These structures do not display significant slip in post-Permian aged rocks, and as a result are commonly concealed. Structures related to the transtensional fault system are responsible for preservation of thick wedges of Antler sequence rocks.

The Permo-Triassic Golconda thrust fault is traceable throughout the entire Battle Mountain range. Onset of the latest crustal extension began in the late Eocene and has continued sporadically to present. The most prominent extensional faults in the district are the range-bounding normal faults that define the Battle Mountain range, including the post-mineralization, south-west striking Oyarbide fault (Doebrich and Theodore, 1996).

At least four generations of folding are recorded in Ordovician rocks of the Roberts Mountain allochthon, including tight-to-isoclinal overturned F1 folds with north-west–south-east fold axes, open and upright F2 folds with west–north-west fold axes, large-scale open and upright F4 folds with north–north-east fold axes, and roll-over anticline style F5 folds that affect the entire rock package. Fold events F1 and F2 pre-date deposition of Antler sequence rocks. The F3 fold event is restricted to the Havallah sequence. F4 folds are thought to be related to Mesozoic tectonics and affect Comus-Preble Formation, Valmy Formation, Antler sequence, and Havallah sequence rocks while F5 folds appear to affect the entire rock package including Tertiary rocks.

6.3 Property Geology

6.3.1 Property Stratigraphy

The Property stratigraphy is summarized in Figure 6.6.

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Figure 6.6 Schematic Tectono-stratigraphic Section of the Rock Units at Marigold

SSR, 2021

Sedimentary Rocks

There are no Mesozoic sedimentary rocks in the Marigold mine area; however, approximately two thirds of the Property is covered by Tertiary to Quaternary intercalated gravel and volcanic material.

Comus-Preble Formation

The assignment of rocks to the Comus-Preble Formation at Marigold is the result of an extensive effort to explore the depths of the Marigold system. On the basis of lithology and deformation style, rocks believed to be positioned below the Roberts Mountain Thrust were assigned to the Comus-Preble Formation.

Valmy Formation

Fold deformation in the Valmy Formation is characterized by tight, east-vergent, and overturned folds. This fold deformation has resulted in shattering of quartzite beds and ductile deformation of argillite. Where the contact is not eroded or structurally displaced, the top of the Valmy Formation is unconformably overlain by rocks of Pennsylvanian age. Silurian and Devonian rocks are not present either due to nondeposition or erosion.

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Antler Sequence

The Antler overlap sequence is composed of Pennsylvanian to Permian-aged rocks assigned to three formations: the basal Battle Formation; the Antler Peak Limestone Formation; and the Edna Mountain Formation. These Formations represent a transgressive sequence of fluvial-to-shallow marine rocks that include conglomerate, sandstone, limestone, siltstone, and debris flows. There is evidence the Antler sequence was locally deposited into sub-basins developed by normal offset on growth faults of likely Late Pennsylvanian to Early Permian age.

Antler sequence rocks are relatively undeformed, except for offset and rotation along Basin and Range normal faults and potentially low-amplitude, long-wavelength (kilometres to tens of kilometres) F4 folding likely related to Mesozoic deformation. The Antler sequence is in thrust contact with the overlying and partially contemporaneous Havallah sequence.

Havallah Sequence

Figure 6.7 Plan View Map Showing Distribution of Paleozoic Units at Marigold.

SSR, 2018

Igneous Rocks

A 2 m interval of an extremely biotite-rich intrusive rock, interpreted to be lamprophyre, was intersected in a single drillhole approximately 1,100 m below the pre-mining topography. Even though the rock is relatively unaltered, the lamprophyre is Jurassic in age (160.7 ± 0.1 Ma Ar-Ar of biotite) (Fithian, 2018) and is age-equivalent to lamprophyre intrusions in northern Nevada.

A series of Late Cretaceous (~92.22 ± 0.05 Ma to 97.63 ± 0.05 Ma, CA-TIMS of zircon) (Fithian, 2015) porphyritic quartz-monzonite dikes crosscut the Paleozoic rock package at Marigold. The intrusions are up to tens of metres wide, and several can be traced along strike for

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hundreds of metres. The dikes strike south-east to north–south and are typically steeply dipping. No alteration aureole related to these intrusive rocks has been identified at Marigold (Fithian, 2015). The dikes contain phenocrysts of plagioclase feldspar, biotite, hornblende, and quartz. The mafic phenocrysts have all been altered to secondary mineral assemblages to varying degrees.

Oligocene (~31.8 ± 0.8, 31.4 ± 1.0 Ma) (Theodore, 2000) basaltic andesite is present on the Property, and forms a small, mesa-like landform between Trout and Cottonwood Creeks. The basaltic andesite is crudely columnar in this location.

Late Oligocene to Early Miocene (22.9 ± 0.7 Ma) (McKee, 2000) post-mineralization rhyolite tuff is intercalated with gravel throughout the Property. The tuff contains phenocrysts of biotite and is typically altered to white clay. The tuff provides a minimum age of mineralization at Marigold, as it is unmineralized and immediately overlies the orebody at the 8S deposit (Theodore, 2000; McGibbon and Wallace, 2000).

6.3.2 Property Structure

Valmy Formation rocks are highly deformed, with interpreted imbricate low-angle intra-plate thrust faults and at least two generations of pre-Pennsylvanian folding. The first generation of deformation related to folding of the Valmy Formation, D1, is characterized by tight, east verging folds with approximately north-west–south-east to north–south striking fold axes. The second deformation event, D2, is defined by open folds with approximately east–west striking fold axes. Folds of this orientation are best defined on the southernmost part of the property, including the Basalt pit area.

Although D1 and D2 folds are described individually because of their unique character, it is possible that these fold sets are the product of the same deformation event. The areas of confluence of D1 and D2 folds are thought to have played a role in the localization of mineralizing fluids.

Argillite beds within the Valmy Formation deformed plastically while brittle quartzite beds shattered, creating open fracture space amenable for precipitation of auriferous iron sulfides. Antler sequence rocks are cut by, and rotated along, Early Permian and Cenozoic normal faults. The timing of the proposed Early Permian growth faults is based on preservation of Battle Formation, Antler Limestone Formation, and a thicker wedge of Edna Mountain Formation in the hangingwall of east dipping normal faults, with little-to-no appreciable offset of the overlying Havallah sequence (Figure 6.16).

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Figure 6.8 The Top Surface of the Valmy Formation with the Current Property Boundary

SSR, 2021. Black lines indicate the position of major structures in the Valmy Formation (dashed where projected)

Figure 6.9 Cross Section 11,200N Highlighting Inferred Permian Growth Fault and Associated Antithetic Normal Faults with a Steep West Dip

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Rocks of the Antler sequence are deformed by F4 and F5 folds, which are not easily recognized in the field. Despite the position between two inferred major allochthonous packages, the Antler sequence does not display more-intense fold deformation akin to F1 and F2 folds.

Havallah sequence rocks were deformed by thrusting and folding related to compression during the Permo-Triassic Sonoma orogeny. An extensive series of thrust faults and folds are documented by Theodore (1991) in the Valmy and North Peak quadrangles west of the Marigold mine area.

Deformation of the Havallah sequence is apparently unrelated to gold mineralization at Marigold. Development of basin and range normal faults and reactivation of Paleozoic faults during the Cenozoic affected the entire stratigraphic section at Marigold, including displacement of post-mineralization Oligocene tuff and Quaternary gravel (Figure 6.10).

Figure 6.10 Normal Displacement of Alluvium and Tuff Immediately South of the Basalt Pit

View is towards the south

Fithian, 2015

6.3.3 Mineralization

The gold deposits at Marigold cumulatively define a north-trending alignment of gold mineralized rock more than 8 km long (Figure 6.11).

Gold mineralizing fluids were primarily controlled by fault structure and lithology, with tertiary influence by fold geometry. Within the Valmy Formation, higher gold grades are observed in the hinge zones of open folds that trend west–north-west and plunge gently. When viewed down plunge, the undulation of these folds is mimicked by gold mineralized horizons. The deposition of gold was restricted to fault zones and quartzite dominant horizons within the Valmy Formation and high permeability units within the Antler sequence.

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In unoxidized rocks, gold occurs in arsenic-enriched overgrowths on pre-ore pyrite (Figure 6.12). Arsenopyrite is also present on pre-ore pyrite grains but is not auriferous. Geochemically, the gold mineralization event is characterized by elevated arsenic, barium, antimony, and mercury, among others. Gangue minerals include quartz, arsenopyrite, stibnite, calcite, clay, and barite. Hypogene sulfide minerals do not occur in ore as these gold-bearing phases are not amenable to heap leaching.

In oxidized rocks, gold occurs natively in fractures associated with iron oxide (Figure 6.13). Rocks within the Marigold mine area are oxidized to a maximum depth of approximately 450 m. The redox boundary is not consistent throughout the property and is substantially influenced by lithology. Shale, argillite, and siltstone units are frequently unoxidized adjacent to pervasively oxidized quartzite horizons.

A silver and base metal mineralizing event at Marigold includes a mineral association of chalcopyrite, argentiferous tennantite, galena, and sphalerite. The absolute age of this event is unclear, although it may be related to late Cretaceous magmatism in the district.

6.3.4 Alteration

Alteration of rocks includes silicification along mineralizing structures and decalcification of carbonate horizons (primarily in the Antler sequence). Argillic alteration of quartz monzonite intrusive bodies occurs in fault zones and areas of high hydrothermal fluid flow (Fithian, 2015). The intensity of alteration decreases towards the core of the intrusions.

Studies have demonstrated a spatial correlation between gold mineralized rock and increased white mica crystallinity index (Kester, 2015). There is evidence for large volumes of quartz precipitation within and outboard of gold mineralized zones, including jasperoid bodies, cryptic silicification, and quartz vein breccias.

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Figure 6.11 Plan View of the Marigold Mine Area showing the Spatial Distribution of 1.0 g/t Au Grade Shells Over an 8 km Northerly Trend

SSR, 2018

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Figure 6.12 Gold in Arsenian Pyrite Overgrowths on Pyrite Grains in Unoxidized Rock

Arsenopyrite (white) does not contain gold

Modified from Fithian, 2015

Figure 6.13 Native Gold Occurs with Iron Oxide in Weathered Rocks

Modified from Fithian, 2015

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6.4 Deposit Geology

Gold at Marigold is currently mined from multiple deposits located on a 10 km by 1.5 km area.

From north to south, historical and future mineral deposits at Marigold include 32 North (32N), 5 Northeast (5NE), 5 North (5N), 8 North(8N), 8 Deep (8D), Terry Zone North (TZN), 8 South (8S), 8 South Extension (8Sx), Terry Zone (Old Marigold), Top Zone, HideOut, Terry Complex (Battle, Red Rock, East Hill), Red Dot, Mackay, Mud, Target, Valmy, Basalt-Antler, East Basalt, and Battle Cry. The majority of these individual mineralization zones have coalesced into the Mackay pit.

6.4.1 Mackay Pit

The Mackay pit contains most of Marigold’s current Mineral Resources. Gold is predominantly associated with iron oxide minerals on fracture surfaces of Valmy Formation quartzite, with lesser amounts of gold in Antler sequence rocks (Figure 6.14). Gold is concentrated within narrow structures with a steep west dip, and the intersection of these structures with favourable quartzite horizons within the Valmy Formation.

On the northern end of the planned Mackay pit, a greater percentage of the ore is hosted in Antler sequence rocks, including the deposits at HideOut (Figure 6.15), 8Sx, and 8N.

Where mineralized, Antler sequence rocks tend to host higher concentrations of gold, likely due to increased chemical reactivity with mineralizing fluids.

Figure 6.14 Cross Section 13,200N Highlighting Distribution of Gold in Antler Sequence and Valmy Formation Rocks

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Figure 6.15 Cross-Section 16,000N Highlighting the HideOut Deposit Hosted by Antler Sequence Rocks

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Figure 6.16 Plan Reference of Cross-Sections in Figure 6.9, Figure 6.14, and Figure 6.15

SSR, 2021

6.5 Deposit Type

Doebrich and Theodore (1996), Theodore (1998), and Theodore (2000) described the deposits at Marigold as distal disseminated silver–gold deposits. These deposits are disseminated equivalents of polymetallic vein deposits, characterized by a geochemical signature that includes silver, gold, lead, manganese, zinc, copper, antimony, arsenic, mercury, and tellurium (Cox and Singer, 1990). Typically, they contain substantially more silver relative to gold than other types of disseminated gold deposits and may feature supergene enrichment of silver if significantly oxidized.

In Nevada, distal disseminated silver–gold deposits are proximal to Jurassic, Cretaceous, and mid-Tertiary granitoid intrusions (Hofstra and Cline, 2000). A fundamental requirement of the distal disseminated silver–gold model necessitates a genetic link between silver–gold mineralization and causative intrusions (Figure 6.1) (Hofstra and Cline, 2000); however, no such relationship has been conclusively demonstrated at Marigold (Fithian, 2015).

A Carlin-type gold deposit (CTGD) is a unique type of disseminated, sedimentary rock-hosted gold deposit. The genesis of CTGDs is currently not well understood. In Nevada, CTGDs occur along several main mineralization trends, including the Carlin trend and Battle Mountain-Eureka trend, and are primarily hosted by silty carbonate rocks.

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Gold in a CTGD occurs in arsenian pyrite rims on pyrite grains and is associated with arsenic, sulfur, antimony, mercury, and thallium (Cline et al., 2005). There is considerable debate regarding the source of gold in CTGDs. Leading theories include a magmatic-hydrothermal origin (e.g., Sillitoe and Bonham, 1990; Johnston and Ressel, 2004; Ressel and Henry, 2006; Muntean et al., 2011) and gold sourced from the sedimentary host package (e.g., Ilchik and Barton, 1997; Emsbo et al., 2003; Large et al., 2011). Even though the genesis of CTGDs remains enigmatic, there is consensus that all CTGDs in Nevada formed during the Eocene period (42 to 36 Ma) (Cline et al., 2005).

Distal disseminated silver–gold deposits may share similarities with CTGDs, including orebody morphology, structural setting, and alteration styles, but drastically differ with respect to alteration zonation, geochemical signature, hypogene mineralogy, and endowment. Distal disseminated silver–gold deposits show a more definitive magmatic signature than CTGDs that includes zoning of alteration relative to felsic hypabyssal intrusions, base metal enrichment, significantly higher Ag:Au ratios, and distinctive hypogene ore mineralogy (e.g., base metal sulfides, native gold and silver, electrum, silver sulfides and silver sulfosalts) (Cox and Singer, 1990; Cox, 1992; Hofstra and Cline, 2000), and are typically much smaller in terms of gold endowment.

There is increasing support for a model that proposes a continuum between CTGDs, distal disseminated silver–gold deposits, and epithermal deposits. This model implies a magmatic source for heat and metal. Those most familiar with the Marigold system support a model invoking an intrusive metal and heat source, despite a lack of definitively magmatic features. The expanded Marigold property boundary enables study of the Marigold system on a considerably broader scale and may enable recognition of large-scale alteration zonation.

Recent work by Fithian (2015) suggests that the gold deposits at Marigold are best classified as CTGDs, based on many similarities with the CTGD model and a lack of evidence for causative hypabyssal intrusions.

Figure 6.17 is a diagrammatic representation of the deposit model.

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Figure 6.17 Model Illustrating Inferred Processes Related to Formation of Carlin-Type Gold Deposits (CTGD) and Distal Disseminated Silver–Gold Deposits

SSR, 2021

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7 EXPLORATION

For a discussion regarding historical exploration programs completed following SSR’s acquisition of the Property in April 2014, refer to Section 5.

7.1 Gravity Surveys

7.1.1 Gravity Survey Pre-2015

After the purchase of Marigold was completed in 2014, SSR completed a gravity survey at a grid spacing of 150 m x 150 m in areas that had not been previously covered. The main objective of this work was to delineate possible fluid conduits or feeder structures for the Marigold mineralization.

The gravity survey was planned and designed by James L. Wright of J L Wright Geophysics, Spring Creek, Nevada. The gravity survey and field data collection were conducted by Magee Geophysical Services LLC of Reno, Nevada.

The gravity measurements were collected from 1,358 stations using two LaCoste & Romberg Model-G gravity meters. Forty planned stations were skipped due to active mining and/or unsafe ground conditions. Figure 7.1 shows the actual station locations from the gravity survey. Topographic measurements were also collected at each station using the RTK GPS method. Where it was not possible to receive GPS-based information via a radio modem, the Fast-Static (post-processing) GPS method was used.

7.1.2 Gravity Survey Post-2015

After finalizing the purchase of Valmy in 2015 (additional Newmont owned land to the east and west of the previous land boundary), SSR expanded the geophysical gravity survey to include this new ground.

The gravity survey was conducted by Magee Geophysical Services in August and September of 2016. The main objective of this work was to extend the detailed coverage of three previous gravity surveys in the vicinity of the Marigold mine.

Relative gravity measurements were made with LaCoste & Romberg Model-G gravity meters. Topographic surveying was performed with Trimble RTK and Fast-Static GPS methods. Gravity measurements were processed to complete Bouguer gravity, merged with existing data, and forwarded to J L Wright Geophysics for further processing and interpretation.

7.1.3 Gravity Stations

In 2016, a total of 1,806 new gravity stations were acquired by Magee Geophysical Services at variable station spacing on a 150 m square grid and a 150 m x 300 m staggered grid. Existing gravity data included 1,358 stations collected in 2014 by Magee Geophysical Services, 1,250 stations collected in 1998 by Zonge International Inc. (Zonge), and 122 stations collected on various dates by Newmont. Additional stations, including repeats, totalled 4,853 stations. Figure 7.2 shows a complete station posting, color-coded by survey date.

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Figure 7.1 Marigold Mine Gravity Survey Stations in 2014

Marigold mine gravity survey stations in 2014 are shown in red over as-mined topography

Magee Geophysical Services, 2014

7.1.4 Terrain Corrections

Terrain corrections were calculated to a distance of 167 km for each gravity station. The terrain correction for the distance of 0–5 m around each station used a sloped triangle method with the average slopes measured in the field. The terrain correction for the distance of 5–2,000 m around each station used a prism method and a sectional ring method with digital terrain from a 5 m digital elevation model (DEM). The 5 m DEM was prepared by merging a 2016 proprietary Marigold DEM with surrounding United States Geological Survey (USGS) 10 m DEMs. The Marigold proprietary elevation data were assumed to be in NGVD 29; some minor edits were made to remove artificial terrain prior to merging with USGS data.

The terrain correction for the distance of 2–167 km around each station used the sectional ring method with digital terrain from shuttle radar topography mission (SRTM) DEM and/or a 90 m DEM.

Terrain corrections for existing data were performed using the same procedures, but with local terrain derived from a 2014 proprietary 5 m Marigold DEM.

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Figure 7.2 Gravity Stations

Stations: Zonge 1998 (●), Magee 2014 (●), Magee 2016 (●), and USGS (●)

James L. Wright, 2016

7.1.5 Interpretation

The complete Bouguer anomaly at 2.55 grams per cubic centimetre (g/cm3) shows a clear north-east–south-west trending feature that corresponds to the Oyarbide Fault cutting the survey’s south-east corner. Dense rocks lie to the south-east of the fault relative to those in the north-west. However, both rock units are mapped as Valmy Formation. A gravity high to the north-east is attributed to carbonate rocks beneath the valley fill. North–south structures extend directly along the middle of the gravity coverage, and gravity lows along the south-west edge are produced by basin fill in the head of Buffalo Valley (Figure 7.3).

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Figure 7.3 Marigold Mine Gravity Survey Compilation, Complete Bouguer Anomaly Oblique Image

James L. Wright, 2016

7.2 Exploration Drilling

The QP is of the opinion that the drilling and sampling procedures adopted at Marigold are consistent with generally recognized industry best practices. The resultant drilling pattern is sufficiently dense to interpret the geometry and the boundaries of gold mineralization with confidence. The reverse circulation (RC) samples were collected by competent personnel using procedures meeting generally accepted industry best practices. The process was conducted or supervized by suitably qualified geologists. The QPs are of the opinion that the samples are representative of the source materials, and there is no evidence that the sampling process introduced a bias. Accordingly, there are no known sampling or recovery factors that could materially impact the accuracy and reliability of drilling results.

As at the end of November 2021, 9,323 drillholes for 1,940,438 m of drilling comprise the current resource database for the Property.

Table 7.1 summarises all of the drilling on the Property from 1968 through 2021.

7.2.1 Exploration Drilling at Marigold (Pre-2014)

For details on drilling activities conducted at Marigold prior to 2014, refer to SSR’s NI 43-101 Technical Report on the Marigold Mine (19 November 2014).

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7.2.2 Exploration Drilling at Marigold (2014–2017)

Shortly after SSR’s acquisition of the Project was complete, an exploration program was initiated with a view to delineating additional Mineral Resources. The program commenced in June 2014 and targeted the discovery of near surface gold mineralization proximal to Marigold’s open pits and had the result of upgrading the Inferred Mineral Resources to Indicated Mineral Resources.

The 2014 to 2017 drilling included:

•706 reverse circulation (RC) drillholes for 170,684 m;

•37 sonic drillholes in rock stockpiles (included in RC totals)

•7 HQ diamond core holes for 7,588 m.

SSR drilled a total of 713 drillholes for 178,272 m from 2014 to 2017.

7.2.3 Exploration Drilling at Marigold (2018–2021)

From 2018 through to the end of December 2021, a further 995 holes have been drilled. This era of drilling included:

•950 RC holes

•45 diamond core holes.

The 2018–2021 drilling adds a further 343,233 m of drilling to the database.

This brings the total drilling in the history of the Marigold project to 9,435 drillholes for 1,988,280 m.

SSR’s drilling has been directed at various targets and resource areas including East Basalt, Battle Cry, Showdown, Valmy SE, Mud & NW, Crossfire, HideOut, 8Sx, TZN, 8D, 5N, Red Dot, North Red Dot, Mackay pit extensions, and the Mackay Herco Keel structure. These areas are shown in Figure 5.2.

Since 2018, the focus of exploration at Marigold has been:

•Exploration drilling to expand Mineral Resources and Mineral Reserves through systematic step out drilling.

•Drilling of 21 core holes to confirm the grades below water table that were originally obtained from RC drilling in the Red Dot area

•Infill drilling to increase the confidence of Mineral Resource estimates specifically targeting areas widely spaced drilling ~35–50m and around drillholes drilled prior to 2006 with missing assays.

•Drilling to confirm the final position of the pit highwall.

•Advancing drilling to define orebody at Trenton Canyon.

7.2.4 Marigold Sulfide Drilling Program

SSR has undertaken a drilling program to test sulfide mineralization at Marigold. To date, nine diamond drillholes have been drilled to test for sulfide mineralization.

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This drilling has been completed across the Property to help understand the overall geology of the Property and to target higher gold grades beyond the oxidation boundary that is currently mined at Marigold.

7.2.5 Exploration Drilling at Valmy (1968–2006)

In 2015, SSR purchased the Valmy property from Newmont, and all previous drilling information for Valmy was incorporated into the Marigold drilling database.

Numerous companies explored the Valmy property from 1968 until Newmont put the Valmy and Mud pits into operation in 2002. These companies included Hecla, Santa Fe Pacific Minerals Limited, and Newmont. As mentioned, this drilling data has been reviewed closely by SSR.

7.2.6 Exploration – Trenton Canyon and Buffalo Valley

Exploration work on the Trenton Canyon and Buffalo Valley properties consists of drilling, geophysical surveying, remote sensing, geochemical surveying, and mapping.

Gold mineralization at Trenton Canyon is structurally controlled with significantly less dissemination than at Marigold. The net result of this change in mineralization style is higher gold grades in a smaller volume of mineralized rock at Trenton Canyon.

Figure 7.4 shows a plan view of the area and extent of the work completed on Trenton Canyon and Buffalo Valley.

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Figure 7.4 Plan View of Drilling Carried out on Trenton Canyon and Buffalo Valley

SSR, 2021

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The drilling completed since SSR’s acquisition of the Project in 2014 is summarized in Figure 7.5 and shown in plan view in Figure 7.6.

Figure 7.5 Chart of Drilling Completed by SSR since 2014

SSR, 2021

Table 7.1 summarises all of the drilling on the entire Marigold property from 1968 through 2021. Figure 7.7 shows the same in plan view.

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Figure 7.6 Plan View of Drilling Carried out by SSR Since 2014

SSR, 2017

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Table 7.1 Summary of Drilling History

Drilling Program	Company	No. of RC Holes	RC Drilling (m)(1)	No. of Diamond Holes	Diamond Drilling (m)(1)	Total Holes	Total Drilling (m)(1)
1968–1985	Various exploration and mining groups	126(2)	7,037(2)	(2)	(2)	126	7,037
1985–1999	Cordex and Rayrock Mines	2,350	333,325	8	2,176	2,358	335,501
1999–2006	Glamis Gold	2,498	484,619	8	2,030	2,506	486,649
2006–2013	Goldcorp	1,856	520,163	14	8,063	1,870	528,226
1968–2006	Newmont and other mining groups (Valmy property)	852	108,326	15	1,037	867	109,363
2014	SSR	116	21,653	1(3)	1,235(3)	117	22,888
2015	SSR	171(5)	39,070	4	4,270(4)	175(5)	43,340(5)
2016	SSR	231	55,147	1	955	232	56,102
2017	SSR	188	54,814	1	1,128	189	55,942
2018	SSR (Marigold)	259	93,276	0	0	259	93,276
2019	SSR (Marigold)	183	63,629	25	10,265	208	73,893
2020	SSR (Marigold)	109	37,955	0	0	109	37,955
20216	SSR (Marigold)	150	52,579	6	1,636	156	52,214
2019	SSR (TCBV)	64	19,112	0	0	64	19,112
2020	SSR (TCBV)	98	28,840	7	5,901	104	34,742
20216	SSR (TCBV)	88	25,213	7	4,827	95	30,040
Total Drilling		9,338	1,944,758	97	43,523	9,435	1,988,280

1. Drill lengths converted from feet to metres.

2. Figures have been rounded and may not match totals.

3. No documentation of drilling method at Marigold is available for these drillholes. However, before RC drilling became widely adopted in the mid-1980s, conventional single-tube drilling was often relied on as the exploration drilling technique. It is suspected that single tube drilling was used during this time period; only occasional diamond drillholes were used. These drillholes are located in areas that have been mined or are outside of the current Mineral Resource area of Marigold.

4. Historical drillholes completed by Newmont at the Trenton Canyon and Buffalo Valley properties are not included in this table as they are currently being validated

5. Drilling to end of December 2021.

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Figure 7.7 Plan View of All Drilling to End of November 2021

SSR, 2021

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8 SAMPLE PREPARATION, ANALYSES, AND SECURITY

Exploration activities conducted by three companies between 1985 and 2013 have contributed to most of the assays in the Marigold database. Sampling and analytical procedures for this period are known and documented, and it can be assumed that analytical information acquired prior to 1985 will not impact the current Mineral Resources because sampled volumes collected prior to 1985 have been mined out.

Most of the samples that inform the resource database were generated from RC drill cuttings. In general, the process for collecting RC samples has changed very little since 1985; however, over time, there have been numerous improvements in sample preparation, security and analysis. As an operating mine, Marigold generally followed and continues to follow industry best practice standards.

At the Property, there is an extensive sample storage facility that preserves the raw sample material that supports the resource database. Most of the laboratory pulp reject (since 1987), coarse reject (since 2006), and split diamond drill core are catalogued and stored securely in shipping containers on the Property.

A detailed account of the pre-2014 sampling and analytical protocols is described in the NI 43-101 Technical Report on the Marigold Mine (19 November 2014). This section briefly describes historical procedures and reviews the current procedures and results that support the QC of data collected since such last technical report.

8.1 Sample Preparation and Analysis

A summary of historical analytical methods and assay results that comprise the Marigold and Valmy database is presented in Table 8.1. Except for the Marigold, Pinson and Dee Mine site laboratories, all laboratories listed in Table 8.1 are commercial laboratories that were independent from SSR.

Until the end of 1999, fire assay (FA) with gravimetric finish was the preferred analytical method for determining gold in samples. Since then, all samples have been subjected to first-pass gold cyanide solution assay, and, if results were greater than 0.17 g/t Au, samples were also subjected to FA determination with gravimetric finish at the on-site Marigold mine laboratory or FA with atomic absorption (AA) finish and FA with gravimetric finish for over limits at commercial laboratories.

All the Newmont-provided samples that inform the resource database for the Valmy area were assayed at various commercial laboratories. The preferred assay method was FA with AA spectroscopy finish, followed by gold cyanide solution assay on select samples within the mineralized zone.

Since 2014, all exploration samples from Marigold and the Valmy property are analysed at American Assay Laboratories (AAL), an ISO 17025 certified facility in Sparks, Nevada. AAL is independent from SSR. All samples are subjected to first pass FA determination with an AA finish and FA with gravimetric finish for over-limits. This is followed by a gold cyanide solution assay with an AA finish on samples that have FA values greater than or equal to 0.03 g/t Au. In 2019 and 2020 samples were also analysed at Paragon laboratories, a privately held corporation located in Sparks, NV. Paragon is independent of SSR. Analytical protocols similar to AAL were utilized.

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Table 8.1 Analytical Methods for Gold for the Marigold Assay Resource Database

Period		Laboratory		Preparation		Analytical Method		Reported DL (Au g/t)
1985–1989		Pinson or Dee Mine site labs		Undocumented		30 g FA, gravimetric finish		0.17
1990–1999		Pinson or Dee Mine site labs or Inspectorate Labs		Undocumented		30 g FA, gravimetric finish		0.17
1987–1998 (Newmont property)		Barringer Laboratories		Undocumented		30 g FA, AA finish 15 g cyanide gold (CN) assay on select samples		FA: 0.17 <br>CN assay: 0.17
	X-Ray Assay Laboratories		Undocumented		30 g FA, gravimetric finish 15 g CN assay on select samples		FA: 0.03 <br>CN assay: 0.03
	Rocky Mountain Geochemical Nevada		Undocumented		30 g FA, gravimetric finish 15 g CN assay on select samples		FA (AA): 0.03–0.003 CN assay: 0.03
	Chemex Labs Ltd.		Undocumented		15 g FA, AA finish 30 g FA, gravimetric finish 15 g CN assay on select samples		FA (AA): 0.06–0.003 CN assay: 0.03
2000–2004 (Newmont property)		Chemex Labs Ltd.		Dry, crush and riffle split for pulverizing; pulverize to 100µ		All samples 30 g FA, AA finish 15 g CN assay on select samples		FA (AA): 0.01 <br>CN assay: 0.03
2000–2006		Marigold Mine laboratory		Dry 6–12 hrs @ 310° F; crush >95% –2 mm; riffle split to collect 250–400 g for pulverizing; pulverize to >90% –75µ		All samples 10 g CN assay, AA finish If CN assay >0.17 g/t, the 2nd pulp split @ 30 g FA, gravimetric finish		0.03
	American Assay or Inspectorate Labs		Dry 6–12 hrs @ 310° F; crush (using jaw and roll) >90% –2 mm; riffle split to collect 500–1,000 g for pulverizing; pulverize to >90% –100µ		All samples 15 g CN assay, AA finish If CN assay >0.17 g/t, the 2nd pulp split @ 30 g FA, AA finish over-limits by 30 g FA, gravimetric finish		0.03
2006–2013		Marigold Mine laboratory		Dry 6–12 hrs @ 310° F; crush >95% –2 mm; riffle split to collect 250–400 g for pulverizing; pulverize to >90% –75µ		All samples 10 g CN assay, AA finish If CN assay >0.17 g/t, the 2nd pulp split @ 30 g FA, gravimetric finish		0.03
	American Assay or Inspectorate Labs		Dry 6–12 hrs @ 310° F; crush (using jaw and roll) >90% –2 mm; riffle split to collect 500–1,000 g for pulverizing; pulverize to >90% –100µ		All samples 15 g CN assay, AA finish If CN assay >0.17 g/t the 2nd pulp split @ 30 g FA, AA finish over-limits by 30 g FA, gravimetric finish		0.03

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2014–2021	American Assay Laboratories	Dry 6–12 hrs @ 310° F; crush (using jaw and roll) >90% –2 mm; riffle split to collect 500–1,000 g for pulverizing; pulverize to >90% –100µ	All samples 30 g FA, AA finish over-limits by 30 g FA, gravimetric finish If FA >0.03 g/t, the 2nd pulp split @ 15 g CN assay, AA finish	FA: 0.003 <br>CN assay: 0.03
2019-2020	Paragon Laboratories	Dry – 6 to 12 hrs @ 310° F; crush (using jaw and roll) >90% minus 2 mm; riffle split to collect 500 to 1,000 g for pulverizing; pulverize to >85% minus 75µ	All samples 30 g FA, AA finish Over-limits by 30 g FA, gravimetric finish If FA >0.03 g/t, the 2nd pulp split @ 15 g CN assay, AA finish	FA, 0.003 CN assay, 0.03

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8.2 Sample Security

8.2.1 Sample Security until 2013

The bulk of the data in the Marigold resource assay database was for samples analysed at the secure on-site Marigold mine laboratory. Samples shipped off site were either delivered to the commercial lab by an MMC Exploration Department geologist or technician, or samples were collected from the mine by a laboratory employee. All samples were sent with a manifest listing the number of samples included in the shipment. Exploration personnel were unaware of any instances of tampering with samples either on site or in transit to a laboratory.

8.2.2 Sample Security Valmy Property

Newmont provided scanned copies of driller’s logs, sample manifest sheets, and signed assay sheets from commercial laboratories and geologist logging sheets for all the drillholes that inform the resource database for the Valmy property. Based on the documented evidence, the chances of tampering with the samples either on site or in transit were negligible.

8.2.3 Sample Security 2014–2021

All exploration samples were collected from the mine site by an employee of AAL. All sample dispatches included a manifest listing the sample identifiers and number of samples included in the shipment. AAL/Paragon Laboratories electronically acknowledged the receipt of the samples within 24 hours after physically reconciling the samples with the manifest. SSR exploration personnel were unaware of any instances of tampering with samples either on site or in transit to a laboratory.

8.3 Quality Assurance/Quality Control (QA/QC) Procedures

8.3.1 QA/QC Procedures Pre-2014

The oldest hole in the Marigold exploration database is from 1968. Over time, QA procedures for the drillhole database have been inconsistent with current industry best practices.

Because the historical QA/QC procedures at Marigold did not meet current-day best practices, SSR selected a spatial and temporal representation of samples from the well-preserved drillhole sample pulps (from the years 1987 to 2013) stored at Marigold. SSR sent these to a commercial laboratory for analyses. The results of this re-assay program were discussed in the 2014 NI 43-101 Technical Report on the Marigold Mine (19 November 2014), and it was concluded that there was no systematic error or bias in the accuracy and precision of analytical assays from the period between 1987 and 2013.

As a part of the QA/QC program, a total of 1,974 samples were assayed for FA with AA finish and gravimetric finish between 1987 and 2003. Of these assay pairs, 1,029 samples were below the as-mined topography and within the mineralized envelopes. This represents 12% of samples that are within the mineralized envelope and below the mined-out topography. The assay results for both the finishes were compared, and results are presented in Figure 8.1 and Figure 8.2.

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Figure 8.1 Scatter Plot Between FA Gold Values with AA Finish and Gravimetric Finish

SSR, 2018

The scatter shown in the data presented in Figure 8.1 and Figure 8.2 is acceptable (R2 = 0.9982), and the reduced major axis (RMA) regression indicates a bias of 3.7% for all the assay pairs that are below the mined-out topography. These indicate that the assays form similar distributions and can be interchanged, but they do not validate the accuracy or precision of the assay value.

Figure 8.2 Q-Q Plot between FA Gold Values with AA Finish and Gravimetric Finish

SSR, 2018

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8.3.2 QA/QC Procedures Valmy Property

As at Marigold, the QA/QC procedures followed between 1987 and 1998 did not meet the current day industry standards. Newmont began inserting certified standards in the sample stream in 2000. A total of three QC samples were used, but SSR was unable to evaluate the assay accuracy without the expected gold values for these samples.

Because the historical QA/QC procedures for the Valmy property did not meet current day industry standards, SSR drilled eight drillholes within a resource block of 200 m x 150 m. A total of 11 historical drillholes were within the same block. The cross section comparing the SSR drilling to the historical drilling is presented in Figure 8.3.

The cumulative normal distribution comparing the SSR drill composites to the composite from the historical drillholes is provided in Figure 8.4.

Figure 8.3 Cross-Section with SSR Drillholes (drillhole number prefix MRA) and Historical Drillholes Along Section 8000N

SSR, 2018

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Figure 8.4 Cumulative Normal Distribution Comparing Composites from SSR Drilling and Historical Drilling

SSR, 2018

The nearest neighbour (NN) gold grade model estimates were also compared to the assay results from historical drilling and the new drilling. To compare historical Newmont data to SSR data, two NN models were developed: one estimate used only assay results from the historical database; and a second estimate used only the assay results from the SSR drillholes within the same mineralized envelope. The percentage difference between historical and SSR results was –4%. The results of the NN estimates are presented in Table 8.2.

Table 8.2 Comparison of the Nearest Neighbour Mean Gold Grades

Estimate	Mean Gold Grade<br>(g/t)
Nearest Neighbour with Historical Composites	0.624
Nearest Neighbour with SSR Composites	0.600

% Difference (SSR-Historical) is –4%.

The infill drill comparison indicates that there is no systematic error in the historical sampling and assaying methodology when compared to current practices, and, therefore, the historical data can be used to develop the Mineral Resources for the Valmy property.

8.3.3 QA/QC Procedures 2014–2017

SSR’s QA/QC protocol involves the insertion of a certified standards every 20th sample and the insertion of a blank sample every 50th sample. Eleven different certified standards purchased from ROCKLABS and Geo Chem Laboratories were used. In addition to the certified standards and blank material, every 50th sample is sampled in duplicate at the drill site and analysed as a field duplicate.

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8.3.3.1 Blanks

Coarse blanks are samples of barren material that are used to detect possible contamination, which is most common during the sample preparation stage. The size of the blanks was similar to the size of the RC samples, and they were processed through the same crushing and pulverizing stages as the drill samples. The blank samples were placed one in every 50 samples. Blank results that were greater than 10 times the lower detection limit (LDL) were typically considered failures that required further investigation and possible re-assaying of associated drill samples. The lower detection limit of AAL analyses is 0.003 g/t, so blank samples assaying in excess of 0.03 g/t were considered to be failures.

Between 2014 and 2017, a total of 1,107 blanks were inserted into the sample stream. The results are shown in the Figure 8.5. An assay value greater than five times the LDL is recorded as a warning, and ten times the LDL is deemed a failure limit. Four samples failed (0.36%), but only two samples were significant enough with assay values of 0.068 g/t.

Figure 8.5 Blank Results

SSR, 2018

Certified Standards

Certified reference material (CRM) standards were used to evaluate the analytical accuracy and precision of AAL. CRMs were inserted every 20th sample, which represents 5% of the total samples submitted. Three different CRMs were used in any one submission. The CRMs were selected based on the cut-off grade and gold distribution at Marigold mine, being:

•cut-off grade (0.1 g/t)

•mean grade (0.45 g/t)

•90th percentile (2.3 g/t)

Most of the CRMs used were purchased from ROCKLABS, and Ore Research & Exploration Pty Ltd. CRMs were only used in 2014 for a short period of time. The CRMs were assigned sample numbers in sequence with their accompanying drill samples and inserted into the drill-sample stream. The list of CRMs used between 2014 and 2017 is shown in Table 8.3.

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Exploration personnel monitor the assay results on a real-time basis and import the data into the Geology database. Internal validation checks in the database highlight any certified standard assay failures. In the case of normally distributed data, 95% of the standard assay results are expected to lie within two standard-deviation limits of the certified value. All samples outside the three standard-deviation limits were considered to be failures. Failures trigger a re-run of five samples above and five samples below the failed standards, including the failed standard.

Table 8.3 List of CRM Standards used between 2014 and 2017

CRM Standard	Expected Gold Value (g/t)	Standard Deviation (g/t)	No. of Samples Assayed
OxD108	0.414	0.012	480
OxJ95	2.337	0.057	361
OxB130	0.125	0.006	1137
OxJ111	2.166	0.058	131
OxJ120	2.365	0.063	627
OxD128	0.424	0.011	758
OREAS 50P	0.727	0.041	37
OREAS 50Pb	0.841	0.031	89
OREAS 6Pb	1.425	0.077	66
OREAS 7Pb	2.770	0.055	13
G312-7	0.220	0.010	111

Field Duplicates

Field duplicate samples were collected every 50th sample, and two sample bags marked “A” or “B” were provided to collect an original and a duplicate sample. The secondary sample was obtained from the secondary opening in the rotary sampler. The duplicate sample inserted into the sample stream monitors the precision of the sample collection, crushing, and pulverizing stages of sample preparation as well as the analytical stage.

Between 2014 and 2017, 1,650 duplicate samples were collected. Absolute relative difference (ARD) was used to estimate precision, as shown in Figure 8.6. Precision was estimated for all the samples to be at ±31%. Because most samples were below the 0.1 g/t grade used to construct mineralized envelopes, precision was also estimated for samples greater than 30 times the LDL. It was 25%. The estimated precision is considered to be reasonable for coarse field duplicates in gold deposits.

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Figure 8.6 Cumulative Frequency Distribution Comparing Original and Duplicate (field) Assay Results

SSR, 2018

8.3.4 QA/QC Procedures 2018–2021

SSR’s QA/QC protocol involves the insertion of a certified reference material standards (CRM) every 20th sample and the insertion of a blank sample every 50th sample. Eleven different CRMs, purchased from ROCKLABS and Geo Chem Laboratories, were used. In addition to the CRMs and blank material, a field duplicate is taken at the drill site for every 50th sample.

8.3.4.1 Blanks

Coarse blanks are samples of barren material that are used to detect possible contamination, which is most common during the sample preparation stage. The size of the blanks was similar to the size of the RC samples, and they were processed through the same crushing and pulverizing stages as the drill samples. The blank samples were placed one in every 50 samples. Blank results that were greater than 10 times the lower detection limit (LDL) were typically considered failures that required further investigation and possible re-assaying of associated drill samples. The lower detection limit of AAL analyses is 0.003 g/t, therefore blank samples assaying in excess of 0.03 g/t were considered to be failures.

Between 2018 and 2021, a total of 1,609 blanks were inserted into the sample stream. The results are shown in the Figure 8.5. An assay value greater than five times the LDL is recorded as a warning, and ten times the LDL is deemed a failure limit.

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Figure 8.7 Blank Results

SSR, 2021

Certified Standards

•around the cut-off grade (0.1 g/t)

•the mean grade (0.45 g/t)

•around 90th percentile (2.3 g/t) or greater

Most of the CRMs used were purchased from ROCKLABS. Ore Research & Exploration Pty Ltd. CRMs were only used in 2014 for a short period of time. The CRMs were assigned sample numbers in sequence with their accompanying drill samples and inserted into the sample stream. The list of CRMs used between 2014 and 2017 is shown in Table 8.3.

Exploration personnel monitor the assay results on a real-time basis and import the data into the geology database. Internal validation checks in the database highlight any CRM assay failures. In the case of normally distributed data, 95% of the CRM assay results are expected to lie within two standard-deviation limits of the certified value. All samples outside the three standard-deviation limits were considered to be failures. Failures trigger a re-run of five samples above and five samples below the failed CRM, including the failed standard.

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Table 8.4 List of CRMs used between 2018 and 2021

CRM Standard	Expected Au Value (g/t)	Standard Deviation (Au g/t)	No. of Samples Assayed
HiSilK2	3.474	0.087	241
OxJ120	2.365	0.063	648
OxB130	0.0125	0.006	1602
OxJ137	2.416	0.069	457
OxD151	0.430	0.009	894
SG84	1.026	0.025	72
OxD144	0.414	0.11	530

Field Duplicates

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Figure 8.8 Cumulative Frequency Distribution Comparing Original and Duplicate (Field) Assay Results

SSR, 2021

8.4 Conclusions and Recommendations

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9 DATA VERIFICATION

The verification for the exploration data collected before SSR acquired Marigold is described in the 2014 NI 43-101 Technical Report on the Marigold Mine (19 November 2014). It includes the results of AMEC Americas Ltd.’s external review and data verification to identify any material issues with the database used to generate the mineral resources.

SSR subsequently acquired the adjacent Valmy property, and the associated data was appended to the Marigold drillhole database.

The appended data for Valmy comprises collar, downhole survey, lithology, and assay information (provided in comma delimited digital files) for 867 drillholes drilled by Newmont, Hecla and Santa Fe Pacific Corp. Newmont provided this information in hardcopy or scanned versions of the originals which were used to verify the database.

MMC’s exploration personnel manually checked the entire drillhole database against the original documents for data entry errors. Less than 1% of the drillholes had any issues, and these were subsequently corrected.

As an additional check, SSR acquired the chip trays for 687 drillholes, pulps from 57 drillholes, and sample rejects from 66 drillholes, of which 5% were reviewed for lithology and alteration. The original logging was deemed accurate and was used to construct the lithological models.

The collar positions of 43 Valmy drillholes were verified using differentially corrected GPS methods. The results showed a maximum variance of 4 m in the X/Y planes (easting and northing) and < 1 m in the Z dimension (elevation). This error-shift is less than half the size of a resource model cell and is not material to any resulting estimate. The Valmy data, as appended, was deemed accurate and precise, and appropriate for resource estimation purposes.

For data collected after April 2014, the following verification steps were completed as part of the generation of the Mineral Resources estimate presented in the Marigold21TRS:

•The location of planned drillholes was compared to the location of as-built drillholes in real time. Regular field checks were completed on drill and sampling systems.

•Downhole survey intervals that encountered major deviations were reviewed and validated (AMEC, 2014).

•Precision and accuracy of laboratory assay results were verified using a QA/QC program that followed an industry standard protocol using the blind insertion of blanks and certified standards.

•The elevation of all surveyed drillhole collar coordinates was checked against the original/current/depleted topographic surface to identify any variations of more than one metre. No discrepancies were found.

•Profiles of all mined-out pits, backfilled pits and dumps were cross checked, updated annually, and incorporated into the current topography.

•All data, including collars, downhole survey, assays, and lithology, were imported directly into the geological database without any keyboard input. Data validation was conducted before the records were uploaded to the main database.

Three technical issues were identified in the Marigold Mineral Resources database (these issues have since been resolved):

•Drillholes were missing downhole surveys.

•Some samples were only assayed by cyanide soluble analysis and not by FA.

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•Assay results for a high percentage of lower grade samples were recorded as 0.0 oz/st gold.

The first two items were described and resolved in the 2014 NI 43-101 Technical Report on the Marigold Mine (19 November 2014).

The third item is described and resolved in Section 9.1.

9.1 Marigold Assay Database

As described in the 2014 NI 43-101 Technical Report on the Marigold Mine (19 November 2014), there have been changes in the lower detection limit for cyanide soluble gold assays over time as the ROM cut-off grade has been reduced. Prior to 2009, assay values below detection were entered into the database as 0.0 oz/t. This data artefact was under-representing the mineralized volume of the Mineral Resources estimate at the low-grade range of the analytical distribution and contributing to the positive reconciliation experienced at Marigold.

The issue of below-detection-limit analyses in the database was addressed through a systematic assay program implemented in 2015 and 2016 (the Assay Program). A total of 153,023 pulp samples from pre-2009 drillholes reporting a 0.0 oz/st gold cyanide soluble result and located within the reserve pits were recovered from storage and analysed for gold at AAL. Certified standards and blanks were inserted into the pulp sample list at a rate of one standard in 20 samples and one blank in 50 samples. The samples were analysed using a 1 assay ton (30 g) FA with an AA finish, followed by a gold cyanide solution assay with an AA finish for those samples that returned FA results of 0.03 g/t or greater.

The Assay Program identified additional mineralized areas, and the incorporation of this lower grade material that had been previously estimated as 0.0 oz/st or deemed as waste, increased the ore tonnage as shown in Figure 9.1.

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Figure 9.1 Cross Section Showing the Increase in Tonnage Estimated as Mineralized

SSR, 2018

9.2 QP Opinion

In the opinion of the QP the data is adequate for the purposes used in the Marigold21TRS.

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10 MINERAL PROCESSING AND METALLURGICAL TESTING

Production began at Marigold in 1989, ore was processed primarily with a rod-and-ball-mill grinding circuit with gold recovery by carbon-in-leach CIL and carbon-in-column (CIC) circuits.

In March 1990, heap leaching commenced at Marigold. Since April 1999, all Marigold ore deposits have been processed via truck dump ROM heap leaching.

Cumulative gold production from the Marigold leach pad (through December 2021) is equivalent to 71.1% recovery, and total gold recovery, including recoverable gold inventory in the pad, is estimated at 74.2%.

Gold production data from the leach pad provide the best possible indicator for future processing recoveries because all future-placed ore is similar to ore that has been processed since 1999. Gold recovery from future ore is estimated to be 74.7% based on a review of historical assay and recovery data as well as metallurgical testwork on future ore.

10.1 Metallurgical Testwork

Metallurgical testwork activities include testing methods to improve gold recovery by testing ore samples to guide short and long-range production planning and optimizing reagent addition to minimise processing costs.

Metallurgical studies continue to be undertaken on Marigold ore types with respect to heap leach recovery. These studies have been based primarily on both small column (25.4 cm diameter by 1.2 m high, with minus 51 mm ore) and standard bottle roll leach testwork. Testwork has been conducted on a variety of Marigold ores, including representative pit samples taken by ore-control geologists, leach pad grab samples from mine production, and various pit blasthole drill cuttings. Bottle roll testwork has also been conducted on exploration RC drill samples to determine expected gold recovery from deposits that will be mined in the future.

Results of gold recovery versus gold grade for all laboratory column tests are shown in Figure 10.1. In addition to undertaking columns bottle roll tests were also completed on the same samples to develop a trend. The correlation between column and bottle roll tests is good, the relationship is shown in Figure 10.2. The adoption of the bottle roll test enables more metallurgical tests to be undertaken in a shorter time frame, months for columns to days for bottle rolls.

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Figure 10.1 Column Test Results – All Marigold Areas

SSR, 2021

Figure 10.2 Bottle Roll vs. Column Recovery – All Marigold Areas

SSR, 2021

10.2 Process Optimization Metallurgical Testwork

Additional testwork has been carried out as required to optimise the processing variables that are controllable on a large heap leach pad and plant.

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These variables include permeability testing undertaken on a number of ore samples with varying clay content. The testing was undertaken on multiple stages to 122 m of compaction. Overall, the blends tested demonstrated relatively consistent permeability on increasing loads. Flow rates for the blends ranged from 178.8–284.2 L/hr/m2 under no load. Under 122 m effective height loading, flow rates ranged from 34.4 L/hr/m2 up to 1880.082 L/hr/m2. All tests resulted in low, but acceptable permeabilities.

10.3 Gold Recovery Modelling

Marigold uses two assay methods: fire assay that measures the total gold in a sample and a second method known as ‘cyanide soluble gold’. This technique generates a value that represents the head grade of the ore in terms of the amount of gold in a finely ground sample that can be dissolved by a strong sodium cyanide solution, or the maximum cyanide soluble gold content.

All Marigold blasthole samples are assayed for cyanide soluble gold. Samples from each ore polygon delineated by ore control are selected for fire assay based on the grade distribution for the polygon tonnage and targeting one sample per every 2,000 short tons of ore. Therefore, some samples have two assay values: an AuCN (cyanide soluble) value; and an AuFA (fire assayed) value. The ratio of AuCN / AuFA provides the theoretical maximum gold recovery that can be achieved.

For example, if the AuFA ore grade is 0.10 g/t, and the AuCN ore grade is 0.08 g/t, the ratio is 0.008/0.010 = 0.80. This indicates that the maximum gold recovery from this ore sample is 80%. A ratio greater than 1.0 (100%) is impossible.

Testwork has demonstrated that, generally, all ore at Marigold behaves similarly. The ratio of AuCN / AuFA is an important characteristic of each ore block.

The most recent assessment of the predicted recovery for Marigold ore was conducted in 2017. The 2017 exploration database contains approximately 155,000 pairs of fire assays (field AUFA in the database) and cyanide soluble assays (field AUAA in the database). These assay pairs represent all the mine ore types. On an individual ore block basis, the ratio AuCN / AuFA includes all the local geological variables for that ore block (rock type, degree of oxidation, head grade, etc.). The result is the best estimate of maximum recovery. Figure 10.3 shows AuFA plotted against AuCN for all data pairs.

A best-fit linear regression shows the AuCN / AuFA ratio is 0.8037:1 (~80% recovery).

The LOM actual leach pad recovery is 74% (including in-process gold inventory through December 2021).

An adjustment factor can be calculated using the chemical maximum AuCN / AuFA recovery and the actual pad recovery:

Actual: 74% / Chemical: 80% = 0.92

Therefore, the estimated recovery from the ROM heap leach can be expressed as:

Heap Leach Recovery = AuCN / AuFA x 0.92

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Figure 10.3 Exploration Database (2017) AuCN vs. AuFA – All Data

SSR, 2018

10.4 Summary and Recommendations

Marigold ore types behave metallurgically very similarly based on testwork and operating performance. To predict future gold recovery it is recommended that the following studies and work be undertaken:

•Assessment of the AuCN : AuFA ratio be undertaken regularly using updated exploration and blast hole data.

•Ongoing column and bottle roll metallurgical test on heap leach feed composites to determine maximum possible gold recovery.

•Metallurgical testwork on any future ore sources to develop geometallurgical properties and parameters.

•Further studies and assessment of heap leach recoverable Au inventory.

10.5 QP Opinion

In the opinion of the QP the data is adequate for the purposes used in the Marigold21TRS and the analytical procedures used in the analysis are of conventional industry practice.

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11 MINERAL RESOURCE ESTIMATES

11.1 Introduction

SSR has prepared the Mineral Resources estimate for Marigold effective as at 31 December 2021. The Mineral Resources estimate is based on all available data for Marigold as of 31 December 2021. The Marigold21TRS QPs have reviewed and accepted this information for use in the Marigold21TRS.

Mineral Resources are reported exclusive of Mineral Reserves. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. Due to the uncertainty that may be attached to Inferred Mineral Resources, it cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resource as a result of continued exploration.

11.2 Drillhole Database

The digital drillhole database used for this estimate contains a total of 9,066 drillholes with a total length of 1,865,818 m. (SSR uses geoXpedite, a commercially available geology database management system.)

The drillhole database includes collar coordinates, downhole surveys, assays, rock types and oxidation details in separate tables. The database included all the gold assays from the Assay Program and all the data from the Valmy property purchased from Newmont. All relevant validation checks were conducted while importing the data into the database. Fire-assay equivalent and cyanide-assay equivalent gold values were calculated, as discussed in Section 9.1, after importing the comma delimited format files into MineSight. Once imported, the database was checked for errors using the validation tools available in MineSight.

11.3 Domains

The gold mineralization at Marigold is closely associated with the intersection of high-angle fault structures and favourable horizons that intersect these structures. Favourable host rocks in the Antler Sequence are the debris flow horizon in the Edna Mountain Formation, the interbedded limestone/sandstone/siltstone and conglomerate in the Antler Peak Formation, and the conglomerate in the Battle Formation. Favourable host rocks in the Valmy Formation are quartzite and interbedded quartzite-argillite.

The Marigold deposit is divided into seven broad domains based on: orientation of the mineralizing structures; density of structures; orientation of the mineralized zones; and grade distribution.

Figure 11.1 shows the following seven broad domain areas, which include the following:

•Domain 1 Basalt and Antler pit areas

•Domain 2 Target

•Domain 3 Mackay (HideOut, East Hill, Herco North)

•Domain 4 Mackay North (8Sx, 8S, 8N)

•Domain 5 5N/5NE

•Domain 6 TZN

•Domain 7 Valmy pit

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Figure 11.1 Location of the Seven Major Domains over Depleted Topography as at 31 December 2017

SSR, 2018

Geological mapping and drillhole data were used to identify the major structural orientations that control the distribution of mineralization at Marigold. These structural orientations trend north–south, north-east and north-west and are shown on Figure 11.2.

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Figure 11.2 Oblique Plan View Showing the Major Structures

NS (blue), NE (green) and NW (red) with respect to pit locations

SSR, 2018

An envelope of 30 m around the high-angle structures was developed around the interpreted structures to represent the high-angle domains. Figure 11.3 shows a typical cross section with interpreted structures and high-angle domain envelopes.

The first drill intersection of the formational contact and the interpreted structural data were used to generate the bottom surface for Alluvium, the bottom of Havallah Formation, the top of Antler Sequence and the top of the Valmy Formation. The Antler and Valmy Formations are considered two different formational domains for the exploratory data analysis and grade estimation process.

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Figure 11.3 Typical East–West Cross-Section along 10,300 N

SSR, 2018

The base of the oxidized and transition zones was interpreted with respect to geological logging and analytical data.

11.4 Geological Interpretation

Geological interpretations of structures and rock types were initially conducted on east–west cross sections every 30 m, with select north–south long sections and oblique sections as part of the iterative process.

Mineralized envelopes were delineated using a breakeven cut-off greater than or equal to 0.1 g/t bench (7.6 m) composite gold values in cross sections (east–west) 30 m apart with a clipping of 15 m on either side. Bench composites were used to define the ore zones instead of mineralized drillhole widths because selective mining is not considered an option. The addition of the lower grade gold values from the Assay Program expanded the mineralized envelopes. The mineralized envelopes define the ore zones within which the gold grades were estimated. All known and interpreted structures were considered when the mineralized envelopes were generated.

The internal waste was delineated within the mineralized envelopes wherever possible. In the previous estimates, the internal waste envelopes were defined by connecting these intervals between drillholes on sections and into the preceding and succeeding sections. Based on the large positive tonnage reconciliation and grade control information gathered over the previous 3–4 year period, no effort was made to connect these intervals unless there was a continuity on the preceding and succeeding cross sections. The internal waste was defined as small envelopes encompassing composites that were less than 0.1 g/t Au inside the mineralized envelope. A typical cross section is shown in Figure 11.3.

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The complex nature of the mineralized envelopes made it impractical to create 3D wireframes. The mineralized and waste envelopes from the cross sections were sliced at 7.6 m bench plans and were used to define the mineralized envelopes on each bench. The mineralized envelopes from the bench plans were reviewed and verified on cross section in an iterative process and any volume discrepancies were corrected on plans and sections. A typical bench plan is shown in Figure 11.4.

Figure 11.4 Typical Bench Plan (level=5000)

SSR, 2018

11.5 Exploratory Data Analysis

Exploratory data analysis (EDA) was conducted to:

•Understand the gold distribution and recognise any systematic spatial variation of gold grade with respect to major structures and rock units;

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•Identify distinctive geologic domains that should be evaluated independently in the resource estimation;

•Identify any data and analytical errors not identified in the data verification process; and

•Improve the quality of the estimation by understanding the classical statistics of the dataset.

The EDA process involved visual inspection of the raw assay data to establish structural and mineralization trends. Bench composites (7.6 m) were created to match mining selectivity; these composites were reviewed, and those composites within the mineralized envelopes were flagged by domain using the following criteria:

•Location – Basalt and Antler Pits, Target II, Mackay, Mackay North 1, Mackay North 2, 5N/5NE and Valmy pits;

•Formation – Antler, Valmy; and

•Structural domain – high-angle or low-angle domain.

There are 31,971 bench composites flagged within the mineralized envelopes. Table 11.1 provides the basic statistics for gold grades by domain.

Table 11.1 Basic Au g/t Statistics of 7.6 m Bench Composites within the Mineralized Envelopes by Domain

Domain Location			Formation			Structural Domain			Statistic (Au g/t)
	No. of Samples				Min.				Max.			Mean			Std Dev			CV
Basalt			Antler			Low Angle			1,398			0			5.64			0.4			0.44		1.1097
	Valmy			4,934				0			13.03			0.63			0.94			1.4921
Target II			Antler			Low Angle			550			0			3.22			0.27			0.31		1.1481
	High Angle				978			0			5.72			0.37			0.38			1.0401
	Valmy			Low Angle			1,047			0			3.97			0.29			0.33			1.1221
		High Angle			1,610			0			4.03			0.34			0.35			1.0294
Mackay			Antler			Low Angle			3,716			0			8.85			0.38			0.57		1.5173
	High Angle				1,089			0			9.04			0.48			0.69			1.3854
	Valmy			Low Angle			13,189			0			21.84			0.43			0.71			1.657
		High Angle			9,196			0			15.8			0.45			0.79			1.779
TZN			Antler			Low Angle			157			0.08			0.62			0.19			0.11		0.6079
	Valmy			1,222				0			9.74			0.56			0.85			1.528
Mackay North (8S, 8Sx, 8N)			Antler			Low Angle			2,116			0			86.62			1.04			2.53		2.4393
	Valmy			166				0			1.59			0.31			0.28			0.9284
5N/5NE			Antler			Low Angle			387			0			7.51			0.65			0.94		1.449
	Valmy			23				0.09			0.91			0.24			0.19			0.7753
Valmy			Valmy			Low Angle			2,936			0			7.65			0.45			0.62		1.3895

11.5.1 Outlier Restriction

Bench composites were examined for the presence of local high-grade outliers, which are closely associated with the high-angle structures and favourable rock types. The high-grade outliers were restricted to a certain grade and distance during the grade interpolation process instead of being capped to a specific grade value (see Table 11.2).

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Table 11.2 Outlier Restriction Values and Distance for Various Domains

Domain Location			Formation			Structural Domain			Outlier Range<br>(m)		Outlier Au<br>(g/t)
Basalt			Antler			Low Angle			15.2		2.23
	Valmy			22.9				4.11
Target II			Antler			Low Angle			15.2		1.71
	High Angle				15.2			1.37
	Valmy			Low Angle			22.9			2.06
		High Angle			22.9			2.40
Mackay			Antler			Low Angle			15.2		2.75
	High Angle				15.2			2.05
	Valmy			Low Angle			22.9			5.14
		High Angle			22.9			6.20
Mackay North (8S, 8Sx, 8N)			Antler			Low Angle			15.2		8.57
	Valmy			15.2				2.06
5N/5NE			Antler			Low Angle			15.2		3.60
	Valmy			15.2				3.60
TZN			Antler			Low Angle			15.2		3.43
	Valmy			15.2				3.43
Valmy			Valmy			Low Angle			15.2		2.74

11.6 Material Density

There has been no change to the methodology used to assign density to different formations described in the 2014 NI 43-101 Technical Report.

The density used in the cell model at depth (from original topographic surface) for different material is summarized in Table 11.3.

Table 11.3 Summary of Density for Different Material

Material	Depth<br>(m)	Density
Alluvium/Backfill	>0.00	2.10
Havallah	>0.00	2.48
Valmy/Antler	0.0–533	y=2.4076+(0.0001*DEPTH)
Valmy	>533	2.64

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11.7 Variograms

Correlograms were used in this estimation of Mineral Resources as a tool to describe the pattern of spatial continuity or strength of the spatial similarity of a variable with separation distance and direction. A correlogram measures the correlation between data values as a function of their separation distance and direction. Correlograms were generated using the domain coded composite data using SAGE2001 software (Isaaks & Co.). Structural information from mapping and interpreted structures from the orientation of gold grades were used as a guide to select the along-strike, across-strike, and along-dip directions.

The correlogram was completed for different domains, and the parameters are shown in Table 11.4.

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Table 11.4 Correlogram Parameters Used to Estimate Different Domains

Domain Location		Structural Domain			First Structure									Second Structure									Direction/Dip									Nugget
	X			Y			Z			X			Y			Z			X			Y			Z			C0			C1			C2
Basalt		Low Angle			77			22			8			90			71			265			261/31			169/3			74/59			0.269			0.47		0.26
Mackay and Target II		High Angle			21			96			11			41			263			176			232/7			322/–2			275/20			0.315			0.44		0.25
	Low Angle		9			15			18			83			290			187			102/–77			348/–5			77/12			0.246			0.54			0.22
Mackay North (8S, 8Sx, 8N) and 5N/5NE		Low Angle			15			112			33			54			235			274			81/76			55/–13			327/6			0.181			0.573		0.246
TZN		Low Angle			47			24			11			93			235			56			292/71			92/18			4/–6			0.279			0.378		0.343
Valmy		Low Angle			27			26			7			169			312			30			70/20			355/15			285/15			0.15			0.55		0.3

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11.8 Cell Model and Grade Estimation

The Mineral Resource cell model was created using MineSight v15.80-3. The model extents and the cell sizes are summarized in Table 11.5.

Table 11.5 Cell Model Limits

Item	Minimum*	Maximum*	Extent*	Cell Size*<br>(ft)	Number of Cells
Eastings	–3,000	29,000	32,000	20	1,600
Northings	–8,000	34,000	32,000	25	1,680
Elevation	3,000	8,500	32,000	25	220

* Expressed in Imperial units

The cell dimension was selected based on drillhole spacing; approximately one-third of the drill spacing and cell heights match the future mine bench heights. The model attributes are shown in Table 11.6.

Table 11.6 Model Attributes

Field	Description
TOPO	Percentage of cell below the 31 December 2021 topography
ORE	Ore or waste cells: Ore=1, Waste = 10
ORE%	Percentage of ore within the cell
AUNN	Gold value for NN model
AUKR	Gold value for kriged estimate
AUPAY	Gold value for payable gold grade
CAT	Resource category: Indicated=2, Inferred=3
SDOM1	Low/high-angle structural domain: low angle=2, high angle=5
SDOM2	Low/high-grade domain: low-grade block=2, high-grade block=1
SDOM3	Location: Basalt=1, Target=2, Mackay=3, Hercules=4, 5N/5NE=5
RCODE	Formation/rock unit: Alluvium=1, Havallah=2, Antler=3, Valmy=4, Backfill/dump=6
REDOX	Oxidation state: Oxides=1, Transitional=2, Sulfides=3
TCF	Tonnage conversion factor
ROYL	Royalty
REC	Recovery

The histograms of the composites within the mineralized envelopes for the various domains were generated. These histograms indicated a skewed distribution, with approximately 20% of the bench composites grades for all the domains with a gold grade below 0.1 g/t, indicating internal dilution. The limits of gold mineralization within the mineralized envelopes are difficult to interpret manually with these lower grade ranges. A probabilistic approach is required to identify the higher grade and lower grade cells to avoid overestimation of tonnages and smearing of higher grades into lower grade cells. The chosen method used indicators that set a value of one to each bench composite that had a gold value greater than or equal to 0.14 g/t Au and a value of zero to composites less than 0.14 g/t Au. The values between zero and one were then estimated into the model cells using ordinary kriging.

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The distribution of the indicator estimates (values between zero and one) was compared to the frequency distribution of the nearest neighbour grade model to determine the probability (percentage) that a cell has a grade of 0.14 g/t or higher (high-grade domain). The percentages are different in different domains and show a close continuity to the composites and NN model. The probability thresholds as percentages used in different domains are shown in Table 11.7.

Table 11.7 Probability Percentages for Cells Au>0.14 g/t

Domain	Probability<br>(%)
Basalt	65
Target II	58
Mackay	38
Mackay North 2 (8S, 8Sx, 8N)	64
5N/5NE	60
Mackay North 1 (TZN)	48
Valmy	36

Before the cells were estimated, the cell model was tagged for the following:

•The depleted pre-mining topography as of 31 December 2017 was used to tag the percentage (TOPO) of in-situ material followed by 31 December 2017 surface topography to incorporate all the dumps and backfills;

•The ore and waste envelopes developed on bench plans were used to tag the ore material /internal waste (ORE) and percentage of ore material (ORE%) in each cell;

•The rock type/formation surfaces were used to tag the RCODE variable in the cell model;

•The surface developed for the top of the transitional zone and fresh material was used to tag the REDOX variable in the model;

•The structural domain (SDOM1) was tagged using the high-angle structural envelopes; and

•The grade domain (SDOM2) was tagged using probability percentages.

The composites were back-tagged using the cell model for the different domains and attributes described here.

The cells were then estimated for gold using ordinary kriging in 90 separate calculations.

HideOut and 8Sx mineral centres identified in 2014 and 2015 are located below the historical waste dumps. The material in these dumps was mined during the late 1990s and early 2000s when cut-off grades were higher than the current cut-off grades. While drilling these mineral centres, the samples from these waste dumps were also assayed for gold. A majority of these samples returned gold values higher than our current cut-off grade.

To confirm the grades, a total of 37 sonic drillholes were drilled in 2016. These drillholes confirmed the gold grades in the dumps or mineralized stockpiles. A total of 372 holes drilled between 2010 and 2017 in the waste dumps was considered for this estimation. This stockpile was demarcated using the original and current topography. The samples within these surfaces were selected and bench composited to 7.6 m. The cells were then estimated for gold using inverse distance cubed (ID3) in two separate calculations. The search parameters used to estimate the cells within the stockpile are shown in Table 11.8.

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Table 11.8 Estimation Parameters for Mineralized Stockpiles

Domain		Min No. of Composites			Max No. of Composites		Outlier Range (m)		Outlier Au (g/t)			Search Ellipsoid Distance and Orientation
	X Search (m)			Y Search (m)				Z Search (m)			Max Search (m)			Z Axis			X Axis			Y Axis
Mineralized Stockpile		1			8		12.2		0.342			150			150			15			150		0		0		0
	3		8			12.2		0.342		91			91			15			91			0		0		0

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11.9 Model Validation

The cell model was validated both visually and statistically. Visual validation compares the composites and the estimated model grades in both plan and section. Plans and sections were also checked for smearing of grades across stacked ore/mineralized zones, and no smearing was identified. This validates the kriging parameters used to estimate the cells. A typical cross section and plan with estimated grades are shown in Figure 11.5 and Figure 11.6, respectively.

Figure 11.5 Typical East–West Cross Section along 10,400N Looking North, with Estimated Cell Grades Au g/t

SSR, 2018

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Figure 11.6 Typical Plan 4950 Elevation, with Estimated Whole Cell Grades Au g/t

SSR, 2018

Checks for global bias were conducted on a domain basis, and the relative percent differences of the kriged mean gold grades were checked against the nearest neighbour estimates; the difference was less than ±/-5%.

Swath plots were generated to compare the nearest neighbour gold grades and the kriged gold grades. These plots shown on Figure 11.7, Figure 11.8, and Figure 11.9 demonstrate good correlation.

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Figure 11.7 Swath Plot Along Eastings

opt = ounces per short ton

Au NN is nearest neighbour estimates; Au Kriged is ordinary kriged estimates

SSR, 2021

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Figure 11.8 Swath Plot Along Northings

opt = ounces per short ton

Au NN is nearest neighbour estimates; Au Kriged is ordinary kriged estimates

SSR, 2021

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Figure 11.9 Swath Plot Along Elevation

opt = ounces per short ton

Au NN is nearest neighbour estimates; Au Kriged is ordinary kriged estimates

SSR, 2021

11.10 Resource Classification

There has been no change to the Mineral Resource classification methodology that was described in the most-recent NI 43-101 Technical Report on the Marigold Mine (SSRTR18).

The model cells were classified as Inferred or Indicated based on the parameters in Table 11.9. The sample spacing and the nature of the mineralization do not warrant classification of any resources in the Measured category.

Table 11.9 Resource Classification Parameters

Category	Minimum Composites	Distance to First Composite<br>(m)	Distance to Second Composite<br>(m)
Indicated (CAT=2)	2	36	50
Inferred (CAT=3)	1	78	–

Two resource classification envelopes/polygons were used to classify the Mineral Resources within the mineralized stockpiles. One polygon was digitized based on a distance of 30 m from the exterior composite for Indicated resources and at a distance of 50 m for Inferred Mineral Resources.

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11.10.1 Ore Reconciliation

Reconciliation between resource model estimates and mined production is the most effective means of validating a cell model estimate.

Production since the acquisition of Marigold by SSR has been mainly in Mackay Phase 1 and Mackay Phase 3. Mining is currently underway in Mackay Phase 2 and Mackay Phase 5. The reconciliation for these phases is presented in Table 11.10.

Table 11.10 Ore Reconciliation for the Period Between 2018 and 2021

Item	Tonnage<br>(Mt)	Gold Grade<br>(g/t)	Contained Gold<br>(Moz)
Actual mined	97.4	0.39	1.22
Resource model	94.5	0.41	1.23
Difference	2.95	–0.02	0.01
% Difference	3%	–4%	–1%

Reconciliation between the Mineral Resources model and the grade control model is reasonable. This demonstrates that the Mineral Resources model is able to adequately predict the tonnages and grades for the previous four-year period and can be used to estimate Mineral Reserves.

11.11 Mineral Resource Statement

Mineral Resources for Marigold were calculated based on an optimized pit at a payable gold grade of 0.065 g/t (Au assay factored for recovery, royalty and net proceeds per cell) using an assumed gold price of $1,750/oz.

The gold price of $1,750/oz was selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The metal price is representative of the range of price estimates publicly reported for Mineral Resource cut-offs. The Marigold Mineral Resource is assumed to be mined by open pit.

The very low cut-off grade applied in 2020 (0.065 g/t Au, assay factored for recovery, royalty and net proceeds per cell) is based on current mining performance, which has largely been above the water table. With a shift in the future to mining a higher proportion of Mineral Resources located below the water table, the 2020 cut-off grade was applied for the 2021 Mineral Resources, while waiting for further mining performance data below the water table. Once the additional mining performance data is available, the Mineral Resource cut-off grade rationale should again be reviewed.

By definition, the estimation of Mineral Resources has considered environmental, permitting, legal, title, taxation, mining, metallurgical, infrastructure, socio-economic, marketing and political factors and other constraints, as discussed in various sections of the Marigold21TRS.

SSR is unaware of any current environmental, permitting, legal, title, taxation, socio- economic, marketing, political, or other relevant factors that could materially affect the Mineral Resources estimate (exclusive of Mineral Reserves) as at 31 December 2021 presented in Table 11.11.

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Table 11.11 Summary of Marigold Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price

		Mineral Resources
	Measured				Indicated				Measured + Indicated				Inferred
	Tonnage<br>(Mt)		Au Grade<br>(g/t)		Tonnage<br>(Mt)		Au Grade<br>(g/t)		Tonnage<br>(Mt)		Au Grade<br>(g/t)		Tonnage<br>(Mt)		Au Grade<br>(g/t)
Marigold		–		–		115.3		0.43		115.3		0.43		21.8		0.36
Total		–		–		115.3		0.43		115.3		0.43		21.8		0.36

1. The Mineral Resource estimate was prepared in accordance with S-K 1300.

3. The Mineral Resources estimate is reported below the as-mined surface as at 31 December 2021 and is exclusive of Mineral Reserves.

4. The point of reference for Mineral Resources is the entry to the carbon columns in the processing facility.

5. Metallurgical recoveries used are, on average, 67% for gold.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

Table 11.12 Details of Marigold Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price

Mineral Resources	Tonnage<br><br>(Mt)	Au Grade<br><br>(g/t)	Cut-off Grade<br><br>(Au g/t)	Metallurgical Recovery<br>(%)
Measured	–	–	–	–
Indicated	115.3	0.43	0.065	66%
Measured + Indicated	115.3	0.43	0.065	66%
Inferred	21.8	0.36	0.065	75%

1. The Mineral Resource estimate was prepared in accordance with S-K 1300.

3. The Mineral Resources estimate is reported below the as-mined surface as at 31 December 2021 and is exclusive of Mineral Reserves.

4. The point of reference for Mineral Resources is the entry to the carbon columns in the processing facility.

5. Metallurgical recoveries used are, on average, 67% for gold.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

11.12 Comparison with Previous Estimates

The Mineral Resources have been compared to the previous Mineral Resources which was based on the EOY 2020 pit surface. Comparison of the 2021 Mineral Resource with the 2020 Mineral Resource shows a net decrease in contained gold of 0.44 Moz (-8%).

Key changes in the Mineral Resource (contained metal) have resulted from:

•Mining depletion (-4.9%)

•Engineering changes (-2.3%)

•Leach pad inventory changes (-0.5%)

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11.13 QP Opinion

The Marigold21TRS QP has not identified any relevant technical and/or economic factors that require resolution with regards to the Mineral Resource estimate.

The Marigold21TRS QP reviewed the assumptions, parameters, and methods used to prepare the Mineral Resources Statement and is of the opinion that the Mineral Resources are estimated and prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300).

11.14 Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects

The Mineral Resources reported in the Marigold21TRS are suitable for reporting as Mineral Resources using the Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

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12 MINERAL RESERVE ESTIMATES

The Mineral Reserve estimates for Marigold were completed by the SSR technical department on site. The Marigold21TRS QPs have reviewed and accepted this information for use in the Marigold21TRS.

This section describes the methodology and parameters used to estimate the Mineral Reserves for Marigold. The Mineral Reserves estimate as of 31 December 2021 considers all information used in the Mineral Resources estimate as at 31 December 2021 presented in Section 11.

Lerchs-Grossman pit optimizations were run on the Mineral Resources cell model at a range of gold prices.

The ultimate pits and subsequent phase designs were developed from the $1,350/oz optimization runs. Interramp angles are 37° in mined fill and range between 47° and 49° in rock. The gold price assumption was based on an internal assessment of recent market prices, long-term forward curve prices, and consensus among analysts regarding price estimates.

Mining costs are based on historical values and budgeted costs that include an incremental haulage component using estimated haul cycle times and pit depths. Processing and general and administrative (G&A) costs were estimated based on historical values and budgeted costs. Estimated sustaining capital costs, royalties, severance taxes, and reclamation costs were also included in the optimization costs.

The Mineral Reserves estimate for Marigold was calculated using the as-mined surface at 31 December 2021 with the following assumptions and parameters:

•The reserve classification converts Indicated Mineral Resources to Probable Mineral Reserves within the pit design. There is no Measured Resources category in the Mineral Resources model, and Inferred Mineral Resources are not considered ore when calculating the Mineral Reserves;

•The mining recovery is 100% within the pit design;

•The Mineral Resources were not diluted (see Section 11 for reconciliation data). Internal dilution included in the Mineral Resource estimate is considered adequate;

•The Mineral Reserves estimate assumes that mining operations will continue to use the current Marigold mining methods, as described in Section 13; and

•The estimated cut-off grade was 0.0019 oz/st payable Au or 0.065 g/t payable Au (Au assay factored for recovery, royalty and net proceeds).

The Marigold21TRS QPs are unaware of any current environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that could materially affect the Mineral Reserves estimate as at 31 December 2021 presented in Table 12.1.

12.1 Mineral Reserves Estimate

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Table 12.1 Summary of Marigold Mineral Reserves Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price

Marigold		Mineral Reserves
	Proven				Probable				Total
	Tonnage<br><br>(Mt)		Au Grade<br><br>(g/t)		Tonnage<br><br>(Mt)		Au Grade<br><br>(g/t)		Tonnage<br><br>(Mt)		Au Grade<br><br>(g/t)		Contained Gold <br>(koz)
In Situ		–		–		203.8		0.48		203.8		0.48		3,173
Leach Pad														237
Total		–		–		203.8		0.48		203.8		0.48		3,410

1. The Mineral Reserve estimate was prepared in accordance with S-K 1300.

2. The Mineral Reserve estimate is based on metal price assumptions of $1,350 gold.

3. The Mineral Reserve estimate is reported at a cut-off grade of 0.065 g/t Au.

4. Economic analysis for the Mineral Reserve has been prepared using long-term metal prices of $1,600/oz.

5. No mining dilution is applied to the grade of the Mineral Reserves. Dilution intrinsic to the Mineral Reserves estimate is considered sufficient to represent the mining selectivity considered.

6. The point of reference for Mineral Reserves is the entry to the carbon columns in the processing facility.

7. SSR has 100% ownership of the Project.

8. Metals shown in this table are the contained metals in ore mined and processed.

9. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

Table 12.2 Details of Marigold Mineral Reserves Estimate (as at 31 December 2021) Based on $1,350/oz Gold Price

Mineral Reserves	Tonnage<br><br>(Mt)	Au Grade<br><br>(g/t)	Contained Gold<br>(koz)	Cut-off Grade<br><br>(Au g/t)	Metallurgical Recovery<br>(%)
Proven	–	–	–	–	–
Probable	203.8	0.48	3,173	0.065	74.69
Leach Pad	–	–	237	–	–
Total	203.8	0.48	3,410	0.065	74.69

1. The Mineral Reserve estimate was prepared in accordance with S-K 1300.

2. The Mineral Reserve estimate is based on metal price assumptions of $1,350 gold.

3. The Mineral Reserve estimate is reported at a cut-off grade of 0.065 g/t Au.

4. Economic analysis for the Mineral Reserve has been prepared using long-term metal prices of $1,600/oz.

5. No mining dilution is applied to the grade of the Mineral Reserves. Dilution intrinsic to the Mineral Reserves estimate is considered sufficient to represent the mining selectivity considered.

6. The point of reference for Mineral Reserves is the entry to the carbon columns in the processing facility.

7. SSR has 100% ownership of the Project.

8. Metals shown in this table are the contained metals in ore mined and processed.

9. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

12.2 Cut-off Grade

The estimated cut-off grade for Mineral Reserves was based on a $1,350/oz gold price and current operating costs and metallurgical performance. The gold price of $1,350/oz was selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The metal price is representative of the range of price estimates publicly reported for Mineral Reserve cut-offs. Factors used to estimate the cut-off grade are outlined in Table 12.3, and include refining charges, royalties, and net proceeds tax.

An average recovery rate of 74.9% was used to estimate the cut-off grade based on the average of model recoveries from the 2019 Strategic Business Plan.

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Table 12.3 Key Economic Parameters for Mineral Reserves Estimate

Material Type	Unit	Ore	Rock Waste	Dump
Gold Price (per ounce)	$/oz	1,350
Oil Price (per barrel)	$/bbl	55
Mining Cost (per tonne)	$/t	1.991	1.813	1.560
Processing Cost (per ore tonne)	$/t	1.669
G&A (per ore tonne)	$/t	0.868
Average Process Recovery (formula)	%	74.9%
Refining Charge (per ounce)	$/oz	0.79
Internal Cut-off	g/t	0.065

12.3 Royalties, Net Proceeds and Excise Tax

NSR royalty payments vary between 0% and 10% of the value of production net of off-site refining costs, which is equal to an annual average range of 3.7% to 10%, as further described in Section 3.

The State of Nevada imposes a yearly tax on the net proceeds of all mining operations conducted within the state, plus a yearly property tax on all fixed and mobile equipment used by the mining operation. The net proceeds tax is based on the income from the sale of all products from the mine minus: the royalties; mine, plant, and administration expenses sourced in the State of Nevada; development expenses paid during the year; prescribed depreciation of tangible assets according to set, pre-defined classifications contained in state regulations; and reclamation expenditures incurred during the year of the tax. A net proceeds tax of 5% was applied to the Mineral Reserves estimation.

In 2021 the State of Nevada enacted Assembly Bill 495, effective 1 July 2021, which is an annual excise tax on gold and silver revenue. Under the bill, the tax rates vary based on the taxpayer’s Nevada gross revenue. A 0.75 % rate is imposed on Nevada gross revenue of more than $20 million but not more than $150 million in a taxable year (defined as the calendar year). A rate of 1.10 % applies to Nevada gross revenue exceeding $150 million in any tax year. The LOM average rate is about 0.9%.

12.4 Dilution

No mining dilution was applied to the grade of the cells. Dilution intrinsic to the Mineral Resources model is considered sufficient to represent the stated mining selectivity.

12.5 Mining Recovery

Mining recovery was assumed to be 100% of the Indicated Mineral Resources. Inferred Mineral Resources were assigned as waste.

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12.6 Comparison with Previous Estimates

The Mineral Reserves have been compared to the previous Mineral Reserve which was based on the EOY 2020 pit surface. Comparison of the 2021 Mineral Reserve with the 2020 Mineral Reserve shows a net decrease in contained gold of 0.28 Moz (-8%). Changes have occurred from mine depletion, leach pad inventory changes, infill drilling results, Resource model updates and design changes. Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects

The Mineral Reserves reported in the Marigold21TRS are suitable for reporting as Mineral Reserves using the Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

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13 MINING METHODS

Drilling and blasting occurs on 15.2 m benches. One grade control sample is taken from each blasthole with the sub-drilling excluded. Mining occurs on 15.2 m benches when pre-stripping waste or mining ore areas with the P&H electric shovel. 7.6 m benches are mined using the smaller hydraulic shovels to minimise the dilution that would otherwise occur from dozing a 15.2 m high face to these smaller shovels Blasting is done with an ammonium nitrate and fuel oil (ANFO) blend and a sensitized ANFO emulsion. The ore control mark-out procedure includes blast movement analysis for 90% of ore production blasts.

Loading operations are currently performed using one electric shovel and three hydraulic shovels. Waste and ore haulage is performed with a fleet of 300 t primary haul trucks.

Equipment maintenance is performed on site for all equipment. There are no contract mining operations on site, other than for blasting as detailed in Section 13.7.

13.1 Geotechnical, Hydrological, Pit, and Other Design Parameters

Historically, Marigold pits have been designed with interramp angles (IRAs) at 48° to 50°. The primary rock, a quartzite in the Valmy Formation, dips in a westerly direction at 40° to 70°. The east highwall, which has rock dipping out of the face, is designed at 48° to 50°. The west highwall, which has rock dipping favourably into the face, is designed at 50°. Achieved IRAs range between 48° and 50°. Because many of the interim and final pit walls are within the Valmy Formation, the steeper 50° angle is thought to be achievable for pit designs within the same rock unit (Knight Piésold, 2014). Call & Nicholas, Inc. (CNI) consultants perform an annual audit of activities and provide guidance if any issues arise with slope stability. A 2019 CNI Slope Stability Study of the Red Dot design based on the results of a 2018 geotechnical core drilling program recommended flattening the slope of the west wall of Red Dot to 47° to 49° and the east wall to 45°. The results of this study were used to inform the ultimate pit design for Mackay / Red Dot pit.

The Marigold geotechnical management plan (GMP) was implemented in 2011. The GMP is continually updated with information as mining progresses.

In 2012, a robotic highwall monitoring station was installed at a primary mining location to survey prisms placed strategically on highwall catch benches. The survey instrument was replaced with a highwall radar monitoring system in 2015, and a second system was added in 2017 and a third system in 2019. These allow for 360° monitoring of highwalls in the Mackay pit or multiple areas within other pits. These three radars provide coverage 24 hours per day. Threshold values with respect to movement are programmed into the system. If these values are exceeded, notifications are sent across the wireless network to the dispatch control centre and to the geotechnical team. If the movement is significant, the notifications are sent to senior management.

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Mining below the regional water table commenced in 2020 using a combination of in-pit sumps and emulsion blasting, These short-term solutions were adopted, pending the completion of permitting and construction of primary dewatering facilities. The Mackay 2019 EIS approved dewatering to allow mining below the water table. The mine dewatering plan is discussed in Section 14.

Haul road and ramp widths are designed for two-way traffic that accommodates 300 t class haul trucks. The total road width, including berms and ditches, is 36.4 m. The roads follow topography external to the pit and do not exceed a 10% grade. Ramps inside the pits are also designed at a 10% maximum grade.

Waste dumps are placed in lifts of 15.2–45.5 m high, with benches left on the outside edges for a final 3:1 slope pushdown. There have been no waste dump stability issues on the Property.

The leach pad is similarly built with lifts of 6.1 m to 12.2 m high, with benches left on outside edges for a final 3:1 slope pushdown. The leach pad is permitted to a 121.2 m height above the plastic liner at the base. As each new leach pad cell is designed and permitted, a geotechnical analysis is completed. There have been no leach pad stability issues on the Property.

13.1.1 Open Pit Geotechnical Reports Review

A review of previous geotechnical studies was conducted in 2021 to confirm that studies completed to date are appropriate and to identify any gaps or areas of residual concern, (PSM, 2021).

The following reports for Marigold were provided and form the basis of the review:

•2018 – NI 43-101 Technical Report on the Marigold Mine (31 July 2018) (SSRTR18)

•2019 – CNI Slope Stability Study of the Red Dot design

•2021 - CNI site visit recommendations

•2021 - CNI analysis of soil slopes

•2021 - Piteau Associates (Piteau) Mackay pit dewatering system design

The available reporting does not represent all the data that may be available, particularly in view that mining has been ongoing since 1988. Moreover, the 2018 NI 43-101 Technical Report on the Marigold Mine (SSRTR18) indicates Knight Piésold involvement in 2014 and with CNI involvement since 2015.

13.1.1.1 Overview of Geotechnical Report Review

Below is a summary of comments that represent perceived gaps in the geotechnical reporting for the Marigold open pit:

•The CNI stability analyses of the overall slopes are considered to have an element of conservatism owing to approach in assigning rock mass strengths and utilizing a linear Mohr-Coulomb strength envelope. With use of Hoek & Brown strengths, higher FOS for overall slopes could be anticipated in some areas.

•Further consideration of the potential impact of faults on large scale pit wall stability is recommended. The stability assessments do not address the potential impact:

◦On western pit walls of thrust faults dipping moderately to the east.

◦Potential wedges between faults parallel to the primary bedding fabric.

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◦Faults dipping steeply to the east which can form shallow wedges plunging to the south and which may impact the north wall once below the water table, where pore pressures may influence wedge stability.

•The CNI batter face angle and berm width designs, without appropriate consideration of blasting are not considered sound. Such designs, with proposed batter face angles nominally 10° steeper than typically achieved would potentially allow loose material to fall whilst faces are being dug and also result in berms being filled with rill. It may be more effective to dig batters to nominally 63° and have berm widths closer to design or presplitting to achieve BFA’s above 70° where steeper interramp slope angles can be considered (south and west walls) and which could also consider double benching.

There are limited concerns with waste dump and leach pads as these are developed with 3H:1V (~18°) overall angles and neither have presented stability issues on the property.

13.1.2 Pit Optimizations and Designs

Pit optimizations and subsequent pit designs were completed by Marigold personnel in 2020 using the current Mineral Resources estimate.

Optimizations used the Lerchs-Grossman algorithm. MMC developed operating mining costs for the existing mining fleet during the pit optimization process. The mining cost for the pit shells was based on the total mining net of haulage mining costs, which are presented in Table 12.3, in addition to ore and waste haulage costs that were incorporated into the cell model.

The ROM leach recovery model, as developed by MMC, was also incorporated into the Mineral Resources cell model. To facilitate the calculations and the Mineral Resources tabulations, variables were incorporated into the model for recovered gold [gold x recovery] and payable gold [gold x recovery x (1–royalty)]. Payable gold cut-off grades were established at 0.065 g/t Au and 0.104 g/t Au, respectively, for incremental cut-off and breakeven cut-off. Incremental cut-off is based on pit rim routing, so the only mining cost change is the increment between the ore and waste mining costs. Breakeven cut-off includes the ore mining cost.

The mining and processing costs for the evaluation include sustaining capital costs. The mining costs also include the Marigold analytical laboratory because most of the on-site lab work involves assaying production blastholes for ore control. The processing costs also include sustaining capital and the full site reclamation costs.

Overall slope angles used in the optimization are presented in Table 13.1.

Table 13.1 Overall Slope Angles by Azimuth

Pit	Slope Angle<br>(Degrees)
All Pits in Reserves	47.0-49.0
East Wall Mackay	45.0
Fill Material	35.0

Twelve Lerchs-Grossman pit optimizations were run at gold prices from $800/oz to $1,500/oz: $100/oz increments from $800/oz to $1,000/oz; $50/oz increments from $1,000/oz to $1,400/oz; and $100/oz increments from $1,400/oz to $1,500/oz.

The $1,350/oz gold price cone was selected as a guide to develop the ultimate pit and subsequent pit phase designs.

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Geotechnical review recommendations provided by Knight Piésold (2014) and confirmed by CNI on pit slope geometry were incorporated into the pit designs. Berm/catch bench widths range from 7.2–8.2 m in rock and from 7.2–15.4 m in fill and are designed for every 15.1 m bench height.

13.2 Pit Phases and Timing

The pit optimization for the LOM plan used a Lerchs-Grossman algorithm with an internal recoverable gold value of 0.065 g/t. The optimized pit was built into an ultimate pit design that includes access and takes into account geotechnical considerations for designed highwall angles.

The overall design has three distinct areas: the main Mackay pit, the North Mackay pits, and the Valmy area pits. Figure 13.1 shows the ultimate pit configurations, including current backfill.

The Mackay ultimate pit is an expansion, consolidation, and deepening below the water table of four existing pits into a single pit of approximately 4.6 km long, 1.8 km wide, and 430 m deep. It contains more than 60% of the mineral reserve tonnage. For sequencing purposes, the ultimate pits are designed into 15 logical development stages.

Tonnages for each mining phase are shown in Table 13.2.

Table 13.2 Mining Phase Design Summary

Phase Name	Ore<br>(kt)	Waste<br>(kt)	Strip Ratio
5N2	3,051	11,030	3.62
8SExt	16,539	61,224	3.70
EB	10,236	59,764	5.84
H1	2,344	24,507	10.46
M4P1	3,229	3,147	0.97
M4P2	35,232	72,496	2.06
M5	779	427	0.55
M7	11,113	35,475	3.19
M9	15,374	40,677	2.65
Mud1	1,311	6,293	4.80
Mud2	958	4,652	4.85
RD	72,386	259,298	3.58
TZ	18,026	98,890	5.49
VS	9,216	39,649	4.30
VN	4,757	10,531	2.21
Total	204,550	728,061	3.56

*Differences between the Mineral Reserve and LOM quantities used in the economic analysis are due to differences between planned and actual 31 December 2021 face positions

* Totals may not match due to rounding

13.3 Production Rates, Mine Life, Dimensions and Dilution Factors

Mining is scheduled 24 hours per day, 363 days per year on a rotation of two 12-hour shifts. The current mine plan provides 17 years of operational life, including 11 years of active mining followed by 6 years of processing the heap leach pad.

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In order to meet near-term LOM production rates, the existing shovel fleet of four units will be maintained by deferring retirement of the smaller EX5500 hydraulic shovel to 2023. The haul fleet averages 25 x 300 t class units and will peak at 28 trucks. Short term variations in mine fleet requirements are managed by delaying retirement of older units when they are scheduled to be replaced. The average sustained mining rate is 90.4 Mtpa over the first 10 years of the remaining 11 year mining life while ore delivery to the ROM leach pad is at an average annual rate of 19.7 Mt. Average payable gold production over the 10 years of full production is approximately 222,000 ounces per year. In general, ore will be mined on 15.2 m benches.

The mineralized zones are structurally controlled and strike in a generally northern direction. They vary in width throughout the Property from one metre or less up to 40 m long and 49 m wide. In the LOM model, there is no dilution or mining loss added to the Mineral Reserves for planning and scheduling.

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Figure 13.1 Marigold Ultimate Pit (end of year 2032)

SSR, 2021

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13.4 Stripping Requirements

The LOM strip ratio is 3.6:1. Stripping requirements are consistent over the life of the main Mackay pit area at an average strip ratio of 3.3:1. The stripping requirements for the other two areas, Mackay North and Valmy, are planned to be above the LOM average at 6.6:1 and 4.6:1, respectively. Table 13.3 and Figure 13.2 show the annual production schedule for the LOM, including ore tonnes mined, waste tonnes mined, and strip ratio.

Table 13.3 Annual Production Schedule Tonnes Mined

Year	Ore<br>(kt)	Waste<br>(kt)	Strip Ratio
2022	21,818	80,800	3.70
2023	22,010	68,638	3.12
2024	21,411	71,419	3.34
2025	15,713	78,276	4.98
2026	16,538	74,096	4.48
2027	20,857	62,370	2.99
2028	20,207	70,152	3.47
2029	26,912	52,024	1.93
2030	18,189	75,674	4.16
2031	13,103	74,301	5.67
2032	7,792	20,313	2.61
Total	204,550	728,061	3.56

*Differences between the Mineral Reserve and LOM quantities used in the economic analysis are due to differences between planned and actual 31 December 2021 face positions

* Totals may not match due to rounding

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Figure 13.2 Mine Annual Production Schedule

SSR, 2021

13.5 Required Mining Fleet and Machinery

The equipment list for the Marigold mining fleet is presented in Table 13.4. Capital replacement of mining equipment is scheduled throughout the LOM plan as sustaining capital when a piece of equipment reaches the end of its useful life and cannot be repaired or rebuilt economically. Sustaining capital is not planned within the last five years of the LOM plan because it is assumed that equipment life can be stretched out and replacements are difficult to justify near the end of the Property life. The sustaining capital replacement costs are included in the reserve optimization calculation costs. Capital costs are discussed in Section 18. As of the date of the Marigold21TRS, MMC does not employ contract mining services, except with respect to blasting, as discussed in Section 13.7.

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Table 13.4 Marigold Mining Fleet Equipment List

Number of Items	Equipment Name and Class
1	P&H 4100 XPC electric shovel
2	Komatsu PC7000 hydraulic shovels
1	Hitachi EX5500 hydraulic shovel
1	Caterpillar 992 wheel loader
8	Hitachi EH5000 300 t haul trucks
17	Komatsu 930E 300 t haul trucks
1	Caterpillar 789B haul truck
3	Caterpillar 789B water trucks
3	Ingersoll Rand DML drills
3	Atlas Copco PV271 drills
4	Caterpillar 834 and 854 wheel dozers
6	Caterpillar D10 and D11 track dozers
3	Caterpillar 16H and 16M motor graders
3	Lube / fuel trucks
1	Caterpillar 637 scraper
1	Caterpillar 789 Lowboy heavy hauler

13.6 Ore Control Drilling and Method

Blasthole sampling is used to define ore zones. A grade control sample is taken every 15.2 m of drilling. The sample is manually collected from a cross-section of the cone of drill cuttings. The procedure includes removal of the sub drill material. Ore Control personnel periodically audit the performance of the blast hole samplers and provide feedback on compliance to standard. Benches are mined 15.2 m with an electric or hydraulic shovel in stripping and bulk ore mining areas.

If ore is encountered in the stripping areas on the 15.2 m benches, it is mined at that bench height to maintain pit productivity. A dilution factor is added to the monthly survey using a 1.0 m rind around ore shapes at the calculated grade for the shape. This is added to the surveyed tonnage for the bench and reported as ore mined during the month.

Each blasthole sample is analysed for gold at the on-site laboratory facility. A cold cyanide digestion is performed on each sample to determine the quantity of cyanide soluble gold contained in the sample. Due to the non-destructive analysis method of the cold cyanide leach, it generally does not measure the total amount of gold in a sample. At Marigold, about one in every five blasthole samples containing 0.10 g/t (historically, 0.003 oz/st) cyanide soluble gold is assayed for total gold content using FA with a gravimetric finish. Samples from each ore polygon delineated by ore control are selected for fire assay based on the grade distribution for the polygon tonnage and targeting one sample per every 2,000 short tons of ore. The FA results (Au g/t) from the blastholes and exploration drillholes in the pit area, and cyanide soluble assay results (Au g/t) are used to determine a fire-assay-to-cyanide-soluble ratio for the pit area. This ratio is applied to all remaining cyanide soluble assays in the blast to calculate a total gold value contained in each blasthole.

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FA grades associated with each blasthole are entered into the grade control (blasthole) model. The blast pattern is then converted to a blasthole cell model with cell sizes of 3.05 m x 3.05 m x 7.6 m. The blasthole data is kriged using ordinary kriging in two dimensions on the bench. If there is sufficient volume above the cut-off grade to make a mineable shape of ore, this is blocked out and surveyed in the pit (indicated by ore flags for mining) to be sent to the leach pad for processing.

13.7 Drilling and Blasting

Blasthole drilling is performed with three Atlas Copco PV271 rigs that drill with both rotary and hammer drill bits as well as three Ingersoll Rand DML rigs that drill with hammer bits. The rigs drill 22.2 cm diameter blastholes. The PV271 rigs can drill to 16.8 m in a single pass. The DML rigs can drill to 10.4 m in a single pass.

The normal explosive is a heavy ANFO (blend of ANFO and emulsion) which is placed by a combination of both contractor and Marigold employees. An emulsion product is also used for wet holes to manage groundwater in the winter and fall and help break up the rock in areas of the pit that are more difficult to dig.

The blast patterns are adjusted for rock conditions. Typically, the patterns are 7.3 m x 6.4 m for the 15.2 m benches. To help break the toe of the bench, 1.5 m of sub drilling is added to each hole. The ore host rock generally breaks easily with blasting, and this provides a good ROM leach feed to the pad. Electronic detonators are used to control the timing of the blasthole detonation. The typical fragmentation is P80 of 20.3 cm.

A trim blast is performed around the limits of the mining on final highwall configurations. This is a four-row pattern that is shot to a free face to minimise blast damage and vibration into the highwalls. Historically, a presplit blasting pattern has been used on final highwalls to ensure good wall conditions and minimise the potential for a wall failure. A new crest and catch bench are formed every 15.2 vertical metres of mining that ranges from 6.7–9.1 m depending on the highwall angle.

13.8 Loading Operations

Loading operations are performed with one electric Komatsu 4100 XPC rope shovel with a 52.8 m3 dipper, two diesel hydraulic Komatsu PC7000 hydraulic shovels, and one diesel hydraulic Hitachi EX5500 shovel. Double-sided loading is typically used where there is adequate working room. Digging faces are defined by ore control and are marked in the field with flags and on maps that are provided to the operators. All loading units are equipped with a high-precision digging screen that is a component of the Modular Dispatch system. The screen, located in the operator’s cab, updates in real time to show the location and grade of the ore material being mined. Dig boundaries are typically adjusted to allow for movement associated with blasting.

13.9 Hauling Operations

Excavated rock is loaded into haul trucks and sent to either a waste dump or a leach heap based on the average gold grade of the material. Waste rock is hauled to the multiple waste stockpile locations or to previously mined-out areas for backfilling pits. Pit backfilling, where not mandated by permit to eliminate pit lakes in certain satellite pits, has positive impacts at Marigold: it reduces costs associated with haulage distance and helps address the lack of dump space due to permitting restrictions and current land position. Backfilling plans are reviewed and adjusted to minimise the potential for sterilizing future mineralization. Minimizing the waste haulage distance to the nearest facility improves mining productivity and minimises haulage costs. Ore is hauled to the leach pad facility and stacked in lifts for processing. Pit and dump progression stages at the end of each of each year of the LOM plan are presented in Figure 13.3 to Figure 13.8.

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Marigold has a mixed fleet of Hitachi and Komatsu 300 t class haulage trucks for ore and waste haulage.

A Modular Dispatch system is used to optimise fleet management. Trucks are sent haulage assignments according to priorities set for the loading units and which loading unit requires a truck at that time.

Annually, from December to February, there is snow, fog, and freezing temperatures at the Property. However, there is a minimal amount of haulage downtime due to the weather in most years.

13.10 Mine Support

Mine support functions are performed using different quantities and types of equipment. These include water trucks, dozers and graders as well as other non-operated ancillary equipment such as the radar highwall monitoring units. Mine support functions include ripping leach pads after a panel is completed, monitoring slope stability, maintaining roads and access points, and developing exploration drill pads. This work is completed with a fleet of Caterpillar D8, D10 and D11 class track dozers and Caterpillar 18, 16H and 16M motor graders.

Current mine support fleet numbers are included in Table 13.4.

13.11 Mine Maintenance

Mine maintenance is an integral function of the mining operations and relates to the day-to-day upkeep of the mining equipment. Activities such as preventive maintenance, equipment rebuilds and fixing equipment on breakdowns are all included in the mine maintenance function. The objective is to provide efficient maintenance of the mining fleet, thereby increasing reliability and availability of this equipment through effective strategies, planning and continuous improvement. High levels of equipment availability and reliability facilitate operational and delivery performance, resulting in asset intensity reduction, and reduced direct operational and maintenance costs.

13.12 Mine General and Administration

Mine G&A refers to all day-to-day supervision and engineering support of mining operation activity. Expenses included in the mine G&A are mine salary labour charges and fringe benefits, mine office supplies, safety supplies, equipment rentals and leases, light-vehicle tires, miscellaneous contract services, travel expenses, training, and tax and freight charges.

13.13 Mine Safety

Marigold has one mine rescue and emergency response team which is trained to competently assess accident conditions and fight fires. There is one ambulance and one small fire truck available on site and a rescue trailer that is used in emergencies. The Property is set up with hydrants and appropriate connectors, hoses, and wrenches at strategic locations. For mobile equipment fires, the Property is set up with large water trucks equipped with water cannons.

Marigold also has access to and can call either the Valmy Fire Department (5 km away) or Battle Mountain Fire Department (24 km away), when required. There is a monthly training session for the Marigold rescue team to ensure effective participation in any recovery operations in the event of a mine incident.

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Figure 13.3 End of Production Year 2022

SSR, 2021

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Figure 13.4 End of Production Year 2023

SSR, 2021

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Figure 13.5 End of Production Year 2024

SSR, 2021

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Figure 13.6 End of Production Year 2025

SSR, 2021

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Figure 13.7 End of Production Year 2026

SSR, 2021

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Figure 13.8 End of Production Year 2027

SSR, 2021

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Figure 13.9 End of Production Year 2028

SSR, 2021

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Figure 13.10 End of Production Year 2029

SSR, 2021

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Figure 13.11 End of Production Year 2030

SSR, 2021

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Figure 13.12 End of Production Year 2031

SSR, 2021

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Figure 13.13 End of Production Year 2032

SSR, 2021

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13.14 Mine Dewatering

The Marigold Mine Plan of Operations (PoO) Amendment for the Mackay Optimization Project record of decision in October 2019, allowed mining to be carried out below the water table in the expanded Mackay pit. The approved dewatering system incorporated a pit dewatering design by Piteau Associates that consisted of a series of water wells to be located around the periphery of the ultimate pit to extract water for mine operations use and infiltration. Infiltration is by means of a series of rapid infiltration basins (RIBs) located in an area of deep alluvium cover about 5 km north of the operations area. The initial area selected for the RIB location was relocated to an adjacent land section. Permitting delays associated with this relocation have resulted in short-term mine water diversion by means of temporary discharge permits into a local drainage. Trial RIBs are expected to be permitted and constructed in mid-2022 and approval for the expanded RIB field is expected to be in place by early 2023.

The dewatering system is developed in stages with the initial design incorporating 14 dewatering wells, each with a nominal sustainable pumping rate of 1.89 m3/min.

Figure 13.14 shows existing and planned well sites. New wells are developed in advance of the mining elevations required to support the LOM plan. Recent monitoring and modelling of pumping and drawdown rates indicate that the current number of wells included in the design is conservative and potentially not all will be required to achieve the required drawdown.

Some dewatering water is diverted for mine use with the majority delivered by pipeline to the RIB field north of the mine for re-infiltration. The RIBs are located in areas of thick alluvium which facilitates rapid infiltration back into the aquifer and also provides the benefit of attenuation of naturally elevated arsenic levels in the ground water before it reaches the existing water table. Initial testwork indicates that water treatment will not be necessary prior to infiltration but the economic modelling includes a provision for water treatment, should it prove necessary in the future.

Trial RIBs will be permitted and constructed in mid-2022 to allow infiltration performance to be verified and the RIB cell design and configuration to be finalized. Initial RIB design criteria are summarized in Table 13.5.

The RIB design being permitted includes a total of 14 RIBs. Figure 13.20 shows the conceptual layout of the RIBs and spoil piles with the majority located on (BLM) Section 30.

The location of the RIB area relative to the mining operation is shown in Figure 15.1.

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Figure 13.14 Existing and Proposed Dewatering Wells

Piteau, 2021

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Table 13.5 RIB Design Criteria

RIB Cell Design Attribute	Unit	Dimensions (m)
Basin floor length	m	210
Basin floor width	m	59
Basin crest length	m	247
Basin crest width	m	106
Minimum basin depth	m	6.1
Excavation/dump slope	H:V	3:1
Minimum spacing between cells	m	122
Access road width	m	7.3
Infiltration capacity	m/day	0.43
Infiltration capacity per cell	m3/min	3.8
Cell availability	%	50

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Figure 13.20 Conceptual Layout of RIBs and Spoil Piles

SSR, 2021

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14 PROCESSING AND RECOVERY METHODS

14.1 Introduction

The Marigold processing facilities combine industry standard run-of-mine (ROM) heap leaching, carbon adsorption, carbon desorption and electro-winning circuits to produce a final precious metal (doré) product.

The Marigold heap leach and gold recovery circuit is typical in the industry for treating solutions containing gold cyanide. A simplified flowsheet for the process is shown in Figure 14.1.

14.1.1 Ore Stacking on Leach Pad

Ore processing is undertaken via stacking and leaching on a ROM heap leach pad. ROM ore is delivered to the leach pad by haulage truck and stacked in 6.1 to 12.2 m lifts. Pebble lime is added to the haulage trucks from a storage silo to control pH prior to dumping. Fresh ore is ripped and cross-ripped prior to commencement of leaching. The available pad area is divided into manageable cells for inventory and irrigation control. These cells provide control of irrigation duration and time between lifts to manage future ore placement.

14.1.2 Leaching Solution to the Pad

A series of header and sub header lines distribute the barren solution to the pad with application by drip-tubing on the surface and impact sprinklers on side slopes. The overall barren solution application rate to the leach pad is 0.122–0.143 L/min/m2.

The Marigold heap leach solution processing facilities consist of two barren ponds, six pregnant ponds, and one stormwater/overflow pond. Pregnant pond 1 is inactive due to a leak. Ancillary facilities include solution pumps and piping, two separate sodium cyanide addition facilities, two sodium hydroxide addition systems (barren solution pH adjustment) and four locations for antiscalant addition.

The heap leach pad was originally constructed in 1990 and has since expanded as required, with ongoing expansion of solution processing facilities to match production rate and leach area. Barren leach solution (cyanide-bearing solution, very low in gold grade) is applied selectively to different areas of the pad, or cells. At any given time, approximately 0.5 Mm2 of pad area is being leached, with other areas draining or being made ready to accept ore for the next lift.

The barren leach solution is pumped to the leach pad by two independent barren solution distribution systems. Combined barren solution flow capacity from the two pumping systems is 53 m3/min. Sodium hydroxide is added to the barren solution in the pond after the carbon column discharge. Sodium cyanide is added to the barren solution at the spillway between barren ponds one and two.

The new ore placed on the heap leach pad in cells, is leached for a targeted 120 days.

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Figure 14.1 Simplified Marigold Processing Flowsheet

SSR 2021

14.1.3 Pregnant Solution

Pregnant solution (gold bearing) from the leach pad is collected into the pregnant solution pond(s) and pumped to seven parallel carbon column trains, each with five columns, to recover gold from solution. Column discharge solution reports to the barren ponds before the solution is recycled back to the leach pad.

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14.1.4 Carbon Processing

Loaded carbon from the carbon columns is transported by a dedicated truck to the carbon processing facility where gold is eluted (re-dissolved) from the carbon in strip vessels (two 3 t capacity carbon vessel). The precious-metal-bearing solution is passed through two 2.1 m3 electro-winning cells in parallel where the metals are plated out of the solution. The electro-winning barren solution is recycled back through the strip vessel until the batch process is complete.

14.1.5 Refining

The plated material (sludge) resulting from electrowinning is collected in a filter press and then retorted for drying and mercury removal. After retorting, the sludge is mixed with flux and smelted in a propane fired furnace for final precious metal recovery.

Stripped carbon is screened to remove the fine carbon, acid washed to remove carbonate scale, and thermally reactivated to remove any organic contamination, as required, before returning to the carbon columns.

14.1.6 Ventilation

Ventilation from the strip circuit pregnant and barren solution tanks, electro-winning cells, retort and smelting furnace is directed to a deep bed scrubber (sulfur-impregnated activated carbon) where any vapourized mercury is recovered prior to exhaust.

The kiln discharge is vented to a wet scrubber that uses water mist to condense mercury and recover it as elemental mercury. After demisting, the air is also passed through sulfur-impregnated carbon to recover any remaining vapourized mercury prior to exhaust.

14.1.7 Planned Processing Upgrade Projects

A number of ongoing improvement project are planned, these include:

•With the increasing height of the heap leach pads, barren booster pumps are planned to be installed to maintain solution flow rates at 53 m3/min as the pad grows taller and further from the pump locations.

•Installation of mobile telemetry and instrumentation to be able to remotely monitor individual area barren application rates. In addition, telemetry on primary pregnant and barren flowmeters.

•Further ventilation upgrades on the refinery to improve mercury removal, decrease mercury volatilization, and improve working temperature within the buildings.

14.1.8 Reagents

Reagent consumption rates are within industry norms for the types of ores processed.

Consumption rates for the two most expensive reagents, sodium cyanide (NaCN) and lime (CaO), vary depending on ore type.

Average annual reagent unit consumption rates of the two key reagents, cyanide and lime (CaO) for the period 2010–2021 are shown in Figure 14.2.

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Figure 14.2 Average Annual Reagent Consumption

SSR, 2021

14.2 Gold Recovery

14.2.1 Recovery from Heap Leaching

From March 1990 through December 2021, gold recovery from the heap leach pad is 71.1%. This recovery was achieved with 90–120 day primary leach cycles and an overall mass-of-solution to mass-of-ore ratio of 1.28:1.

The gold recovery trend achieved from March 1990 through December 2021 from the Marigold heap leach is shown in Figure 14.3.

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Figure 14.3 Marigold Heap Leach Pad Gold Recovery Curve from March 1990 through December 2021

SSR, 2021

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15 INFRASTRUCTURE

The overall site layout of the Property is shown in Figure 15.1. Additional details on the LOM plan can be found in Section 13.

15.1 Site Access, Power and Water

15.1.1 Site Access

15.1.2 Power

15.1.3 Operations Water Supply

Water for Marigold is supplied from three existing groundwater wells located near the access road to the Property. Marigold owns groundwater rights and collectively allows up to 3.134 Mm3 of water consumption annually, the majority of which is used as makeup water for process operations. On average, total freshwater makeup is 2.4 m3/min. The well pump parameters are listed in Table 15.1, and the locations of the pumps are shown in Figure 15.2.

Table 15.1 Pump Assets

Pump Asset	Pump Capacity<br>(hp)	Power Consumption<br>(kW)
793-PMP-001	75	56
793-PMP-002	150	112
793-PMP-003	150	112

Discussion of the extraction and infiltration of pit water is included in Section 13.14.

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Figure 15.1 LOM Site Map Showing Final Pit Limits, Waste Dumps, and Leach Pad

MMC, 2021

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15.2 Buildings and Facilities

15.2.1 Buildings and Facilities in Main Plant and Offices Area

The buildings and facilities described below are located in the main plant and offices area as shown in Figure 15.3 and Figure 15.4:

•Truck shop and mobile maintenance warehouse: The Marigold truck shop complex is located near the mine entrance. It is a four-bay shop sized for 300 t class haul trucks. The shop contains a tool crib, oil and lubricant bulk storage, ten offices, locker rooms, training room and warehouse. A covered warehouse storage yard is located adjacent to the admin building complex.

•Mill building: The mill building consists of facilities supporting the metal recovery operations, including the refinery and metallurgical laboratory. Adjacent to the mill building is the thickener water storage tank and remaining CIL tanks from the 1989 flowsheet.

•Crushing plant: The crushing plant is used to produce stemming for blastholes, road material and over liner for heap leach pad. The crusher is a remnant from the 1989 flowsheet.

•Heap leach carbon columns: The heap leach carbon columns are an integral part of the gold recovery process, which is detailed in Section 14.

•Wash bay: The wash bay is located next to the truck shop and consists of one covered bay. The wash bay building also contains a settling pond for water recycling.

•Administration building and light vehicle (old) shop: The main administration building encompasses most site-support departments and includes a small warehouse facility, the shovel and drill shop (former truck shop), light-vehicle maintenance bay and the assay laboratory. Adjacent to this building are trailers which provide additional office space.

•Assay laboratory: The assay laboratory supports ongoing mine operations, including grade control and gold solution analysis.

•Motor control centre (MCC): The motor control centre houses controls for the pumps and boosters for the barren and pregnant solution ponds.

15.2.2 Additional Buildings and Facilities on Site

Additional buildings and facilities on site include:

•Site access building

•Potable water treatment building

•Process line-out building

•Radio shop

•Safety building

•Hose shop and storage

•Tire pad

•Fuel stations

15.2.3 Additional Facilities on Section 20

Additional facilities are located on Section 20, which is identified in Figure 15.1. These facilities include:

•Welding and fabrication shop

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•Dispatch/MineCare office and mine operations line-out building

•GPS dispatch receiver

•Diesel tanks and fuelling station

15.3 Explosives Magazine

The explosives magazine is located a safe distance from the plant and offices area as identified in Figure 15.1.

15.4 Tailings Storage Facility and Water Diversion

The TSF was decommissioned and reclaimed. The only remaining activity concerning the TSF is ongoing monitoring.

The Trout Creek water diversion structure and flood control dam is located west of the former Basalt Pit. It is designed for a 100-year storm event.

15.5 Leach Pads and Solution Ponds

The leach pad is discussed in detail in Sections 14 and is shown in Figure 15.1.

Details on the barren and pregnant solution ponds can be found in Section 14.

15.6 Waste Dumps

Details on completed, in progress, and future waste dumps can be found in Section 13. The general location of planned and current waste dumps is shown in Figure 15.1.

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Figure 15.2 Well Sites

SSR, 2018

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Figure 15.3 Main Infrastructure Area

SSR, 2018

Figure 15.4 Plant, Shops and Offices

SSR, 2018

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16 MARKET STUDIES

16.1 Marketing and Metal Prices

The metal prices used in the Marigold21TRS are based on an internal assessment of recent market prices, long-term forward curve prices, and consensus among analysts regarding price estimates. The metal prices selected have taken into account the current project life. For the economic analysis in the Marigold21TRS, the metal prices shown in Table 16.1 were used.

Table 16.1 Metal Prices

Metal	Unit	Average	2022	2023	2024	2025	2026	>2026
Gold Price	$/oz	1,647	1,800	1,740	1,710	1,670	1,600	1,600
Silver Price	$/oz	21.56	24.00	23.00	22.00	21.00	21.00	21.00

The doré is securely transported by road freight to a refinery where it is refined into gold bullion. The bullion is sold by SSR to banks that specialise in the purchase and sale of gold bullion.

No external consultants or market studies were directly relied on to assist with the sales terms and commodity price projections used in the Marigold21TRS. The QP for this Section 16 agrees with the assumptions and projections presented.

16.2 Contracts

There are a number of acceptable refineries with the capacity to refine doré. Currently, SSR has entered into a non-exclusive refining agreement with Asahi Refining USA, Inc., and the terms and conditions of this contract are within industry norms. The transportation and refining costs for the doré and other operating costs are also in accordance with industry standards.

16.3 QP Opinion

The Marigold21TRS QP considers it reasonable to rely on SSR because SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

Additionally, the project has been in operation for a number of years, and following a review of the current supplied information, the opinion of the QP is the current plans and input parameters appear adequate for use as inputs to the Marigold21TRS.

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17 ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS

17.1 Environmental Studies

Significant portions of the Property exist on public lands administered by the BLM. Therefore, the majority of environmental studies related to mining activities are conducted under BLM authority as part of the NEPA regulations, which require various degrees of environmental impact analyses dictated by the scope of the proposed action. Marigold has undergone several significant NEPA actions in the normal course of operational planning; the most recent is an amendment to the existing PoO to permit the future mining of all pits to their planned maximum depths.

The environmental baseline studies to support the Environmental Impact Statement (EIS) process were initiated in 2013. These baseline studies completed in preparation for the Plan of Operations – Mackay Optimization Project Amendment included, but were not limited to, socioeconomics, air quality impacts, cultural and archaeological resources, groundwater model, pit lake model, screen-level ecological risk assessment (SLERA), waste rock/material characterization, water characterization, sage grouse habitat evaluation, evaluations for flora and fauna, and feasibility evaluation and pilot testing for rapid infiltration basins. A list of the baseline studies and reports is shown in Table 17.1.

Table 17.1 Baseline Studies

Study Media	Documents/Reports Included Baseline Studies and Data Compiled for Marigold Mine Mackay Optimization EIS
Hydrology/Water Quality/Geochemistry	Groundwater Model Report, Waste rock Management Plan, Water Characterization Report, Water Management Plan, Pit Lake Model Report, Screening Level Ecological Risk Assessment Report
Air Quality	Air Quality Assessment
Flora/Fauna	Habitat Evaluations (including sensitive special surveys), Migratory Bird Surveys, Plant Surveys, Weed Management Plan, Raptor Nest Survey, Bat Survey, Sage Grouse Habitat Survey
Socio-Economic	Economic Impact Report
Cultural Resources	Cultural Resource Survey

The final EIS record of decision approving the amended plan of operations was received 30 October 2019.

Scope of the amended plan of operations included:

•Increasing surface disturbance by 833 ha acres on private and public lands.

•Consolidation of multiple pits into three larger pits with associated expansion of pits, waste rock storage areas and leach pads.

•Mining below the historical water table requiring installation of facilities to extract and dispose of excess groundwater.

The approval allowed for infiltration of excess dewatering water by way of rapid infiltration basins (RIBs) on private (leasehold) land north of the mine. An alternative location on adjacent BLM land is in the process of being approved. RIB testing, approval and construction, and the associated water pollution control permit issuance is expected by early 2023. In the interim, mine dewatering and infiltration is proceeding according to the LOM plan by means of temporary surface discharge permits allowing water diversion into local watercourses.

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SSR has a reasonable expectation that all necessary operating permits will be granted within the required timeframes to meet the LOM plan.

17.2 Environmental Permits

Specific federal, state and local (Humboldt County, Nevada) regulatory and permitting requirements apply to Marigold activities. Marigold currently holds active, valid permits for all current facets of the mining operation, including, but not limited to, those permits listed in Table 17.2.

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Table 17.2 Marigold Mine Permits 31 December 2021

Permit Name		Permit Number
Plan of Operations		NVN-065034 (previously N26-88-005P)
Reclamation Permit and Bond		0108
Water Pollution Control Permit (including Petroleum Contaminated Soils Permit)		NEV0088040
Stormwater General Discharge Permit		NVR300000
Title V Air Quality Operating Permit		AP1041-2967
Class II Air Quality Operating Permit		AP1041-3666
Mercury Operating Permit to Construct: Phase II air)		AP1041-2254
EPA/RCRA ID		NVD986766954
Industrial Artificial Pond Permit		39502
Jurisdictional Waters of the U.S. Determination		N/A (no jurisdictional waters) <br>(August 2019 determination)
Class III Landfill Waiver		SW1764
Hazardous Materials Permit (State of Nevada)		97207
Potable Water Permit		HU-1103-NTNC
Septic Permit		GNEVOSDS09-0016
	GNEVOSDS09-S0341
	GNEVOSDS09-L0252
DOT Hazardous Materials Registration		061521550469DF
Liquefied Petroleum Gas – Class 5 License		5-3482-01
Trout Creek Dam Permit <br>(including Dam/Impoundment Permit)		J-666
Water Rights		83256 (Changed 3691 (Certificate 583))
	2324 (Certificate 584)
	86582
	86583
	865841
	865851
	87235-87242
	76425S01, 76425S02, 76425S03
	88986
		90787
		911411
		90788
		3282 (Certificate 499)
		V01898
		2216 (Certificate 498)
County Conditional Use Permit		UH-15-07
MSHA ID		26-02081

1. As at November 2021, permit is pending with the State of Nevada Division of Water Resources

Certain permits listed here are renewed annually and may be issued under a different permit number.

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Given the number of active permits at Marigold, some degree of permit modification or renewal effort is typically underway. With the exception of water-related permits already discussed, approved permits are in place for all planned mining activities.

17.3 Environmental Impacts

At present, there are no known environmental issues that impact the ability to extract Mineral Resources at the Property. Specifically, no threatened or endangered species are known to exist at the site; there are no year-round watercourses on the Property; groundwater impact of mining has been addressed and all environmental regulations and permit conditions are continuously being met. Cultural resource surveys have been conducted across the Property, and an approved program of avoidance, distance buffer and mitigation measures are in place as part of the existing PoO.

Waste rock is managed in several designated surface storage areas within the Property boundary, concurrently reclaimed to 3:1 slopes when the sequence of mining operations allows, and then re-vegetated with native seed mixes. When possible, older pits are backfilled with waste rock. To date, all waste rock encountered at Marigold has been oxide in nature and non-acid-generating as confirmed by quarterly sampling. There are no waste rock areas with observed runoff or stability concerns.

The only tailings area at Marigold operated during a limited period from 1989 to 1999; this area has been reclaimed and revegetated, and the State Engineer’s office no longer lists it as a permitted dam.

17.4 Environmental Monitoring Program

Marigold has an extensive monitoring program in place for both groundwater quantity and quality and seasonal surface water quantity and quality. Results from this program as well as long-term trend data are reported to both state and federal agencies. Air, geochemical, vegetation, wildlife, and industrial health monitoring are also conducted regularly according to permit requirements. Agency representatives from the Nevada Division of Environmental Protection, Nevada Department of Wildlife, and Bureau of Land Management also conduct routine compliance inspections on a quarterly basis.

17.5 Reclamation and Closure

MMC engages in concurrent reclamation practices and is bonded for all permitted features, as part of the Nevada permitting process. Current bonding requirements are based on third-party cost estimates to reclaim all permitted features at the Property. Both the BLM and State of Nevada review and approve the bond estimate, and the BLM holds the financial instruments providing the bond backing.

State regulatory requirements mandate a formal closure plan be filed two years before the facility initiates closure. Both the BLM and State require a tentative closure plan as part of normal NEPA and operating permit requirements. Marigold has filed and maintained these closure plans, which, in conjunction with standard reclamation and re-vegetation of all disturbed areas, include discussions on removal of most infrastructure, monitoring, and notably long-term heap leach drain down solution management. Marigold’s currently approved closure plan describes a series of evapotranspiration cells to manage long-term solution drain down following an approximate two-year period of active solution volume reduction through evaporation.

Costs associated with all reclamation and closure activities are discussed in Section 18 and are reflected in the agency-approved bond amount.

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17.6 Community Relations and Social Responsibilities

There are currently no outstanding negotiations or social requirements regarding operations at the Property. The nature of NEPA and large-scale state permits involve public comment periods as well as public meetings. Recently held meetings generated minimal concern from the community, and local county government has been consistently supportive of continued mine operations at Marigold. There are no formal discussions required with local stakeholders or Native American tribal representatives, but mine management does meet informally to provide general updates and to discuss proposed donation/support requests.

Community support and engagement is well established at Marigold, and mine management provides regular updates with respect to the Property to local stakeholders and regulators.

17.7 QP Opinion

Following a review of the information supplied, the opinion of the QPs is that it is reasonable to rely on the information provided by SSR as outlined above for use in the Marigold21TRS because a significant environmental and social analysis has been conducted for the project over an extended period, the project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

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18 CAPITAL AND OPERATING COSTS

18.1 Introduction

Costs related to the development of reserves are based on a combination of historical site costs for fixed costs and a zero-base cost method for calculating variable costs. The variable costs are based on tonnage mined, tonnage processed, or hours worked for mining, maintenance, process and administration costs. The total planned spend is divided by tonnes mined for mining and maintenance unit costs, and ore tonnes stacked for process and administration unit costs.

18.2 Capital Costs

As the project has been in operation for a number of years, the level of project definition for the capital cost estimate is very high. Given the remaining capital scope, and the level of project definition, no contingency was included in the cost estimate. The QP considers the capital estimate to be in the accuracy range of +/-15%.

LOM project capital costs are summarized in Table 18.1

Table 18.1 Summary of Capital and Reclamation Costs

Capital Costs	Total<br>($M)
Exploration and Development	9.1
Sustaining Capital	348.9
Mine Development	10.3
Total Capital Costs	368.3
Reclamation	71.8
Total Capital and Reclamation	440.1

Sustaining capital costs include:

•Costs associated with ongoing expansion of the leach pad and associated process infrastructure represents about 8% of future capital.

The costs associated of reclamation and closure activities at Marigold were estimated to be $71.8M with the majority of the costs incurred from 2033 through to 2045.

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18.3 Operating Costs by Category

As the project has been in operation for a number of years, the level of project definition for the operating cost estimates is very high. Given the available project performance data and the high project definition, no contingency was included in the cost estimate. The QP considers the operating cost estimate to be in the accuracy range of +/-15%.

The LOM operating costs estimate is $10.05/t of processed ore. Operating costs per tonne for the LOM and next five years of operations are shown in Table 18.2.

Table 18.2 Summary of Operating Costs

Operating Costs		Total LOM (M)	/t Ore
	Years 1–5
Mining		1,469	7.53
Processing		373	1.57
Site Support		214	0.92
Total Operating Cost		2,056	10.01

All values are in US Dollars.

Totals may vary due to rounding

18.3.1 Mine Operating Costs

The LOM operating cost estimates include:

•Hauling

•Blasting

•Loading

•Road and dump maintenance

•Drilling

•Mine engineering and administration

•Maintenance

•Dewatering

Production Drilling

Depending on rock conditions, a combination of hammer drilling and rotary drilling is used at the Property.

The major operating cost categories for drilling in the LOM plan are labour, fuel and consumable supplies, including drill hammers, drill bits and drill steel.

Blasting

The major operating cost categories for blasting in the LOM are labour, ammonium nitrate, emulsion, contract blasting labour and support, and blasting accessories. These categories comprise more than 95% of the total blasting costs. Ammonium nitrate is the primary blasting agent, and, in areas of hard rock or when meteoric water exists from perched groundwater, rain or snow, an emulsion product is used as a supplementary blasting agent.

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The major operating cost categories for loading are labour, fuel for the PC7000 shovels, power for the P&H 4100 shovel, and ground engaging tools (GET), which include bucket teeth and other steel parts for all loading equipment.

Hauling

The major operating cost categories for hauling are labour, fuel, tires, and wear parts.

Roads And Dumps

The major operating cost categories for the support equipment fleet are labour, fuel, tires, dust suppression, and GET.

Maintenance

Mine maintenance costs associated with the LOM plan are for preventative and repair maintenance on the mine operations mobile production equipment and mobile support equipment. Maintenance costs are developed from historical data and planned work that is based on hours that the equipment accumulates during normal mining activities.

The major operating costs for mine maintenance are labour, maintenance repair parts, on-site contract labour, lube oils and greases, filters, hydraulic hoses, maintenance supplies, small tools and welding supplies.

Dewatering

Dewatering operating costs are mainly associated with (mains) power consumption for the dewatering wells.

Processing

Processing costs over the LOM include all costs required to recover the gold from the rock after it is mined and placed on the leach pad. This includes the cost of chemicals to process the ore, pumping costs to get the barren solution to the leach pad, pumping costs to get the pregnant solution to the carbon columns for gold recovery after it returns from the leach pad, and the costs associated with the extraction of the gold from the carbon to produce the final doré product shipped from Marigold.

A total of 85% of the operating costs for processing are labour, cyanide, lime, power, maintenance supplies and leach supplies. The remaining 15% of the costs are for other supplies, reagents and final off-site refining costs required to produce doré to a standard that meets the criteria defined by the London Bullion Market Association (LBMA).

The Marigold laboratory is under the direction of the Processing Department and operating costs include expenses associated with sampling, assaying and supplies related to leaching and refining.

18.3.2 G&A

G&A costs for the LOM include accounting and site administration, warehousing, safety, human resources, and environmental. These costs are related to supporting the operations groups in the mine, maintenance and processing departments.

The major operating cost for this group is labour, taxes, insurance, transportation expenses, and legal and audit expenses make up a large portion of the remaining costs.

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19 ECONOMIC ANALYSIS

19.1 Introduction

This economic analysis presents the key economic performance indicators for Marigold, including cash costs, all-in sustaining costs (AISC) and net present value (NPV), based on a 5% discount rate and mid-year cash flows approach. Cash flow projections commenced on 1 January 2022 and are estimated over the remaining LOM based on estimates of sales revenue, site production costs, capital expenditures, and other cash flows, including taxes and reclamation expenditures, all presented on a real cash flow basis.

This economic analysis is based on the latest Marigold LOM plan that excludes Inferred mineralization. Differences between the economic analysis quantities and the Mineral Reserve in Section 12 are the result of actual end of 2021 face positions used to report the Mineral Reserve being different to those predicted in the LOM plan. These differences are not considered material.

Marigold produces gold doré which is refined into gold bullion and, in turn, sold to bullion banks. The financial model includes recoverable gold on the leach pad and gold doré on hand as at 1 January 2022, all of which is sold over the remaining LOM. There is expected to be approximately 2,375 koz recoverable gold stacked over an active mining period of 11 years. LOM production includes an additional 161 koz payable gold that are on the leach pad as at 1 January 2022, for a total production of 2,536 koz payable gold. Reclamation is expected to continue for thirteen years after the last mining is completed. Gold production continues through 2038. The final reclamation occurs in 2045.

Cash inflows from sales assume all production within a period is sold, with minimal working capital movements, using a LOM average gold price of $1,647/oz.

The estimates for site production costs, sustaining capital and reclamation expenditures have been developed specifically for Marigold and are presented in earlier sections of the Marigold21TRS.

Based on SSRs projections as set forth in the Marigold21TRS, Marigold will incur LOM cash costs of $1,009 per payable ounce of gold sold and AISC of $1,154 per payable ounce of gold sold over the LOM to 2038. The after-tax NPV using a 5% discount rate and mid-year cash flows approach is $860M over the LOM.

Key project economic indicators are summarized in Table 19.1.

19.1.1 QP Opinion on Inputs

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Table 19.1 Marigold Key Economic Indicators

Item	Unit	Amount
Oxide Processed
Heap Leach Quantity Placed	Mt	205
Gold Feed Grade	g/t Au	0.48
Total Gold Produced
Total Payable Gold	koz	2,536
Gold Recovery	%	74.71
5-Year Annual Average
Total Payable Gold Produced	koz/yr	215
Free Cash Flow	$M/yr	72
Total Cash Costs (CC)	$/oz payable Au	1,042
All In Sustaining Costs (AISC)	$/oz payable Au	1,278
Key Financial Results
Life-of-Mine (LOM Total Cash Costs) CC	$/oz payable Au	1,009
LOM AISC	$/oz payable Au	1,154
LOM Site Operating Costs2	$/t treated	10.07
After-Tax NPV	$M	860
Discount Rate	%	5
Mine (processing) Life	years	17

1. Recovery includes impact of starting pad inventory

2. Includes operating cost plus $0.02/t refining cost

3. Differences between the Mineral Reserve and LOM quantities used in the economic analysis are due to differences between planned and actual 31 December 2021 face positions

19.2 Mine and Leaching Production Statistics

Mined material is either placed on the waste dumps or directly onto the leach pad over the course of 11 years of active mining. SSR has estimated its gold grades and recovery rates for each period to determine the recoverable ounces stacked. The annual production figures were obtained from the LOM plan. Total LOM production includes 164 koz recoverable gold that is on the leach pad as of 1 January 2022.

A summary of estimated mine production and gold production over the LOM is shown in Table 19.2.

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Table 19.2 Mining and Leaching Production

Year	Total Mined <br><br>(kt)	Ore Mined <br><br><br>(kt)	Au Grade <br><br><br>(g/t)	Waste Mined <br><br>(kt)	Gold Recovery <br><br>(%)	Recover-able Gold Stacked (koz)	Gold Produced <br><br>(koz)
2022	102,616	21,818	0.53	80,798	76.4%	282.2	230.0
2023	90,646	22,010	0.38	68,637	76.2%	203.6	260.1
2024	92,828	21,410	0.42	71,418	76.4%	218.3	200.4
2025	93,988	15,713	0.50	78,275	74.9%	190.1	201.9
2026	90,633	16,538	0.40	74,095	76.1%	161.3	182.3
2027	83,225	20,857	0.40	62,369	74.0%	198.5	138.3
2028	90,358	20,207	0.64	70,151	73.8%	305.5	302.7
2029	78,934	26,911	0.43	52,023	71.5%	265.5	278.6
2030	93,861	18,188	0.47	75,673	72.4%	198.2	220.3
2031	87,403	13,103	0.68	74,299	75.6%	215.6	209.8
2032	28,105	7,792	0.72	20,313	75.6%	136.0	162.1
2033	–	–	–	–	–	–	77.7
2034	–	–	–	–	–	–	29.9
2035	–	–	–	–	–	–	15.0
2036	–	–	–	–	–	–	10.0
2037	–	–	–	–	–	–	7.0
2038	–	–	–	–	–	–	10.0
Total	932,597	204,547	0.48	728,050	74.7%	2,374.9	2,535.9

1. Gold produced from 2033 onwards is derived from the residual recoverable gold remaining in the leach pad when mining is completed and is recovered through continued leaching from 2033 to 2038.

3. The LOM production quantities differ from the Mineral Reserve due to differences between planned mining and actual end of 2021 face positions.

4. Overall leaching recovery excludes impact of previously placed recoverable ounces.

5. Differences between the Mineral Reserve and LOM quantities used in the economic analysis are due to differences between planned and actual 31 December 2021 face positions.

6. Totals may vary due to rounding.

19.3 Sales and Refinery Process

The gold doré is poured at site and is transported by road via a secure vehicle to Asahi Refining USA, Inc. (Asahi) in Salt Lake City, Utah, which is approximately five hours away. SSR has entered into a non-exclusive refining agreement with Asahi, and the terms and conditions of this contract are within industry norms. The transportation and refining costs for the doré are also in accordance with industry standards.

Marigold or its agent sells all the gold (doré or refined bullion) to bullion banks.

19.4 Revenue

Annual revenue is determined by applying forecast metal prices to the estimated annual payable metal for each operating year. Sales prices have been applied to all LOM production without escalation.

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To determine the metal price assumptions used to calculate revenue, SSR reviewed consensus forecasts. Consistent with the financial modelling approach, these consensus forecasts and metal price assumptions are expressed in constant 2022 dollars.

Forecast metal prices used to inform the economic model are shown in Table 19.3.

Table 19.3 Forecast Metal Prices

Metal	Unit	Average	2022	2023	2024	2025	2026	>2026
Gold Price	$/oz	1,647	1,800	1,740	1,710	1,670	1,600	1,600
Silver Price	$/oz	21.56	24.00	23.00	22.00	21.00	21.00	21.00

19.5 Operating Costs

Operating costs for Marigold, which include mining, processing, and site support, have all been estimated. Unit LOM and 5-year operating costs are summarized in Table 19.4.

Table 19.4 Unit Operating Costs

Operating Costs		Total LOM (M)	/t Ore
	Years 1–5
Mining		1,469	7.53
Processing		373	1.57
Site Support		214	0.92
Total Operating Cost		2,056	10.01

All values are in US Dollars.

19.6 Royalties

Marigold is subject to a variety of NSR royalty payments, payable to various parties under the terms of the leases, as described in Section 3. The annual average NSR royalty payments range from 3.7% to 10.0%.

19.7 Cash Costs and AISC

Over the production life cash costs are estimated to average $1,009 per payable ounce of gold sold, and AISC is estimated to average $1,154 per payable ounce of gold sold as shown in Table 19.6.

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Table 19.5 Operating Costs per Payable ounce of Gold Sold

Operating Costs	Value<br>($/payable oz of gold sold)
Mine Operations	579
Processing	147
General and Administration	84
Inventory Adjustment	54
Royalties and Refining (net of silver credits)	144
Subtotal Cash Costs	1,009
Sustaining Capital	138
Development/Exploration	8
Total AISC	1,154

2. Payable ounces of gold sold over the LOM total 2,535.9 koz.

3. Totals may vary due to rounding.

4. Cash costs and AISC per payable ounce of gold sold are non-GAAP financial measures.

Average annual cash costs per payable ounce of gold sold range from $868 to $1,201 during the 11 years of active mining and leach pad stacking. Table 19.7 summarizes the cash costs and AISC over the LOM.

Table 19.6 Cash and AISC Unit Costs

Year	Cash Costs <br>($/payable oz of gold sold)	AISC <br>($/payable oz of gold sold)
2022	988	1,261
2023	913	1,233
2024	1,081	1,287
2025	1,116	1,318
2026	1,167	1,307
2027	1,201	1,449
2028	952	1,042
2029	868	937
2030	912	972
2031	919	990
2032	878	912

1. Cash costs include mine operations, processing, G&A, inventory adjustment, royalties and refining charges (net of silver credits). AISC includes cash costs and sustaining capital.

2. Totals may vary due to rounding.

3. Cash costs and AISC per payable ounce of gold sold are non-GAAP financial measures.

19.7.1 Taxation

Marigold is subject to Nevada Net Proceeds of Minerals Tax, Nevada property and sales taxes, and U.S. federal income tax. The economic analysis calculates these taxes in accordance with legislation enacted as at 1 January 2022. Property and sales taxes are accounted for in the operating costs of the mine.

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Nevada Net Proceeds of Minerals Tax

The State of Nevada imposes a 5% net proceeds tax on the value of all minerals extracted in the State. This tax is calculated and paid based on a prescribed net income formula applied only to income and expenses from mining, disallowing deductions for exploration and related-party financing costs. This tax is a deductible expense for U.S. federal income tax.

Nevada Excise Tax

Nevada Property Tax

Humboldt County assesses property tax on 35% of the total appraised value of Marigold’s real and personal property. The effective LOM current property tax rate is adjusted to reflect planned increases in the mine mobile fleet and leach pad areas. This property tax is a deductible expense for U.S. federal income tax.

Humboldt County Sales Tax

The Nevada sales tax rate for Humboldt County is 6.85%. Supplies and materials used in mining operations are taxed by the vendor at this rate. This sales tax is not recoverable but is a deductible expense for US federal income tax.

US Federal Income Tax

Federal income tax is determined under regulations that came into effect on 1 January 2022. Under these regulations, which removed alternative minimum tax, the mine is subject to a federal income tax rate of 21%.

19.8 Excluded Costs

Exploration costs unrelated to the delineation of existing Mineral Reserves have been excluded.

19.9 Sensitivity Analysis

The after-tax NPV calculation is based on the cash flows for the Property from and after 1 January 2022. Marigold is expected to generate $1,315M in pre-tax cash flow and $1,166M in after-tax cash flow over the LOM. The after-tax NPV using a 5% discount rate is $860M over the LOM.

Table 19.7 and Table 19.8 shows the results of sensitivity analysis from changes in discount rate, gold price, operating costs and sustaining capital. The cash flow used to evaluate Marigold is presented in Table 19.8.

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Table 19.7 Marigold21TRS Gold Price Sensitivity

After-Tax NPV ($M)	Long-Term Gold Price (/oz)
Discount Rate	1,000	1,350	1,400	1,600	1,750	1,800	2,000
Undiscounted	516	895	949	1,166	1,329	1,383	1,600
5%	472	796	842	1,027	1,166	1,212	1,397
10%	418	676	713	860	970	1,007	1,155
12%	350	530	556	659	737	762	866
15%	329	487	509	599	667	689	779

All values are in US Dollars.

Table 19.8 Marigold21TRS Gold Price Sensitivity

Variable		Change from Base NPV5% (M)
	-20%			–		10%		20%
Capital Cost		1,096	978		860		742		623
Operating Costs		915	887		860		832		805

All values are in US Dollars.

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Table 19.9 Cash Flow

Item
	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M	M
Revenue
Heap Leach - Gold Revenue
By-Product Revenue
Net Revenue
Realization Costs
Freight and Refining
Royalties
Total - Realization Costs
Operating Costs
Mining
Processing - Heap Leach
Site Support
Total - Site Operating Costs
Operating Surplus / (Deficit)
Capital Costs
Sustaining
Closure
Development/Exploration/Working and Other
Total - Capital Costs
Net Cash Flow <br>Before Tax
Tax
Net Cash Flow <br>After Tax

All values are in US Dollars.

Totals may vary due to rounding

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20 ADJACENT PROPERTIES

Marigold is located near the northern limits of a regional belt of ore deposits commonly referred to as the Battle Mountain-Eureka trend. This north–north-west striking alignment of mines and prospects. It is the second most prolific gold belt in Nevada after the Carlin trend, and it includes variants of Carlin-Type Gold Deposits (CTGD), distal type sediment hosted deposits as well as skarn and copper–gold porphyry systems.

Three major gold deposits lie adjacent to the SSR property. Nevada Gold Mines’ Phoenix mine is ~10km south-east of the Buffalo Valley property, i-80 Gold Corp’s Lone Tree mine is ~7 km north-east of Marigold, and Waterton Global Resource Management’s Converse project is ~6 km west of Marigold. There are also several inactive mines and exploration and/or development projects that can be found within a 19 km radius of the property.

Reported production and Mineral Resources for these adjacent properties are presented in

Table 20.1.

Table 20.1 Past Production and Mineral Resources for Adjacent Properties

Property		Owner		Years of Production	Gold Produced (Moz)		Stated Mineral Resources and Mineral Reserves (gold unless otherwise stated)
	Mineral Reserves		Measured and Indicated Mineral Resources			Inferred Mineral Resources
Phoenix1		Nevada Gold Mines		2006–Present	unknown		2.9 Moz gold (0.58 g/t)<br><br>840 Mlb copper @ 0.18%	5.28 Moz	0.34 Moz
Lone Tree Complex2		i-80 Gold Corp.		1991–2012	4.53		n/a	610 koz of Au @ 1.51g/t	2.76 Moz @ 1.6 g/t
Converse3		Waterton		–	–		n/a	6.12 Moz	0.59 Moz

1. Nevada Gold Mines, May 2021; Investor Day Presentation

2. i-80 Gold Corp., 2021; Technical Report, filed 21 October 2021

3. Chaparral Gold, 21 October 2014; website, deposit sold to Waterton Global Resource Management in 2014

Phoenix mine is currently operating by Nevada Gold Mines and is polymetallic Au-Cu-Ag porphyry system that has been in production since 2006. The mine includes various deposit types, all structurally controlled by north-west trending faults.

Lone Tree is considered a distal-disseminated deposit that may be genetically related to a porphyry-type system, mineralization was structurally controlled by north–north-west trending faults.

At Converse, gold mineralization is hosted within a skarn that developed within the Havallah Formation. No production has occurred at Converse to date.

A plan map of mine properties adjacent to Marigold is presented in Figure 20.1.

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Figure 20.1 Plan Map Showing Marigold Property Outline and Mineralization Relative to Adjacent or Nearby Mines or Published Deposits

The outer property boundary is shown as white outline

SSR, 2017

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21 OTHER RELEVANT DATA AND INFORMATION

There is no other relevant data or information

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22 INTERPRETATION AND CONCLUSIONS

The estimate of Mineral Resources presented in Table 11.11 and the estimate of Mineral Reserves presented in Table 12.1 were prepared for Marigold with an effective date of 31 December 2021.

The estimate of Mineral Resources was prepared using a domain-controlled, ordinary kriging technique with verified drillhole sample data derived from exploration activities conducted by various companies from 1968 to 2021.

Based on an evaluation of the available data from the Marigold mine, the authors of this report have drawn the following conclusions.

Possible areas of uncertainty that could materially impact the estimate of Mineral Reserves at Marigold include the commodity price assumptions, capital and operating cost estimates, estimation methodology, pit dewatering, and the geotechnical slope designs for the pit walls. These reasonably foreseeable impacts of the uncertainties in the cost, operations and estimation assumptions are discussed here:

•Commodity price assumptions: If the price of gold drops significantly below the cost of production for a significant period of time, it becomes uneconomic to extract the gold.

•Capital/operating cost estimates: If the operating cost of a major contributor to the operation, such as explosives, labour or fuel, increases more than has been reasonably estimated, the profit generated from the sale of gold ounces will decrease. And similarly, if the estimated capital cost to expand a heap leach pad or rebuild equipment, for example, is significantly more than anticipated, the additional capital input required may impact the profitability of the operation.

•Mineral Resource estimation methodology: The impact of the estimation methodology on the economic viability will be minimal because the applied methodology meets industry standards and has been verified by independent/external consultants.

•Mineral Reserves estimation methodology: The impact of the estimation methodology on the economic viability will be minimal because the applied methodology meets industry standards and has been verified by independent/external consultants and has been validated by historical reconciliation with mine production.

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•Geotechnical slope designs for pit walls: Marigold has operated for approximately 30 years, and the mining conditions and stable wall angles for the different rock types are well understood. There may be a risk that unidentified fault plane(s) and wall conditions encountered below the water table require the angle of a pit slope wall to be lowered to overcome potential multi-bench failure. Lowering the slope angle of the wall would mean that more waste material would need to be mined to reach the ore zone. Mining more waste than anticipated will increase the cost of production per ounce of gold and will negatively impact the project economics. Alternatively, ore defined as Mineral Reserves could be left in the ground unmined if the cost to remove overlying waste rock exceeds the value of the recoverable metal.

There are a number of active environmental permits at Marigold, and some degree of permit modification or renewal effort is typically always underway. The Marigold21TRS was prepared with the latest information regarding environmental and closure cost requirements and has indicated that work is in progress with regard to the renewal or extension of additional environmental permits.

The Marigold21TRS presents the LOM plan for Marigold as of 31 December 2021. Mining commenced on the Marigold deposit in 1988 with an expected mine life of eight years; now, approximately 30 years of continuous gold production later, the latest LOM plan still foresees an 11-year mine life. The future development for Marigold is planned as a large open pit ROM heap leach operation, which exploits Mineral Resources exceeding 5 Moz contained gold.

In total, the LOM plan states that Marigold will produce 2,536 koz of gold over an active mining period of 11 years with residual leaching over a further six years to 2038. LOM production includes an additional approximately 164 koz payable gold sold that is on the leach pad as at 1 January 2022.

Marigold will operate at an average total material movement rate of 250,000 tpd, or 90.4 Mtpa over the next ten years. Reclamation is expected to continue for an additional seven years following the last gold production. Going forward, operational efficiency and cost control measures remain key areas of focus for optimum margins, increasing Marigold’s medium to long-term potential and enabling the conversion of additional Mineral Resources into Mineral Reserves.

Based on SSR’s projections as set forth in the Marigold21TRS, Marigold will incur average annual cash costs of $1,009 per payable ounce of gold sold and AISC of $1,154 per payable ounce of gold sold over the processing LOM to 2038. The after-tax NPV using a 5% discount rate is $860M over the LOM.

Several optimization studies were initiated in 2017 to investigate opportunities to further increase Marigold’s operating efficiency. These studies include haulage profile optimization, expansion equipment studies and equipment productivity improvements. Indications from the operational excellence program over the past four years show improvements that have translated into improved per unit operating costs.

SSR has initiated exploration and Mineral Resources and Mineral Reserves development activities to enhance Marigold’s operating margins and extend the mine life. Further studies will examine the deep sulfide-hosted gold and could include further drilling evaluation and metallurgical testwork.

All QPs have reviewed the conclusions and agree with the findings of the Marigold21TRS.

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23 RECOMMENDATIONS

23.1.1 Processing

Pursue an upgrade to the barren pumping system in order to maintain a solution to ore ratio in excess of 1.5 as the leach pad increases height and new expansions are further from the barren ponds. An upgrade to the pumping system will aid in reducing the current WIP inventory and decrease the likelihood of building inventory in the future. Perform a trade -off study between CIC efficiency loss at high flow vs addition CIC trains with increased efficiency over the life of mine plan. The additional column trains to create a one-pass recovery system are a significant improvement to the system however there is still opportunity to optimize the flow rates through the columns and pull ounces forward through increased efficiency.

23.1.2 Metallurgy / Analytical

23.1.3 Mineral Resources

Incorporate geological data (from pit mapping) and hard boundaries (from faults that offset mineralization) into the resource model. Costs associated with this project are minor.

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23.1.4 Mine Planning

Develop and evaluate a digital twin of the mine haulage network utilizing industrial mathematics to iterate on material destinations over the LOM and optimize haulage profiles. Code projections of dewatering progress to the mine planning model. Record weekly plan variances, explanations, and associated actions for trending.

23.1.5 Exploration Drilling

23.1.6 Mine Operations

The Marigold operations team anticipates undertaking work focused on improving quality of ore delivered to leach pads and tactical fleet resourcing optimizations for improved cost efficiencies in the haulage cycles. To improve utilization of existing dispatch tools onsite as well as implementation of industrial mathematics-based haulage simulation tools for strategically optimized efficiencies throughout the LOM. Training of dispatch personnel for operation of updated fleet management systems onsite to optimise load / haul fleet resourcing and positively improve site productivity should be undertaken. Site will also deploy simulation software for strategic haulage network planning. This haulage simulator can be used to identify opportunities for mine planners and operations personnel to optimise material destinations.

23.1.7 Maintenance Operations

The Marigold maintenance teams is committed to remain focused on improved maintenance operations at the site with the aim of increasing equipment availabilities and reducing unit costs. Projects underway include disciplined work planning and execution and consumables wear optimization.

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24 REFERENCES

This section lists the references used in support of the Marigold21TRS.

AMEC Americas Ltd., 2014. Marigold Drill Hole Deviation Study. Memorandum, Prepared for Silver Standard Resources Inc., Dated 21 May 2014. (AMEC, 2014)

Call & Nicholas Inc, 2019. Slope Stability Study of the Red Dot Design, Marigold Mine, July 2019. (CNI, 2019a)

Call & Nicholas Inc,2019. Marigold Mine Site Visit Recommendations 6–7 July 2021, August 2019. (CNI, 2019b)

Call & Nicholas Inc, 2019. Analysis of Soil Slopes – H1, 5N1 and 5N2, October 2019. (CNI, 2019c)

Carver, J.N., Rathnam, K., Rice, T., and Yeomans, T.J., 2018. NI 43-101 Technical Report on the Marigold Mine, Humboldt County, Nevada, USA, 31 July 2018. (SSRTR18)

Cline, J.S., Hofstra, A.H., Muntean, J.L., Tosdal, R.M., and Hickey, K.A., 2005. Carlin-type gold deposits in Nevada: Critical geologic characteristics and viable models, Economic Geology 100th Anniversary Volume, p. 451–484. (Cline et al., 2005)

Cook, H.E., 2015. The Evolution and Relationship of the Western North American Paleozoic Carbonate Platform and Basin Depositional Environments to Carlin-type Gold Deposits in the Context of Carbonate Sequence Stratigraphy, in Pennel, W.M., and Garside, L.J., eds., New Concepts and Discoveries, Geological Society of Nevada Symposium Proceedings, Reno/Sparks, Nevada, May 2015, v. 1, p. 1-80. (Cook, 2015)

Cook, H.E. and Corboy, J.J., 2004. Great Basin Paleozoic carbonate platform; facies, facies transitions, depositional models, platform architecture, sequence stratigraphy, and predictive mineral host models; field trip guidebook; metallogeny of the Great Basin Project, 17–22 August 2003, 135 p. (Cook and Corboy, 2004)

Cook, H. E. and Taylor, M. E., 1977. Comparison of continental slope and shelf environments in the Upper Cambrian and lowest Ordovician of Nevada, in The Society of Economic Paleontologists and Mineralogists, Special Publication No. 25 pp. 51 – 81. (Cook and Taylor, 1977)

Cox, D.P., 1992. Descriptive model of distal-disseminated Ag-Au, U.S. Geological Survey Bulletin 2004, 19 p. (Cox, 1992)

Cox, D.P. and Singer, D.A., 1990. Descriptive and grade-tonnage models for distal-disseminated Ag-Au deposits: A supplement to U.S. Geological Survey Bulletin 1693, U.S. Geological Survey Open-File Report 90-282, 7 p. (Cox and Singer, 1990)

Doebrich, J.L. and Theodore, T.G., 1996. Geologic History of the Battle Mountain Mining District, Nevada, and Regional Controls on the Distribution of Mineral Systems in Coyner, A.R., and Faney, P.L., eds., Geology and Ore Deposits of the American Cordillera, Geological Society of Nevada, Symposium Proceedings, Reno/Sparks, Nevada, April 1995, pp. 453-483. (Doebrich and Theodore, 1996)

du Bray, E.A., 2007. Time, space, and composition relations among northern Nevada intrusive rocks and their metallogenic implications, Geosphere, v.3, p. 381–405. (du Bray, 2007)

Emsbo, P., Hofstra, A.H., Lauha, E.A., Griffin, G.L., and Hutchinson, R.W., 2003. Origin of high-grade gold ore, source of ore fluid components, and genesis of the Meikle and neighboring Carlin-type deposits, northern Carlin trend, Nevada, Economic Geology, v. 98, p. 1069– 1100. (Emsbo et al., 2003)

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Fithian, 2015. Geology, Geochemistry, and Geochronology of the Marigold Mine, Battle Mountain-Eureka Trend, Nevada, M.Sc. Thesis, Golden, Colorado, Colorado School of Mines, 120 p. (Fithian, 2015)

Fithian, M.T., Holley, E.A., and Kelly, N.M., 2018, Geology of gold deposits at the Marigold mine, Battle Mountain district, Nevada: Reviews in Economic Geology, v. 20, p. 121–155.

Glamis Gold Ltd. 2001. Glamis Marigold Mine Millennium Project Section 31 Technical Report and Reserve Summary for the Glamis Marigold Mine July 2001. (Revised). (Glamis Gold, 2001)

Grauch V.J.S., Rodriguez B. D., and Wooden J.L., 2003. Geophysical and Isotopic Constraints on Crustal Structure Related to Mineral Trends in North-Central Nevada and Implications for Tectonic History Economic Geology, April 2003, v. 98, pp. 269-286. (Grauch et al.,2003)

Hamilton, W., 1987. Crustal extension in the Basin and Range Province, southwestern United States, in Coward, M.P., Dewey, J.F., and Handcock, P.L., eds., Continental Extensional Tectonics, Geological Society Special Publication No. 28, pp. 155-176. (Hamilton, 1987)

Hofstra, A.H. and Cline, J.S., 2000. Characteristics and models for Carlin-type gold deposits, Reviews in Economic Geology, v. 13, p. 163–220 (Hofstra and Cline, 2000).

Ilchik, R.P. and Barton, M.D., 1997. An amagmatic origin of Carlin-type gold deposits, Economic Geology, v. 92, p. 269-288. (Ilchik and Barton, 1997)

Johnston, M.K. and Ressel, M.W., 2004. Carlin-type and distal disseminated Au-Ag deposits: Related distal expressions of Eocene intrusive centers in north-central Nevada in Controversies on the origin of World-class gold deposits, Part 1: Carlin-type gold deposits in Nevada, by J.L. Muntean, J. Cline, M.K. Johnston, M.W. Ressel, E. Seedorff, and M.D. Barton: Society of Economic Geologists Newsletter, v. 59, p. 12-14. (Johnston and Ressel, 2004)

Kester, M., 2015. On infrared absorption band position variation as a result of gold mineralization in a distal disseminated gold deposit at the Marigold Mine, Humboldt Co., Nevada, Senior Thesis, South Dakota School of Mines and Technology, 29 p. (Kester, 2015)

Ketner, K.B, 2008. The Inskip Formation, the Harmony Formation, and the Havallah Sequence of Northwestern Nevada—An Interrelated Paleozoic Assemblage in the Home of the Sonoma Orogeny. US Department of the Interior U.S. Geological Survey. Professional Paper 1757. (Ketner, 2008)

Ketner, K.B, 2013. Stratigraphy of Lower to Middle Paleozoic Rocks of Northern Nevada and the Antler Orogeny US Department of the Interior U.S. Geological Survey. Professional Paper 1799. (Ketner, 2013)

Knight Piésold Ltd. (KPL), 2014. Marigold Mine – Review of Rock Mechanic Considerations, dated 24 July 2014. REF. NO. NB101-201/24. Vancouver, British Columbia, Canada. (Knight Piésold, 2014)

Large, R.R., Bull, S.W., and Maslennikov, V.V., 2011. A carbonaceous sedimentary source-rock model for Carlin-type and orogenic gold deposits, Economic Geology, v. 106, p. 331–358. (Large et al., 2011)

Magee Geophysical Services LLC, 2014. Gravity Survey over the Marigold Mine Property, Humboldt County, Nevada. August 2014. (Magee, 2014)

Magee Geophysical Services LLC, 2014. Gravity Survey over the Marigold Mine Property, Humboldt County, Nevada. September 2016. (Magee, 2016)

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McKee, E.H., 2000. Potassium-argon chronology of Cretaceous and Cenozoic igneous activity, hydrothermal alteration, and mineralization, in Theodore, T.G., Geology of pluton-related gold Mineralization at Battle Mountain, Nevada, Monographs in Mineral Resource Science, no. 2, p. 121–143. (McKee, 2000)

McGibbon, D.H., 2004. Marigold Summary and Tour Guide (Internal report), pp. 1-3. (McGibbon, 2004)

McGibbon, D H., 2005, Geology of the Antler and Basalt Gold Deposits, Glamis-Marigold mine, Humboldt County, Nevada, in Rhoden, H.N., Steininger, R.C., and Vikre, P.G., eds., Geological Society of Nevada Symposium 2005: Window to the World, Reno, Nevada, May 2005, pp. 399–409. (McGibbon, 2005)

McGibbon, D.H., and Wallace, A.B., 2000. Geology of the Marigold Mine area: in Theodore, T.G., 2000, Geology of pluton-related gold mineralization at Battle Mountain, Nevada; Monographs in Mineral Resource Science No. 2: Center for Mineral Resources, Tucson, Arizona, pp. 222–240. (McGibbon and Wallace, 2000)

Muntean, J.L., Cline, J.S., Simon, A.C., and Longo, A.A., 2011. Magmatic-hydrothermal origin of Nevada’s Carlin-type gold deposits, Nature Geoscience, v. 4, p. 122–127. (Muntean et al., 2011)

Muntean, J.L. and Cline, J.S., 2018, Diversity of Carlin-Style Gold Deposits: Reviews in Economic Geology, v. 20, p. 1-5

OreWin, 2022. Marigold21TRS is the Marigold 2021 Project Update Report

Pells Sullivan Meynink, 2021. Geotechnical Review: Marigold Gold Project. Internal memo by PSM Consult Pty Limited. (PSM, 2021)

Piteau Associates, Marigold Mine Mackay Pit Dewatering System Design 4180-R03, January 2021. (Piteau, 2021)

Ressel, M.W., and Henry, C.D., 2006. Igneous Geology of the Carlin Trend, Nevada: Development of the Eocene Plutonic Complex and Significance for Carlin-Type Gold Deposits, Economic Geology, v. 101, p. 347-383. (Ressel and Henry, 2006)

Roberts, R.J., 1964. Stratigraphy and Structure of the Antler Peak Quadrangle, Humboldt and Lander Counties, Nevada: U.S. Geological Survey Professional Paper 459-A, 93 pp. (Roberts, 1964)

Roberts, R.J. 2002. A Passion for Gold, an Autobiography. University of Nevada press 1st Edition. Reid, R.F. (Roberts, 2002)

Sillitoe, R.H., and Bonham, H.F., 1990. Sediment-hosted gold deposits: Distal products of magmatic-hydrothermal systems, Geology, v. 18, p. 157–161. (Sillitoe and Bonham, 1990)

Theodore, T. G., 1991. Preliminary geologic map of the North Peak Quadrangle, Humboldt and Lander counties, Nevada. USGS Open-File Report. 91-429. (Theodore, 1991a)

Theodore, T. G., 1991. Preliminary geologic map of the Valmy Quadrangle, Humboldt and Lander counties, Nevada. USGS Open-File Report. 91-430. (Theodore, 1991b)

Theodore, T.G. (2000). Geology of Pluton-related Gold Mineralization at Battle Mountain, Nevada. Tucson, Arizona: centre for Mineral Resources, the University of Arizona. (Theodore, 2000)

21013Marigold21TRS220927Rev0.docx Page 181 of 184

Wallace, A.R., Ludington, S., Mihalasky, M.J., Peters, S.G., Theodore, T.G., Ponce, D.A., John, D.A., and Berger, B.R., 2004. Assessment of metallic mineral resources in the Humboldt River Basin, northern Nevada, U.S. Geological Survey Bulletin 2218, 309 p. (Wallace, 2004)

Waterton Global Resource Management website, 2018: www.watertonglobal.com

Wright, James L., 2016 Marigold Property Gravity Survey. (James L. Wright, 2016)

Wyld, S.J., Rogers, J.W., and Copeland, P., 2003. Metamorphic Evolution of the Luning-Fencemaker Fold-Thrust Belt, Nevada: Illite Crystallinity, Metamorphic Petrology, and 40Ar/39Ar Geochronology, The Journal of Geology, v. 111, p. 17-38. (Wyld et al., 2003)

Zoback, M.L., McKee, E.H., Blakely, R.J., and Thompson, G.A., 1994. The Northern Nevada rift: Regional tectonomagmatic relations and middle Miocene stress direction, Geological Society of America Bulletin, v. 106, p. 371-38.

21013Marigold21TRS220927Rev0.docx Page 182 of 184

25 RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT

The Marigold21TRS QPs have relied on the following information provided by SSR in preparing the findings and conclusions in this Technical Report Summary regarding the following aspects of modifying factors:

•Macroeconomic trends, taxes, royalties, data, and assumptions, and interest rates.

◦This has been used in Section 19 as described in this section. The QPs have relied exclusively on SSR for this information.

•Marketing information and plans within the control of the registrant.

◦This has been used in Sections 16 and 19 as described in those sections. The QPs have relied exclusively on SSR for this information.

•Legal matters outside the expertise of the qualified person, such as statutory and regulatory interpretations affecting the mine plan.

◦Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

•Environmental matters outside the expertise of the qualified person.

◦Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

•Accommodations the registrant commits or plans to provide to local individuals or groups in connection with its mine plans.

◦Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

•Governmental factors outside the expertise of the qualified person.

◦Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

Following a review of the information supplied, the opinion of the QPs is, that it is reasonable to rely on the information provided by SSR as outlined above for use in the Marigold21TRS, because a significant environmental and social analysis has been conducted for the project over an extended period, the project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

21013Marigold21TRS220927Rev0.docx Page 183 of 184

Document

Exhibit 96.3

Explanatory Note

SSR Mining Inc. (the “Company”) previously filed the Seabee 2021 Technical Report Summary (the “Seabee21TRS”), with an effective date of December 31, 2021 and issued on February 23, 2022, as Exhibit 96.3 to its Annual Report on Form 10-K for the year ended December 31, 2021, as amended. The Seabee21TRS has been amended to reflect certain revisions in compliance with Subpart 1300 of Regulation S-K, which revisions consist of adding confirmatory statements and other modifications that SSR does not consider material. The amended Seabee21TRS has been reissued as of September 29, 2022 and is presented with an effective date of December 31, 2021. The information in this amended Seabee21TRS has not been updated to reflect events, information or developments occurring after the effective date.

This page does not constitute a part of the amended Seabee21TRS.

Title Page

Project Name:	Seabee Gold Operation
Title:	Seabee 2021 Technical Report Summary
Location:	Saskatchewan, Canada
Effective Date of Technical Report Summary	31 December 2021
Effective Date of Mineral Resources:	31 December 2021
Effective Date of Mineral Reserves:	31 December 2021

Qualified Persons (QPs):

•Bernard Peters, BEng (Mining), FAusIMM (201743), employed by OreWin Pty Ltd as Technical Director - Mining, was responsible for the overall preparation of the Seabee21TRS and, the Mineral Reserve estimates, Sections 1 to 5; Sections 10; Section 12 to 25.

OreWin Pty Ltd ACN 165 722 574

140 South Terrace Adelaide 5000

P +61 8 8210 5600 E orewin@orewin.com W orewin.comi

Signature Page

Project Name:	Seabee Gold Operation
Title:	Seabee 2021 Technical Report Summary
Location:	Saskatchewan, Canada
Date of Signing:	29 September 2022
Effective Date of Technical Report Summary:	31 December 2021

/s/ Bernard Peters

Bernard Peters, Technical Director – Mining, OreWin Pty Ltd, BEng (Mining), FAusIMM (201743)

/s/ Sharron Sylvester

Sharron Sylvester, Technical Director – Geology, OreWin Pty Ltd, BSc (Geol), RPGeo AIG (10125)

21014Seabee21TRS220927Rev0.docx ii

TABLE OF CONTENTS

1EXECUTIVE SUMMARY	1
1.1Introduction	1
1.2Land Tenure and Ownership	2
1.3Property Description and Location	2
1.4Geological Setting and Mineralisation	3
1.5Exploration	3
1.5.1History	3
1.5.2Exploration Activities	4
1.6Development and Operations	5
1.7Processing and Recovery	5
1.7.1Reasonable Prospects for Eventual Economic Extraction	5
1.8Mineral Resources Estimate	6
1.9Mineral Reserves Estimate	7
1.10Metallurgy and Processing	8
1.11Environment, Communities, and Permitting	9
1.12Production	9
1.13Capital and Operating Costs	12
1.14Economic Analysis	12
1.15Interpretation and Conclusions	15
1.15.1Mineral Resources	15
1.15.2Mineral Reserves	15
1.16Recommendations	15
1.16.1Further Assessment	15
2Introduction	17
2.1Terms of Reference	17

21014Seabee21TRS220927Rev0.docx iii

2.2Qualified Persons	18
2.3Qualified Persons Property Inspection	18
2.4Units and Currency	18
2.5Effective Dates	18
3PROPERTY DESCRIPTION	19
3.1Location	19
3.2Ownership	20
3.3Mineral Tenure	20
3.4Underlying Agreements	25
3.5Environmental Considerations	25
3.6Permits and Authorisation	26
3.7Other Significant Factors and Risks	26
4ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	27
4.1Accessibility	27
4.2Physiography, Climate, and Vegetation	27
4.3Local Resources and Infrastructure	29
5HISTORY	30
5.1Previous NI 43-101 Technical Reports	31
6GEOLOGICAL SETTING, MINERALISATION, AND DEPOSIT	33
6.1Regional Geology	33
6.2District Geology	35
6.3Structural Setting	36
6.4Mineralisation	38
6.5Deposit Types	41
7EXPLORATION	42

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7.1Surficial Exploration	42
7.1.1Geochemistry	42
7.1.2Geophysical Surveys	43
7.1.3Airborne Magnetic and Radiometric Survey 2016	45
7.2Drilling	45
7.2.1Drilling by Cominco, Claude Resources, and Placer 1947–1988	47
7.2.2Drilling by Claude Resources 1989–2015	48
7.2.3Drilling by Claude Resources and SSR 2016	49
7.2.4Drilling by SSR Mining 2017 Onwards	49
7.3SSR Drilling Procedures	50
7.3.1Underground Drilling Procedures	50
7.3.2Surface Exploration Drilling Procedures	51
7.3.3Drill Sampling	52
7.4Density	54
8SAMPLE PREPARATION, ANALYSES, AND SECURITY	55
8.1Historical Samples	55
8.2Diamond Core Samples (1989 to Present)	55
8.3Chip and Muck Samples	55
8.4Quality Assurance and Quality Control Programmes	55
8.5Conclusions and Recommendations	56
9DATA VERIFICATION	57
9.1Verifications by SSR	57
9.2Verifications by OreWin	59
9.2.1Site Visit	59
9.2.2Verifications of Analytical Quality Control Data	60
9.2.3Discussion	60

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9.3QP Opinion	63
10MINERAL PROCESSING AND METALLURGICAL TESTING	64
10.1Style of Mineralisation	64
10.2Metallurgical Investigations	64
10.2.1Metallurgical Testwork	64
10.2.2Process Plant Improvements	64
10.3Recovery Estimates	66
10.4QP Opinion	67
11MINERAL RESOURCES ESTIMATES	68
11.1Resource Modelling Methods	68
11.1.1Santoy Mine	68
11.1.2Porky Deposit Area	69
11.2Cell Model Validation	70
11.3Mineral Resource Classification	70
11.4Reasonable Prospects for Eventual Economic Extraction	71
11.5Mineral Resource Statement	71
11.6Reconciliation	73
11.7Comparison with Previous Estimates	73
11.8QP Opinion	73
11.9Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects	73
12MINERAL RESERVES ESTIMATES	74
12.1Mineral Reserves Statement	75
12.2Comparison with Previous Estimates	75
12.3Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects	76
13MINING METHODS	77

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13.1Introduction	77
13.2Mining Methods	80
13.3Primary Access	81
13.4Level Design	81
13.5Material Handling	82
13.6Ventilation	82
13.7Backfill	82
13.8Dewatering	83
13.9Hydrology Considerations	83
13.10Geotechnical Considerations	83
13.10.1Rock Mass Quality and Rock Properties	83
13.10.2Stress Regime and Most Likely Mode of Failure	83
13.10.3Specific Geotechnical Risk	85
13.10.4Current Mitigation Measures Used to Minimise the Geotechnical Risk Support System	85
13.10.5Geotechnical Reports Review	86
13.11Mine Schedule	89
13.12Mobile Equipment	91
14PROCESSING AND RECOVERY METHODS	92
14.1General	92
14.2Crushing	95
14.3Grinding	95
14.4Gravity Recovery	95
14.5Cyanide Leaching	95
14.6Carbon-in-Pulp	95
14.7Carbon Elution and Electrowinning	95

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14.8Gold Refining	96
14.9Carbon Regeneration	96
14.10Mill Tailings	96
15INFRASTRUCTURE	97
15.1Site Access Roads	101
15.2Product Loadout	101
15.3Utilities	102
15.3.1Water	102
15.3.2Sewage Disposal	102
15.3.3Power	102
15.3.4Fuel Storage	102
15.3.5Explosives Storage	102
15.4Tailings Management Facilities	102
15.4.1East Lake Tailings Management Facility	103
15.4.2Triangle Lake Tailings Management Facility	103
15.5Waste Rock Structures	104
15.6Rock Quarry	105
15.7Water Facilities	105
16MARKET STUDIES	106
16.1Marketing and Metal Prices	106
16.2Contracts	106
16.3QP Opinion	106
17ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS	107
17.1Regulatory Setting	107
17.2Federal Environmental Assessment Process	107

21014Seabee21TRS220927Rev0.docx viii

17.3Provincial Environmental Assessment Process	108
17.4SGO Environmental Assessments	108
17.5Environmental Permits/Authorisations	109
17.6Environmental Considerations	109
17.7Mine Closure	110
17.8Social and Community Impact	113
17.9Safety	113
17.10QP Opinion	114
18CAPITAL AND OPERATING COSTS	115
18.1Capital Costs	115
18.2Operating Costs	116
19ECONOMIC ANALYSIS	117
19.1Economic Assumptions	117
19.1.1Pricing and Discount Rate Assumptions	117
19.1.2QP Opinion on Inputs	117
19.2Overview and Results	118
19.2.1Production and Cost Summary	118
19.2.2Project Financial Analysis	121
20ADJACENT PROPERTIES	126
21OTHER RELEVANT DATA AND INFORMATION	128
22INTERPRETATION AND CONCLUSIONS	129
22.1Mineral Resources	129
22.2Mineral Reserves Estimation	129
23RECOMMENDATIONS	130
23.1Further Assessment	130

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24REFERENCES	131
25RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT	135

TABLES

Table 1.1Summary of Seabee21TRS Mineral Resource Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price	6
Table 1.2Summary of Cut-off Values and Metallurgical Recoveries, of Seabee21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price	6
Table 1.3Summary of Seabee21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,600/oz Gold Price	7
Table 1.4Summary of Cut-off Values and Metallurgical Recoveries, of Seabee21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,600/oz Gold Price	8
Table 1.5Mining Production Statistics	10
Table 1.6LOM Average Operating Costs Estimate	12
Table 1.7Gold Price Assumptions	12
Table 1.8Key Economic Assumptions	13
Table 1.9Seabee21TRS Results Summary	13
Table 1.10Financial Results	14
Table 1.11After-Tax NPV Sensitivity to Gold Price and Discount Rates	14
Table 3.1Mineral Tenure Information – All Tenements 100% SGO Mining Inc. Owned	20
Table 5.1Historical Production from the SGO (1996–2021)	32
Table 6.1Key Stratigraphic and Structural Elements Controlling Mineralisation at the Seabee, Santoy, and Porky Deposits (SSR, 2017b)	40
Table 7.1Surface and Underground Drilling Completed on the SGO to 31 December 2021	46
Table 9.1Summary of Analytical QA/QC Data	61
Table 11.1Capping Values at Santoy Mine	69

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Table 11.2Parameters for Mineral Resource Classification	71
Table 11.3Summary of Seabee21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price	72
Table 11.4Summary of Cut-off Values and Metallurgical Recoveries of Seabee21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price	72
Table 11.5Annual Grade Reconciliation at Santoy for 2020 and 2021	73
Table 12.1Mineral Reserves Input Parameters	74
Table 12.2Summary of Seabee21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,600/oz Gold Price	76
Table 12.3Summary of Cut-off Values and Metallurgical Recoveries, of Seabee21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,600/oz Gold Price	76
Table 13.1Excavation Dimensions	81
Table 13.2Santoy Mine Dewatering Requirements	83
Table 13.3Summary of Testing Results for the Hangingwall Structure at Seabee Mine	84
Table 13.4Summary of Testing Results for the Footwall Structure at Seabee Mine	84
Table 13.5Summary of Testing Results for the Orezone Structure at Seabee Mine	84
Table 13.6Development, Waste Rock, and Backfill Summary	91
Table 14.1Seabee Mill Production Statistics 2006–2021	93
Table 16.1Seabee21TRS Economic Analysis Gold Price Assumptions	106
Table 18.1Capital Costs Estimate	115
Table 18.2LOM Average Operating Costs Estimate	116
Table 19.1Seabee21TRS Economic Analysis Gold Price Assumptions	117
Table 19.2Seabee21TRS Key Economic Assumptions	117
Table 19.3Seabee21TRS Results Summary	118
Table 19.4Production Statistics	119

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Table 19.5Cash Costs	121
Table 19.6Operating Costs and Revenues	122
Table 19.7Total Project Capital Cost	122
Table 19.8Financial Results	123
Table 19.9After-Tax NPV5% Sensitivity to Gold Price and Discount Rates	123
Table 19.10After-Tax NPV5% Sensitivity to Operating and Capital Cost Changes	123
Table 19.11Estimated Cash Flow	125

FIGURES

Figure 1.1Project Location Map	1
Figure 1.2Mining Production Profile	10
Figure 1.3Process Feed Profile	11
Figure 1.4Gold Production	11
Figure 1.5After-Tax Annual and Cumulative Cash Flow	14
Figure 3.1Location of the Seabee Gold Operation	19
Figure 3.2Seabee Corporate Structure	20
Figure 3.3SGO Mining Land Tenure Map	24
Figure 4.1Infrastructure at SGO and Typical Landscape of Project Area	28
Figure 6.1Cree Lake Zone and Reindeer Zone of the Trans-Hudson Orogen	33
Figure 6.2Regional Geology of the South-Western Trans-Hudson Orogen	34
Figure 6.3Local Geology Setting	36
Figure 6.4Integrated Structural Analysis of the SGO by SRK (2009) Based on Goldak’s 2007 Aeromagnetic Survey	38
Figure 6.5Typical Mineralisation Observed at the SGO	39
Figure 7.1Historical Rock Samples Collected at the SGO	42
Figure 7.2Historical Rock Samples Collected at the SGO	43

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Figure 7.3Map Showing the Distribution of Surface Drilling in Relation to the Seabee, Santoy and Porky Deposits, and other known Gold Occurrences on the Seabee Property	47
Figure 10.1Historical Annual Milled Daily Throughput	65
Figure 10.2Historical Annual Head and Tailings Grade Trend	65
Figure 10.3Mill Feed Grade and Gold Recovery 2017–2020	67
Figure 13.1Santoy Mine – Existing Development	78
Figure 13.2Santoy Mine – 2021 Life-of-Mine Development Design	79
Figure 13.3Santoy Mine – 2021 Life-of-Mine Stope Designs	80
Figure 13.42017 Structural Data	89
Figure 13.52021 Structural Data	89
Figure 13.6Production Plan Tonnage	90
Figure 13.7Processing Schedule	90
Figure 13.8Production Plan Recovered Gold Ounces	91
Figure 14.1Seabee Mill Flow Sheet	94
Figure 15.1Seabee Gold Operation Major Infrastructure	98
Figure 15.2Seabee Gold Operation Mill Site Infrastructure	99
Figure 15.3Seabee Gold Operation Tailings Management Facility Infrastructure	100
Figure 15.4Santoy Mine Major Infrastructure	101
Figure 19.1Production Plan Tonnage	119
Figure 19.2Processing Schedule	120
Figure 19.3Production Plan Recovered Gold Ounces	120
Figure 19.4Cumulative Cash Flow	124
Figure 20.1Location of the Six Properties included in the Taiga Gold Transaction with Respect to the Seabee Gold Operation	127

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1 EXECUTIVE SUMMARY

1.1 Introduction

The Seabee 2021 Technical Report Summary (Seabee21TRS) is an independent Technical Report Summary that has been in prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300) for SSR Mining Inc. (SSR), on the Seabee Gold Operation (SGO, or the Project).

The Project is located in Saskatchewan, Canada, at the northern end of Laonil Lake, approximately 125 km north-east of the town of La Ronge (Figure 1.1).

The SGO property hosts the Santoy mine, which has been in continuous commercial production since 2014. Commercial production at the now-depleted Seabee mine commenced in 1991 and was ceased in 2018.

Figure 1.1 Project Location Map

Google Earth, 2022

The SGO is directly owned (100%) by SSR through its wholly owned subsidiary company, SGO Mining Inc. (SGO Mining). SSR acquired the SGO on 31 May 2016 as a result of the acquisition of Claude Resources Inc.

21014Seabee21TRS220927Rev0.docx 1

The Seabee21TRS Qualified Persons (QPs) have reviewed the supplied data and information and accept this information as being accurate and complete and suitable for use in the Seabee21TRS. Information and data supplied by SSR that were outside the areas of expertise of the QPs and was relied upon when forming the findings and conclusions of this report are detailed in Section 25. Any individual or entity referenced as having completed work relevant to the Seabee21TRS, but not identified therein as a QP, does not constitute a QP for the Seabee21TRS.

The Seabee21TRS should be construed in light of the methods, procedures, and techniques used to prepare the Seabee21TRS. Sections or parts of the Seabee21TRS should not be read, or removed from, their original context.

1.2 Land Tenure and Ownership

The SGO is comprised of seven mineral leases and 102 mineral claims that cover an area of approximately 62,158 ha.

SGO Mining holds a 100% interest in the property.

1.3 Property Description and Location

Activities at the property are centred at approximately 55.7° latitude north and 103.5° longitude west.

Access to the SGO is by fixed-wing aircraft to the 1,275 m airstrip located on the property. During the winter months, a 60 km winter road is built between the mine site and Brabant Lake to transport heavy supplies and equipment by truck. Mining operations are conducted year-round.

The climate is borderline subarctic. Winters are long, dry, and cold (average –24°C) while summer is short, wet, and moderately warm (average +24°C). Precipitation is low, with an annual average of 486.2 mm.

The site is relatively flat, with much of the area comprised of irregular, hummocky, rocky exposures. Overburden soils are thin in this area, and often the rock outcrops are exposed.

21014Seabee21TRS220927Rev0.docx 2

1.4 Geological Setting and Mineralisation

The SGO is located within the northern portion of the Pine Lake greenstone belt. The belt has a strike length in excess of 50 km and comprises a variety of geochemically distinct tholeiitic mafic volcanic rocks formed in juvenile island arc settings, along with contemporaneous mafic intrusive rocks, volcaniclastics, sediments, and felsic intrusions of varying age. Metamorphic grade across the Pine Lake greenstone belt ranges from upper greenschist to upper amphibolite, with the SGO hosted in the latter. The belt has been complexly folded by at least four major phases of deformation that are observed across the SGO site and elsewhere in the Glennie domain of the Proterozoic Trans-Hudson Orogen.

The SGO can be subdivided into three main geological domains:

•The now-ceased Seabee mine is hosted within a coarsely layered mafic intrusion dominated by gabbro in the mine sequence.

•The Santoy mine area is hosted within a sequence of mafic volcano-sedimentary and intrusive rocks separated by generally north–south trending thrust faults.

•The Porky deposit area is a mineralised trend hosted along a 12 km long openly folded unconformity, separating arenaceous sedimentary rocks of the Rae Lake synform to the north from mafic volcanic rocks of the Seabee mine area to the south.

Gold mineralisation at the Santoy mine is hosted within calc-silicate altered shear structures with diopside-albite ±titanite-bearing quartz veins and occurs in gold-sulfide-chlorite-quartz veins in the shear zones, near or in the granodiorite and granite sills. Diopside-albite calc-silicate alteration facies are the main host to gold mineralisation in the Santoy 8A and Santoy 9A, 9B, and 9C zones. The Gap Hangingwall (GHW) deposit is hosted within a shallow dipping, north plunging, folded limb of the Lizard Lake Pluton. Mineralisation is concentrated near the fold hinge within centimetre to metre-scale quartz veining that strikes roughly north–south and dip sub-vertically.

At the Porky deposit, the brittle-ductile lode gold system is hosted along a thick corridor of calc-silicate altered mafic volcanics and arenaceous sedimentary rocks that straddle a major unconformity along the southern margin of the Rae Lake synform. Both the Porky Main and Porky West deposits are characterised by the same calc-silicate alteration package; however, the unconformity and arenites host most of the auriferous quartz veins at the Porky West deposit.

1.5 Exploration

1.5.1 History

The Laonil Lake region has been intermittently explored since the 1940s, with the first gold discovery made in 1947. Cominco conducted an extensive prospecting, geological mapping, trenching, and diamond drilling programme between 1947 and 1950, and in 1958 was granted 10 quartz mining leases covering the property on which the SGO is located. From 1974 through 1983, Cominco conducted detailed drilling and exploration, and in 1983 sold the property to BEC International Corporation (BEC). BEC subsequently sold the property to Claude Resources in 1985.

Claude Resources completed bulk sampling and drilling as part of a feasibility study for the Seabee deposit and reported a Mineral Reserve estimate in December 1988. Construction of the Seabee mill was completed in late 1991, and mining commenced in December 1991.

21014Seabee21TRS220927Rev0.docx 3

In 1998, prospecting and mapping identified several new discoveries including the Porky West zone in 2002, the Santoy 7 deposit in 2004, the Santoy 8 and Santoy 8 East deposits in 2005, and the Santoy Gap deposit in 2010.

Commercial production at the Santoy 7 deposit was achieved in 2007, and an economic study to evaluate the Mineral Resource at the Santoy 8 deposit was conducted in 2008. Portal construction and surface infrastructure development of the Santoy mine was initiated in late 2009, and environmental studies and permitting for commercial mining of the Santoy 8 and Santoy 8 East deposits were completed in 2010. Underground development continued in 2010, and the Santoy mine advanced towards commercial production in the second quarter of 2011.

On 31 May 2016, SSR acquired Claude Resources, thereby taking ownership of the SGO.

SSR filed the previous NI 43-101 Technical Report on the SGO in October 2017, with an effective date of 31 December 2016.

1.5.2 Exploration Activities

Since 1947, exploration at the SGO has comprised of surficial geochemical sampling, airborne and ground geophysical surveys, and extensive drilling. To 31 December 2021, drilling completed on the SGO property (by SSR and previous operators) includes:

•2,324 surface drillholes totalling 496,197 m and

•6,139 underground drillholes totalling 1,161,184 m.

Exploration surface drilling and infill surface and underground drilling completed by SSR since 2017 has been executed in the Carruthers, Herb Lake, Porky Main, Porky West, Seabee, and Santoy areas.

The objective of ongoing exploration conducted by SSR is to delineate, increase, and upgrade Mineral Resources. Underground drilling since 2016 focused on Santoy 8 and 9, GHW, and Santoy Hangingwall.

At the SGO, the 3-year budget calls for an average of 80 km of combined surface and underground drilling per year between 2022 and 2024. This drilling is for testing of targets to maximise Mineral Resource potential at the mine as SSR develops its long-term strategy for continuing to replenish its 3–5-year reserve inventory in the same way it has for more than 20 years, with particular focus on bringing higher grade zones on stream to displace lower tenor inventory that currently occurs in the schedule from 2024 onwards.

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1.6 Development and Operations

The life-of-mine (LOM) plan of the Mineral Reserve at the SGO, commencing 1 January 2022, includes 2.68 Mt at an average grade of 6.7 grams of gold per metric tonne (g/t Au). A total of 580 koz of gold will be delivered to the mill.

There is currently one operating mine as part of the SGO, that being Santoy. Mining will continue at the Santoy mine to provide feed to the mill located near the old Seabee mine.

Access underground at the Santoy mine is provided from the surface at the Santoy portal via a main ramp with sublevels spaced between 17–20 m vertically. Mining is carried out using sublevel open stoping mining methods with backfill. Stopes are filled with a combination of rock fill (RF) and cemented rock fill (CRF), mined in a bottom-up mining sequence. Sill pillars are mined on retreat once the stopes below and above have been mined (stopes above filled with CRF and allowed to cure). The mining sequence will continue to proceed in several longitudinally retreating, bottom-up advancing mining fronts. Current practice for material handling will remain with ore being truck hauled to surface and then hauled 14 km to the mill.

The major infrastructure at the SGO site includes roads and an airstrip, powerhouse and electrical distribution system, mill buildings and related services facilities, portal and ventilation raises, fuel storage, explosive storage, water supply and distribution, water management ponds and water treatment plant, tailings management facilities (TMFs), administrative buildings, and camp accommodation.

There are currently two TMFs that are being used by the mill: the East Lake TMF and the Triangle Lake TMF. Tailings deposition alternates between the two TMFs where winter deposition occurs in the Triangle Lake TMF and summer deposition is in the East Lake TMF. The remaining storage capacities of both facilities, based on the planned production rates, will potentially reach maximum capacity towards the end of 2030. To ensure that water treatment volumes are attained, a water treatment plant was constructed at East Lake TMF.

1.7 Processing and Recovery

SGO was originally developed based on bench scale metallurgical testwork that characterised the Seabee deposit as a lode gold style of mineralisation that was free milling and that would respond to a standard flow sheet employing gravity recovery and cyanidation. The Seabee deposit was processed for 25 years in the mill constructed immediately adjacent to the Seabee shaft and the plant is now used to process ore from the Santoy mine.

The mill flow sheet is a conventional crushing and grinding circuit employing gravity gold recovery and cyanide leaching with carbon-in-pulp for recovery and production of doré gold on site. The initial capacity was 500 tonnes per day (tpd), which was later expanded to 1,000 tpd with the addition of a third grinding mill.

Historical recovery at the Seabee mill was in the 94%–96% range, with routine low levels of losses both in the tailings solids and solution. Future recovery estimates are 98% and are based on the recent mill performance with mill recoveries of more than 98%. These improvements are attributed to the better condition of the leach equipment as well as the restored leach capacity.

1.7.1 Reasonable Prospects for Eventual Economic Extraction

The Mineral Resources in the Seabee21TRS were assessed for reasonable prospects for eventual economic extraction by reporting only material that fell within conceptual underground shapes and using a cut-off grade of 2.07 g/t Au that is based on a gold price of $1,750/oz.

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1.8 Mineral Resources Estimate

The Mineral Resources have been estimated for the SGO by the SSR technical department on site. The QPs reviewed the assumptions, parameters, and methods used to prepare the Mineral Resource Statement and is of the opinion that the Mineral Resources are estimated and prepared in accordance with S-K 1300.

The Mineral Resources are estimated based on cell models representative of the mineralised veins and using an assumed gold price of $1,750/oz.

The Mineral Resources estimates are based on all available data as of 31 December 2021. The Mineral Resources are reported exclusive of Mineral Reserves in Table 1.1 and Table 1.2.

Table 1.1 Summary of Seabee21TRS Mineral Resource Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price

Area		Mineral Resources Classification
	Measured				Indicated				Measured + Indicated				Inferred
	Tonnage<br>(kt)		Au<br>(g/t)		Tonnage<br>(kt)		Au<br>(g/t)		Tonnage<br>(kt)		Au<br>(g/t)		Tonnage<br>(kt)		Au<br>(g/t)
Santoy Mine		71		19.75		745		12.74		816		13.35		2,238		6.43
Porky West		–		–		52		5.03		52		5.03		516		4.42
Total SGO		71		19.75		797		12.23		869		12.85		2,754		6.05

1. Mineral Resources are reported based on 31 December 2021 as-mined survey data.

2. Mineral Resources are reported exclusive of Mineral Reserves.

3. Mineral Resources are shown on a 100% basis.

4. The Mineral Resources estimates are based on a 2.07 g/t Au cut-off with a gold price assumption of $1,750/oz.

5. Santoy Mine includes Santoy 8, Santoy 9, and GHW lodes.

6. The Mineral Resources in the Seabee21TRS were assessed for reasonable prospects for eventual economic extraction by reporting only material that fell within conceptual underground shapes.

7. SSR has 100% ownership of the Project.

8. The point of reference for Mineral Reserves is the point of feed into the processing facility.

9. Tonnage is metric tonnes and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

Table 1.2 Summary of Cut-off Values and Metallurgical Recoveries, of Seabee21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price

Mineral Resources <br>Classification	Tonnage<br><br>(kt)	Au<br><br>(g/t)	Contained Gold<br><br>(koz)	Cut-off<br><br>(Au g/t)	Metallurgical Recovery<br><br>(%)
Measured	71	19.75	45	2.07	98
Indicated	797	12.23	313	2.07	98
Measured + Indicated	869	12.85	359	2.07	98
Inferred	2,754	6.05	536	2.07	98

1. Mineral Resources are reported based on 31 December 2021as-mined survey data.

2. Mineral Resources are reported exclusive of Mineral Reserves.

3. Mineral Resources are shown on a 100% basis.

4. The Mineral Resources estimates are based on a 2.07 g/t Au cut-off with a gold price assumption of $1,750/oz.

5. Santoy Mine includes Santoy 8, Santoy 9, and GHW lodes.

6. The Mineral Resources in the Seabee21TRS were assessed for reasonable prospects for eventual economic extraction by reporting only material that fell within conceptual underground shapes.

7. SSR has 100% ownership of the Project.

8. The point of reference for Mineral Reserves is the point of feed into the processing facility.

9. Tonnage is metric tonnes, ounces represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

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1.9 Mineral Reserves Estimate

The SGO Mineral Reserves estimate was completed by the SSR technical department on site. The QPs reviewed the assumptions, parameters, and methods used to prepare the Mineral Reserve Statement and is of the opinion that the Mineral Reserve is estimated and prepared in accordance with S-K 1300.

The Mineral Reserve Statement is reported in Table 1.3 and Table 1.4. The reference point at which the Mineral Reserve is identified is where ore is delivered to the processing plant (i.e., mill feed). The QPs are unaware of any environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant issues that may materially affect the Mineral Reserve estimate. However, the Mineral Reserve may be affected by further infill and exploration drilling that may result in increases or decreases in subsequent Mineral Resource and Mineral Reserve estimates. The Mineral Reserve may also be affected by subsequent assessments of mining, environmental, processing, permitting, taxation, socio-economic, and other factors. The effective date of the Mineral Reserve Statement is 31 December 2021.

Table 1.3 Summary of Seabee21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,600/oz Gold Price

Area		Mineral Reserve Classification
	Proven				Probable				Total
	Tonnage<br>(kt)		Grade (Au g/t)		Tonnage<br>(kt)		Grade (Au g/t)		Tonnage<br>(kt)		Grade (Au g/t)
Santoy Mine		304		9.16		2,379		6.40		2,684		6.72

1. Mineral Reserves are reported based on 31 December 2021as-mined survey data.

2. The Mineral Reserves estimates are based on metal price assumptions of $1,600 gold.

3. The Mineral Reserves estimates are reported at a cut-off grade of 2.52 g/t Au.

4. Economic analysis for the Mineral Reserves has been prepared using long-term metal prices of $1,600/oz gold.

5. No mining dilution is applied to the grade of the Mineral Reserves. Dilution intrinsic to the Mineral Reserves estimate is considered sufficient to represent the mining selectivity considered.

6. Processing recoveries vary based on the feed grade. The average recovery is estimated to be 98%.

7. SSR has 100% ownership of the Project.

8. Santoy Mine includes Santoy 8, Santoy 9, and Gap Hangingwall lodes.

9. Metals shown in this table are the contained metals in ore mined and processed.

10. The point of reference for Mineral Resources is the point of feed into the processing facility.

11. Tonnage is metric tonnes and g/t represents grams per metric tonne.

12. Totals may vary due to rounding.

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Table 1.4 Summary of Cut-off Values and Metallurgical Recoveries, of Seabee21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,600/oz Gold Price

Mineral Reserve Classification	Tonnage<br><br>(kt)	Grade <br><br>(Au g/t)	Contained Gold<br><br>(koz)	Cut-off <br>Value<br><br>(Au g/t)	Metallurgical Recovery<br><br>(%)
Proven	304	9.16	90	2.52	98
Probable	2,379	6.40	490	2.52	98
Total Mineral Reserves	2,684	6.72	580	2.52	98

1. Mineral Reserves are reported based on 31 December 2021as-mined survey data.

2. The Mineral Reserves estimates are based on metal price assumptions of $1,600 gold.

3. The Mineral Reserves estimates are reported at a cut-off grade of 2.52 g/t Au.

4. Economic analysis for the Mineral Reserves has been prepared using long-term metal prices of $1,600/oz gold.

5. No mining dilution is applied to the grade of the Mineral Reserves. Dilution intrinsic to the Mineral Reserves estimate is considered sufficient to represent the mining selectivity considered.

6. Processing recoveries vary based on the feed grade. The average recovery is estimated to be 98%.

7. SSR has 100% ownership of the Project.

8. Metals shown in this table are the contained metals in ore mined and processed.

9. The point of reference for Mineral Reserves is the point of feed into the processing facility.

Tonnage is metric tonnes, ounces represent troy ounces, and g/t represents grams per metric tonne.

11. Totals may vary due to rounding.

The 2021 Mineral Reserves are a net increase of 86 koz (18%) total contained gold ounces as compared with the 2020 Mineral Reserves. Although mining depletion has occurred in the Santoy 8A and 9A mining zones, the 2021 Mineral Reserve has increased with the conversion of the Santoy Mineral Resources in the GHW zone into Mineral Reserves. An increase in the gold commodity price has also resulted in a decrease in the Mineral Reserve cut-off grade.

1.10 Metallurgy and Processing

The Seabee Gold Operation (SGO) was originally developed based on bench scale metallurgical testwork that characterised the Seabee deposit as a lode gold style of mineralisation that was free milling and that would respond to a standard flow sheet employing gravity recovery and cyanidation. After the successful commissioning of the Seabee mill and the operation matured the mill became the reference flow sheet for other mineralisation that was identified as a possible mill feed source.

The SGO deposits, are classified as lode gold style deposits with the gold in quartz veins typically in shear zones with some variations of the host rock mineralisation, with gabbros at Seabee and mafic metavolcanics at the Santoy and Porky deposits. As the satellite deposits advanced to potential development, bench scale testing was employed to confirm the free milling potential and the presence of any deleterious elements.

With the consistent long-term metallurgical response of the Seabee and Santoy deposits processed to-date, the focus of metallurgical investigations has been on improvements to process capacity constraints and process operating cost reductions.

The Seabee process plant was originally built as a 500 tpd operation. Subsequent capital projects have included the addition of the primary ball mill, addition of a second Knelson concentrator and Acacia gravity gold recovery. Process improvements have included, improved grind size control, improved gravity circuit utilisation, improved leach feed thickener chemistry and reduction in flocculant addition, and carbon and cyanide management.

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The Seabee operation is characterised by coarse gold making the gravity recovery circuit critical to the overall gold recovery of the process plant. Historically gravity recovery was approximately 40%. In recent years with incorporation of gravity circuit improvements including the Acacia circuit gravity gold recovery has improved to 60%–70% of recovered gold, with the CIP accounting for 30%–40%. Overall gold recovery is estimated at 97%–98.5%.

1.11 Environment, Communities, and Permitting

SSR has successfully completed three environmental assessments for the SGO to date. The site is regulated by both the Saskatchewan Ministry of Environment and Environment and Climate Change Canada. In accordance with provincial environmental legislation and regulations, the operation must adhere to the terms and conditions of an Approval to Operate a Pollutant Control Facility (Approval to Operate). The SGO is in compliance with all the terms and conditions of its current Approval to Operate number PO19-193, issued in October 2019 with an expiry date of September 2022. SSR is responsible to apply to renew this Approval to Operate a minimum of 90 days prior to the expiry date.

The dominant environmental liability at the SGO is the management of the mill tailings and associated tailings effluent. Appropriate infrastructure and operational plans are in place to reduce operational and closure risks associated with these liabilities to acceptable levels.

In 2016 SSR initiated a thorough stakeholder engagement plan designed to strengthen its relationship with communities impacted by the SGO, and the existing social licence to continue operations of the facility. No significant public concern with the SGO was expressed during the stakeholder engagement process.

There are no known environmental concerns at the SGO that cannot be successfully mitigated through the implementation of the various approved management plans that have been developed based on accepted scientific and engineering practices.

In accordance with provincial regulations, SSR has submitted an updated decommissioning and reclamation plan and cost estimate every five years, since 1996. Following initial regulatory review and subsequent edits by SSR, the 2020 revision to the preliminary decommissioning and reclamation plan was approved by the Ministry of Environment. Work on a revision is currently underway to cover the expanded Triangle Lake TMF. The total cost to implement the closure plan using a third-party contractor is currently C$12.0M. This cost estimate incorporates costs to cover release of the property, following the successful implementation of the closure plan, back to the province by way of Saskatchewan’s Institutional Control Program.

1.12 Production

Future proposed mine production has been scheduled to optimise the mine output and meet the plant capacity.

The mining production forecasts are shown in Table 1.5.

Mine, process, and metal production are shown in Figure 1.2 through Figure 1.4.

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Table 1.5 Mining Production Statistics

Item	Unit	Total LOM	2-Year Annual Average	LOM Annual Average
Gold Feed – Tonnes Processed
Quantity Ore Tonnes Treated	kt	2,684	424	424
Au Feed Grade	g/t	6.72	9.34	6.72
Gold Recovery	%	98.0	98.0	98.0
Metal Produced
Gold	koz	568	125	90

Figure 1.2 Mining Production Profile

OreWin, 2021

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Figure 1.3 Process Feed Profile

OreWin, 2021

Figure 1.4 Gold Production

OreWin, 2021

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1.13 Capital and Operating Costs

The cost estimate was prepared by the SSR technical department at both the SGO site and Saskatoon office. The QPs reviewed the assumptions, parameters, and methods used to prepare the cost estimate and is of the opinion that they are sufficient for the purposes of validating the economics of the Mineral Reserves. Total capital expenditure is estimated to be $162M.

The life of mine operating costs are approximately $155/t of ore milled, as summarised in Table 1.6.

Table 1.6 LOM Average Operating Costs Estimate

Cost Component	$/t milled
Mining	46
Surface Haulage	6
Milling (incl. Fixed Plant)	35
G&A	68
Total Operating Cost	155

Sum of individual values may not match total due to rounding

1.14 Economic Analysis

The estimates of cash flows have been prepared on a real basis as at 1 January 2022 and a mid-year discounting is used to calculate NPV.

The projected financial results include:

•After-tax NPV at a 5% real discount rate is $249M

•Mine life of six years

The estimated total cash costs for the first two years of production is $538 per payable ounce of gold, with a LOM average of $735. The all-in sustaining costs (AISC), which includes infrastructure capital and capital development, is $868 per payable ounce of gold for the first two years of production, with a LOM average of $1,021.

The gold prices assumptions used for the economic analysis are shown in Table 1.7. Gold provides the only revenue included in the analysis.

Table 1.7 Gold Price Assumptions

Commodity	Unit	2022	2023	2024	2025	Long- Term
Gold	$/oz	1,800	1,740	1,710	1,670	1,600

Other key economic assumptions for the discounted cash flow analyses are shown in Table 1.8.

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Table 1.8 Key Economic Assumptions

Model Assumption	Unit	Value
Refinery Charge	$/oz gold	0.45
Gold Payability	%	99.5
Tax Rate	%	25.9

The key results of the Seabee21TRS are summarised in Table 1.9. The projected financial results for undiscounted and discounted cash flows, at a range of discount rates are shown in Table 1.10. The estimates of cash flows have been prepared on a real basis as 1 January 2022 and a mid-year discounting is used to calculate net present value (NPV).

The results of NPV5% sensitivity analysis to a range of changes in gold price and discount rates is shown in Table 1.11. A chart of the cumulative cash flow is shown in Figure 1.5.

Table 1.9 Seabee21TRS Results Summary

Description	Unit	Total LOM
Ore Processed
Ore Tonnes Treated	kt	2,684
Au Feed Grade	g/t	6.72
Gold Recovery	%	98.0
Metal Produced
Gold	koz	568
Key Financial Results
Site Operating Costs	$/t milled	155
Mine Site Cash Cost	$/oz payable gold	734
Royalties and Refining	$/oz payable gold	0.5
Total Cash Costs (CC)	$/oz payable gold	735
All-in Sustaining Costs (AISC)	$/oz payable gold	1,021
Average Gold Price	$/oz payable gold	1,701
NPV	$M	249
Discount Rate	%	5
Project Life	years	6

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Table 1.10 Financial Results

Discount Rate		NPV (M)
	Before-Tax
Undiscounted		372	274
2%		358	263
5%		338	249
10%		309	228
12%		299	221
15%		285	211
18%		273	201
20%		265	196

All values are in US Dollars.

Table 1.11 After-Tax NPV Sensitivity to Gold Price and Discount Rates

Discount Rate		Relative Gold Price(/oz)
	–400			–200		–100		–		+100		+200		+300		+400
Undiscounted		106	148		190		232		274		316		358		400		442
2%		104	144		184		224		263		303		343		383		422
5%		101	138		175		212		249		286		323		360		396
10%		96	129		162		195		228		261		294		327		359
12%		94	126		158		189		221		252		284		315		347

All values are in US Dollars.

Figure 1.5 After-Tax Annual and Cumulative Cash Flow

OreWin, 2021

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1.15 Interpretation and Conclusions

1.15.1 Mineral Resources

Mineral Resources for the Seabee21TRS have been estimated in and prepared in accordance with S-K 1300.

Areas of uncertainty that may materially impact the Mineral Resource estimates include:

•Assumptions used to generate the data for consideration of reasonable prospects of eventual economic extraction for the Seabee deposit.

GHW mining recovery could be lower, and dilution increased. Early stoping in GHW should be used to confirm mining method parameters for the GHW zone in terms of costs, dilution, and mining recovery. Early development will also provide access to data and metallurgical samples at a bulk scale that cannot be collected at the scale of a drill sample.

•Environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues.

•Commodity prices and exchange rates.

•Cut-off grades.

1.15.2 Mineral Reserves

Mineral Reserves for the Seabee21TRS have been estimated in and prepared in accordance with S-K 1300.

Areas of uncertainty that may impact the Mineral Reserve estimate include:

•Any changes to the resource model as a result of further definition drilling at the site.

•Changes to mining conditions that have an impact to operating costs, production rates or mining recovery factors.

•Commodity prices and exchange rates.

1.16 Recommendations

The QPs are not aware of any significant risks and uncertainties that could be expected to affect the reliability or confidence in the information discussed herein.

1.16.1 Further Assessment

The key areas for further studies / work are:

•Ongoing drilling to expand the Mineral Resource aimed to increase mine life and optimise grade in years 2024 and beyond, as Seabee has managed to do for many years.

•Ongoing geotechnical drilling and logging will be required to increase the confidence in geotechnical data as the project develops.

•Ongoing geotechnical mapping should take place at regular intervals in the planned developments to verify the rock mass conditions determined and to assess the rock mass quality where there is currently little information. This will also allow for the identification of localised weak zones and potentially unstable wedges which should be appropriately supported.

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•While the structural analysis provides an impression of the major joint sets across the project area, further geotechnical scanline mapping should be conducted regularly as mining commences to allow for the identification of low angle joints in the hangingwall, localised joint sets and for potential wedges and instabilities.

•Update the Santoy geotechnical model to include the expanded GHW mining zone.

•Early stoping in GHW should be used to confirm mining method parameters for the GHW zone in terms of costs, dilution, and mining recovery. Early development will also provide access to data and metallurgical samples at a bulk scale that cannot be collected at the scale of a drill sample.

•Update site standard operating procedures to include a more transparent Mineral Resource and Mineral Reserve process, clearly documenting the key input parameters applied, and an audit trail of approvals for each phase of the work performed.

•Implementation of Operational Excellence projects identified based on SSR’s recent operational review may present incremental improvements to production and operating costs.

•Continue with ongoing review of capital and operating cost estimates and performance and productivity tracking.

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2 INTRODUCTION

The Seabee21TRS has been prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300).

The Seabee Gold Operation (SGO, or the Project) is owned and operated by SGO Mining Inc., a wholly owned subsidiary of SSR. In most cases, the parent company will be referred to as SSR throughout this Technical Report Summary. SSR has reported that the total cost of the gross mineral properties, plant and equipment as of 31 December 2021 was $465.5M.

2.1 Terms of Reference

The Seabee21TRS is an independent Technical Report Summary (TRS) on the Project, prepared for SSR by the Seabee21TRS Qualified Persons (QPs). The Seabee21TRS is based on information and data supplied to the QPs by SSR and other parties where necessary. Any individual or entity referenced as having completed work relevant to the Seabee21TRS, but not identified therein as a QP, does not constitute a QP. Seabee21TRS QPs have reviewed the supplied data and information and it appears accurate and complete and accept this information for use in the Seabee21TRS. The primary source of data for the Seabee21TRS is the Seabee 2021 Project Update.

Section 25 describes any information and data supplied by SSR that was outside the areas of expertise of the QPs and was relied upon when forming the findings and conclusions of this report.

The QPs have used their experience and industry expertise to produce the estimates and approximations in the Seabee21TRS. It should be noted that all estimates and approximations contained in the Seabee21TRS will be prone to fluctuations with time and changing industry circumstances.

The purpose of this Seabee21TRS is to report the Mineral Resources and Mineral Reserves for the project. This report is a Feasibility Study (FS) that represents forward-looking information. The forward-looking information includes metal price assumptions, cash flow forecasts, projected capital and operating costs, metal recoveries, mine life and production rates, and other assumptions used in the FS. Readers are cautioned that actual results may vary from those presented. The factors and assumptions used to develop the forward-looking information, and the risks that could cause the actual results to differ materially are presented in the body of this report under each relevant section.

The conclusions and estimates stated in the Seabee21TRS are to the accuracy stated in the Seabee21TRS only and rely on assumptions stated in the Seabee21TRS. The results of further work may indicate that the conclusions, estimates and assumptions in the Seabee21TRS need to be revised or reviewed.

The Seabee21TRS should be construed in light of the methods, procedures, and techniques used to prepare the Seabee21TRS. Sections or parts of the Seabee21TRS should not be read in isolation of, or removed from, their original context.

21014Seabee21TRS220927Rev0.docx 17

The Seabee21TRS is intended to be used by SSR, subject to the terms and conditions of its contract with OreWin. Recognising that SSR has legal and regulatory obligations, OreWin has consented to the filing of the Seabee21TRS with US SEC. Except for the purposes legislated, any other use of this report by any third party is at that party's sole risk.

A list of the references used to prepare the Seabee21TRS is provided in Section 24.

2.2 Qualified Persons

The following people served as the QPs as defined in subpart 1300 of US Regulation S-K Mining Property Disclosure Rules (S-K 1300):

•Bernard Peters, BEng (Mining), FAusIMM (201743), employed by OreWin Pty Ltd as Technical Director – Mining, was responsible for the overall preparation of the Seabee21TRS and, the Mineral Reserve estimates, Sections 1 to 4; Sections 5 and 6; Section 13; and Sections 15 to 27.

2.3 Qualified Persons Property Inspection

OreWin personnel, Sharron Sylvester Technical Director – Geology and QP, and Graeme Baker Principal Mining Consultant visited the Project 6 February 2020. The site visit included briefings from mining, geology, and exploration personnel; site inspections of potential areas for mining, including underground; discussions with staff; and review of the existing infrastructure and facilities around the Project site.

Bernard Peters has not visited the site due to travel restrictions.

2.4 Units and Currency

This Report uses metric measurements except where otherwise noted. The currency used is US dollars ($) unless otherwise stated.

2.5 Effective Dates

The report has a number of effective dates, as follows:

•Effective date of the Technical Report Summary: 31 December 2021

•Drillhole database close-out date for Mineral Resource estimate: 15 November 2020

•Effective date of Mineral Resources: 31 December 2021

•Effective date of Mineral Reserves: 31 December 2021

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3 PROPERTY DESCRIPTION

3.1 Location

The Seabee Gold Operation (SGO) is located at the northern end of Laonil Lake, approximately 125 km north-east of the town of La Ronge, in Saskatchewan, Canada (Figure 3.1). The centre of the property is located at approximately 55.7° latitude north and 103.5° longitude west.

The mine is a remote operation with access to the mine site by fixed wing aircraft to a 1,275 m airstrip located on the property. Equipment and major resupply items are transported to the site via a 60 km winter ice road, which is typically in use from end of January through to the end of March.

Figure 3.1 Location of the Seabee Gold Operation

SSR, 2017

SGO has been in continuous operation since 1991. Ore is currently produced from the Santoy underground mine from a ramp access / surface portal and is hauled 14 km to the mill located at the Seabee site. A second underground mine, also having ramp access, was operated from 1991–2018 at Seabee.

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3.2 Ownership

SSR Mining Inc. (SSR) holds a 100% interest in the property through its wholly-owned subsidiary, SGO Mining Inc. (SGO Mining). SSR acquired the SGO on 31 May 2016 as a result of the acquisition of Claude Resources Inc.. The structure is shown in Figure 3.2.

SSR is a gold mining company with four producing assets, located in the USA, Turkey, Canada, and Argentina, and with development and exploration assets in the USA, Turkey, Mexico, Peru, and Canada. SSR is listed on the NASDAQ (NASDAQ:SSRM), the Toronto Stock Exchange (TSX:SSRM), and on the Australian Stock Exchange (ASX:SSR).

Figure 3.2 Seabee Corporate Structure

3.3 Mineral Tenure

The SGO is comprised of seven mineral leases and 102 mineral claims that cover an area of approximately 62,158 ha (Table 3.1 and Figure 3.3).

SSR holds a 100% interest in the property through its wholly owned subsidiary, SGO Mining.

Table 3.1 Mineral Tenure Information – All Tenements 100% SGO Mining Inc. Owned

Area	Tenement Number	Expiry Date	Area<br>(ha)
Seabee Area	CBS 7058	08 May 2031	1,230
	CBS 7076	31 May 2031	856
	ML 5535	01 July 2035	45
	ML 5536	01 August 2025	50
	ML 5543 iP	24 January 2033	86

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	ML 5551 iP	31 December 2024	115
	ML 5557	01 February 2029	42
	ML 5558 iP	01 February 2029	36
	ML 5559	21 November 2034	333
	S-97986	19 June 2031	250
	S-100748	13 November 2029	930
	S-101660	13 November 2029	280
	S-101661	13 November 2029	425
	S-102737	08 May 2029	360
	S-102738	08 May 2029	130
	S-102739	08 May 2029	380
	S-106678	23 September 2029	1,880
	S-106771	24 June 2029	196
	S-106772	24 June 2029	193
	S-106773	27 December 2029	328
	S-110855	07 May 2031	1,321
	S-110856	04 December 2029	693
	S-111431	22 November 2030	774
	S-111432	22 November 2029	847
	S-113347	26 October 2029	1,309
	S-113350	05 December 2029	197
	S-113993	05 December 2029	29
	S-113994	05 December 2029	341
Seabee Area Subtotal			13,657
Seabee, Carina	S-99942	31 October 2029	65
Seabee Fisher	MC00000999	16 November 2027	2,757
	MC00001042	07 January 2029	513
	MC00001165	16 February 2029	675
	MC00002559	17 December 2027	329
	MC00002560	17 December 2027	429
	MC00002561	17 December 2027	641
	MC00002598	18 December 2027	643
	MC00002602	18 December 2028	702
	MC00002603	18 December 2028	711
	MC00002746	14 January 2029	280
	MC00002750	14 January 2028	232
	MC00002758	15 January 2029	517
	MC00002759	15 January 2028	544
	MC00002760	15 January 2028	675
	MC00002761	15 January 2029	496
	MC00002762	15 January 2029	559
	MC00002763	15 January 2028	528
	MC00002794	22 January 2028	495

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	MC00002795	22 January 2029	197
	MC00002796	22 January 2029	69
	MC00002868	02 February 2028	498
	MC00002869	02 February 2028	507
	MC00003512	29 July 2027	495
	MC00003514	29 July 2029	524
	MC00003515	29 July 2029	626
	MC00003541	05 August 2027	495
	MC00003542	05 August 2027	492
	MC00003543	05 August 2027	461
	MC00003544	05 August 2028	480
	MC00003545	05 August 2027	564
	MC00003546	05 August 2027	439
	MC00003547	05 August 2027	616
	MC00003548	05 August 2027	563
	MC00003549	05 August 2027	460
	MC00003550	05 August 2027	654
	MC00003553	06 August 2027	575
	MC00003568	09 August 2029	526
Seabee Fisher, cont.d	MC00003584	11 August 2027	591
	MC00003585	11 August 2027	414
	MC00003605	17 August 2027	710
	MC00003628	23 August 2027	1,031
	MC00003630	23 August 2027	731
	MC00003668	31 August 2029	265
	MC00004671	19 March 2029	2,739
	MC00012708	13 May 2028	17
	S-111184	16 February 2029	526
	S-111185	16 February 2029	150
	S-111186	16 February 2029	529
	S-111400	06 October 2029	300
	S-111401	06 October 2029	791
	S-111402	06 October 2029	434
	S-111403	06 October 2029	143
	S-111404	06 October 2029	155
Seabee Fisher Subtotal			30,493
Seabee Fisher S	MC00007135	13 November 2028	251
	MC00007136	13 November 2028	214
	MC00007290	20 November 2028	1,167
	MC00007291	20 November 2028	934
	MC00007293	20 November 2028	705
	MC00007294	20 November 2028	598
	MC00007295	20 November 2028	296

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Seabee Fisher S Subtotal			4,165
Seabee New	MC00000028	14 March 2030	262
	MC00000030	14 March 2030	392
	MC00000069	19 March 2028	905
	MC00000070	19 March 2029	1,226
	MC00003517	30 July 2029	113
	MC00003518	30 July 2029	216
	MC00003532	04 August 2029	163
	MC00003551	05 August 2028	494
	MC00003552	06 August 2029	1,382
	MC00003564	09 August 2027	260
	MC00003571	10 August 2029	526
	MC00003593	13 August 2029	574
Seabee New, cont.d	MC00003631	23 August 2029	783
	MC00003716	10 August 2029	244
	MC00003717	10 August 2029	330
	MC00012589	27 February 2030	497
	MC00012591	27 February 2030	393
	MC00012592	27 February 2030	682
Seabee New Subtotal			9,442
Seabee Shane	S-105301	07 November 2033	642
Seabee Truscott	MC00000093	19 March 2030	3,695
GRAND TOTAL			62,158

iP Mineral leases from which the SGO is currently producing

Note: Work filings have been submitted to the Saskatchewan Ministry of the Economy and are pending review

Claude Resources initially staked or acquired the SGO mineral leases and mineral claims prior to SSR’s acquisition of the property on 31 May 2016.

In January 1999, after Claude Resources fulfilled the conditions of an option agreement and obtained a 100% interest in the adjoining Currie Rose property, a portion of a previous claim CBS 7057 was converted to a mineral lease (ML 5520). The original 10 quartz mineral claims covering the Seabee mine site were consolidated into a single mineral lease (ML 5519) granted by the Provincial Crown in November 1999. In July 2021, a formal request from SGO operations to consolidate ML 5519 and ML 5520 into a single mineral lease ML 5559, a non-producing lease expiring in 2034, was granted.

Additional mineral leases were added at the Santoy 7 deposit (ML 5535) and Porky West deposit (ML 5536) in 2007, at the Santoy 8 deposit (ML 5543) in 2009, and at the Santoy Gap deposit (ML 5551) in 2013. The SGO is currently producing from mineral leases ML 5558, ML 5543, and ML 5551.

Annual rental and mining land taxes, and the fulfillment of work commitments, are required by SSR to ensure that the mineral leases and mineral claims remain in good standing.

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Figure 3.3 SGO Mining Land Tenure Map

SGO, 2021

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3.4 Underlying Agreements

The SGO is subject to production and net smelter return (NSR) royalties payable to third parties.

Claude Resources entered into a royalty agreement with Orion Mine Financial Fund (Orion) in 2014 to grant a 3% NSR royalty on gold sales from the SGO. Payments are to be paid quarterly in cash or in physical gold at the average price of gold in each calendar month. This royalty has subsequently been transferred by Orion to Osisko Gold Royalties Ltd.

In the first quarter of 2016, Claude Resources also granted an aggregate 1% NSR royalty on gold production from certain mineral dispositions to an individual and a private company. These dispositions include MC00003518, MC00003532, MC00003571, MC00003573, MC00003594, MC00003631, MC00003716, and MC00003717 from which the SGO is not currently producing. SSR has an option to re-purchase one half of this NSR royalty for C$1.0M.

The SGO is also subject to certain royalty payments to the Province of Saskatchewan that are calculated on 10% of net operating profits and are payable once capital and exploration costs are recovered. No royalty payments have been made to the Province of Saskatchewan to date.

To the extent known, no other significant factors or risks affect access, title, or the right or ability to perform work at the SGO.

3.5 Environmental Considerations

Further discussion is provided in Sections 15 and 17 of this Seabee21TRS. The primary environmental considerations and potential liabilities with the SGO are related to the operation’s solid waste (mill tailings) and the treatment and release of mine and mill effluent.

The tailings produced at the mill are currently managed in permanent management facilities (the East Lake tailings management facility (TMF) and the Triangle Lake TMF). The operation of these two facilities is conducted in accordance with the SGO’s Tailings Operation, Maintenance, and Surveillance Manual (SRK, 2020) and the Canadian Dam Safety Guidelines. In addition, the current approved SGO Preliminary Decommissioning and Reclamation Plan, 2016 Update (SRK, 2017b) addresses all potential long-term environmental and physical stability issues of the containment structures in accordance with the Canadian Dam Association Guidelines. The SGO cost estimate for closure activities were updated in 2020 and approved by the Ministry of Environment in July 2020 (Ministry of Environment, 2020).

With respect to water management and treatment, three discharge points exist at the operation. Mine water from the old Seabee mine (also referred to as the 2B mine, not currently in operation) is pumped to surface settling ponds that discharge to Laonil Lake. Mine water collected in the Santoy mine is pumped to surface and discharged to the Santoy settling ponds, which is treated in a Moving Biological Bed Reactor (MBBR) water treatment plant in order to remove ammonia and nutrients from the water prior to discharge to Lizard Lake.

In addition, mill effluent accumulating in the two TMFs that is not recycled to the mill as make up process water is treated in a chemical treatment plant through the addition of lime, hydrogen peroxide and ferric sulfate. The treated water from this plant currently discharges to the East Pond which flows through a series of wetlands and ultimately reports to the northern arm of Laonil Lake. A new chemical treatment plant combined with a MBBR was recently constructed to replace the existing chemical treatment plant. Both water treatment plants operate in compliance with the SGO’s Approval to Operate. All water discharges to the environment are in compliance with applicable provincial and federal regulations.

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3.6 Permits and Authorisation

Following a successful environmental assessment for a proposed gold mine development in the Province of Saskatchewan, applicants must secure a Surface Lease Agreement and subsequently an Approval to Operate a Pollutant Control Facilities (Approval to Operate) both issued from the Province of Saskatchewan’s Ministry of Environment.

The SGO currently has a valid surface lease with the Province of Saskatchewan, which was amended in March 2010. This surface lease provides SSR the Crown Land surface rights necessary to carry out the mining, milling, and associated operations at the SGO. The existing surface lease is in effect from March 2010 to its expiry date of 31 May 2040 (SMOE, 2010).

The SGO also holds an Approval to Operate No. PO19-193. This approval is issued by the Province of Saskatchewan’s Ministry of Environment pursuant to The Environmental Management and Protection Act, 2010 and its regulations. This approval was issued in October 2019 and is valid until September 2022. Renewal of this approval is triggered through an application submitted to the Ministry of Environment at least 90 days prior to its expiry date. Subject to the terms and conditions of this approval, SSR is authorised to operate all pollutant control facilities associated with the operation’s mine and mill (SMOE, 2016).

The SGO is also obligated to operate in compliance with the Canadian Metal and Diamond Mining Effluent Regulations issued pursuant to the Canadian Fisheries Act.

The SGO is currently in compliance with all environmental approvals and authorisations.

3.7 Other Significant Factors and Risks

SSR have advised that there are no other known significant risks that may affect access, title or the right or ability to perform mining related work on the Property.

Legal matters such as statutory and regulatory interpretations affecting the mine plan and environmental matters are outside the expertise of the QPs (see Section 25).

The Seabee21TRS QPs considers it reasonable to rely on SSR because SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

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4 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

4.1 Accessibility

Access to the SGO is by fixed-wing aircraft from the town of La Ronge, Saskatchewan to a 1,275 m airstrip located on the property (Figure 4.1). During the winter months, a 60 km winter road is built between the mine site and Brabant Lake on Highway 102, approximately 120 km north of La Ronge, to transport heavy supplies and equipment by truck.

4.2 Physiography, Climate, and Vegetation

The SGO is located within the Precambrian Canadian Shield environment. The site is relatively flat, with much of the area comprised of irregular, hummocky, rocky exposures. Low areas between hummocks that may have 5–9 m of relief are commonly filled with pockets of glacial till, and occasionally with muskeg. Overburden soils are thin in this area, and often the rock outcrops are exposed (Golder 2009).

The province of Saskatchewan is generally considered to have a continental climate, with temperatures and precipitation that vary significantly between seasons; winter is typically cold and dry, while summer is warm and wet. The majority of the province’s precipitation comes from summer rainfall, however, cool winters with long-surviving snowpack also contribute to greater precipitation.

The climate at the SGO is similar to that of the nearby Environment Canada weather station at Island Falls. The mean monthly temperatures recorded at this station between 1981 and 2010 range from –22.2°C in January to 17.3°C in July. Daily maximum temperatures have ranged on average from –15.9°C in January to 22.9°C in July, while daily minimum temperatures have ranged on average from –28.4°C in January to 11.6°C in July.

In the spring and summer months, historical total rainfall ranges on average from 6.8 mm in April to 84.6 mm in July, with mean annual rainfall totalling 347.9 mm. The winter months can experience significant snowfall, with historical monthly averages of 17.9 cm in February and March and up to 26.9 cm in November, with mean annual snowfall totalling 138.5 cm. A mixture of rain and snowfall is commonly experienced during the spring and fall.

Water inflow is well understood at the SGO based on actual data and is not expected to change during the life of mine. The current dewatering infrastructure system adequately manages water inflows and the system is expected to be expanded as the footprint of the Santoy mine expands.

The site is vegetated with a mixture of deciduous and coniferous trees and shrubs typical of a boreal forest, as shown in Figure 4.1. The area has been glacially scoured and is comprised of rocky, ice moulded ridges separated by lakes or muskeg filled depressions. Local relief in the surrounding area can be high, with the shoreline rising sharply to an elevation of 15–20 m above the lake surface (Golder 2009).

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Figure 4.1 Infrastructure at SGO and Typical Landscape of Project Area

A: SGO on-site airstrip apron

B: Seabee mine site

C: Seabee mine camp

D: Core shack

E: Typical landscape with view of Laonil Lake

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4.3 Local Resources and Infrastructure

SSR employs a workforce of approximately 350 employees who work on rotating schedules at the SGO. Up to 251 employees can be accommodated at the mine camps, which are equipped with kitchen and dining facilities, and a recreation room.

Electrical power to the property is provided by the provincial power authority, the Saskatchewan Power Corporation, via a 138 kV hydroelectric transmission line from Island Falls.

Potable water is obtained locally through SSR’s on-site potable water system.

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5 HISTORY

The Laonil Lake region has been intermittently explored since the 1940s, with the first gold discovery made in 1947 by prospectors working on behalf of Cominco Inc. (Cominco). Cominco conducted an extensive prospecting, geological mapping, trenching and diamond drilling programme between 1947 and 1950, and in 1958 was granted 10 quartz mining leases covering the property on which the Seabee Gold Operation (SGO) is located. From 1974 through 1983, Cominco conducted detailed drilling and exploration, and in 1983 sold the property to BEC International Corporation (BEC). BEC subsequently sold the property to Claude Resources in 1985.

In June 1985, Claude Resources optioned the property to Placer Development Limited (subsequently Placer Dome Inc. (Placer)). Placer conducted an extensive exploration programme, which involved geological mapping, trenching and stripping, geophysical, geochemical, environmental, and metallurgical studies, as well as surface and underground drilling. Upon completion of the programme, Placer allowed its option to expire and returned the property to Claude Resources in June 1988.

Claude Resources performed a geological review and analytical study to validate the work completed by Placer, and Cominco Engineering Services Limited (Cominco Engineering) subsequently completed bulk sampling and drilling as part of a feasibility study for the Seabee deposit. A Mineral Reserve estimate was completed in December 1988 and a positive feasibility study was completed in August 1989, which was further revised in May 1990. In the summer of 1990, Claude Resources placed the Seabee deposit into production and construction of the Seabee mine was initiated. Mill construction was completed in late 1991, and mining commenced in December 1991.

In 1998, prospecting and mapping was conducted by Claude Resources on the SGO site and a number of new discoveries were made, including the Porky West zone in 2002, the Santoy 7 deposit in 2004, the Santoy 8 and Santoy 8 East deposits in 2005, and the Santoy Gap deposit in 2010. Permit applications were submitted in 2005 to build an all-weather access road and conduct bulk sampling, and permission was subsequently granted to bulk sample the Santoy 7 and Porky West zones.

Claude Resources’ 2012 and 2013 exploration programmes focused on the Santoy deposit and establishing its geological and structural relationship to the Santoy 8 deposit. In February 2013, a shaft extension project was completed at the Seabee mine to reduce trucking distance and ore handling. In 2014, the ventilation raise at the Santoy deposit was completed and production was initiated. During 2015, an underground drill chamber was completed to begin drill testing the plunge continuity of the Santoy 8 deposit.

On 31 May 2016, SSR acquired Claude Resources and the SGO.

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5.1 Previous NI 43-101 Technical Reports

Mineral Resource and Mineral Reserve estimates have been prepared at various stages for the SGO. The two most recent are described below.

•The 2013 NI 43-101 Technical Report (Claude 2013) for the SGO, filed prior to SSR’s ownership, reported (as at 31 December 2012):

•Measured plus Indicated Mineral Resources of 469.6 kt at a grade of 5.10 g/t Au for 77 koz of contained gold, and

•Inferred Mineral Resources of 2,957.6 kt at a grade of 6.35 g/t Au for 603.4 koz of contained gold.

•Proven and Probable Mineral Reserves of 2,785.2 kt at a grade of 6.19 g/t Au for 554.1 koz of contained gold.

•The 2017 NI 43-101 Technical Report (SGOTR17) for the SGO:

'NI 43-101 Technical Report for the Seabee Gold Operation, Saskatchewan, Canada; Michael Selby, P. Eng; Dominic Chartier, P. Geo; Mark Liskowich, P. Geo; Jeffrey Kulas, P. Geo; reviewed by: Gary Poxleitner, P. Eng and Glen Cole, P. Geo, with Trevor Podaima, P. Eng., G. Ross MacFarlane, P. Eng,, and Caitlyn Adams, GIT, dated 20 October 2017,

filed prior to SSR’s ownership, reported (as at 31 December 2016):

•Measured plus Indicated Mineral Resources of 2,074 kt at a grade of 8.02 g/t Au for 535 koz of contained gold, and

•Inferred Mineral Resources of 2,495 kt at a grade of 7.66 g/t Au for 615 koz of contained gold.

•Proven and Probable Mineral Reserves of 1,371 kt at a grade of 8.19 g/t Au for 361 koz of contained gold.

These earlier reports are superseded by the Mineral Resource and Mineral Reserve estimates documented in this Seabee21TRS.

The SGO has produced over 1.6 Moz of gold since production began in 1991. Production has steadily increased to achieve a peak output of 84 koz, 96 koz, and 112 koz of gold during 2017, 2018, and 2019, respectively. A drop in gold production was experienced in 2020 due to impacts from the COVID-19 pandemic, resulting in less tonnes being processed. A summary of the production history of the SGO since 1996 is presented in Table 5.1.

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Table 5.1 Historical Production from the SGO (1996–2021)

Year		Milled Ore					Recovery<br><br><br>(%)	Gold Produced<br><br>(oz)	Cash Cost<br><br><br>($/oz)	Kitco <br>Gold<br>Price<br>($/oz)
	ktpa		tpd		Grade (Au g/t)
1996		194		531		6.45		36,709	345	388
1997		211		579		9.36	92.2	58,467	215	331
1998		225		615		9.27	92.6	60,200	168	294
1999		245		672		7.30	92.3	54,100	193	279
2000		238		651		8.58	87.9	58,300	190	279
2001		275		753		6.13	88.8	46,300	221	271
2002		202		553		6.59	93.7	41,500	246	310
2003		209		572		7.95	94.7	50,800	253	363
2004		187		512		7.15	95.2	41,200	297	410
2005		236		648		6.32	92.9	42,200	358	445
2006		246		674		6.16	93.6	46,300	396	603
2007		228		624		6.35	95.4	44,323	586	695
2008		228		626		6.46	95.8	45,466	683	872
2009		248		678		6.17	95.3	46,827	613	972
2010		204		559		7.55	95.5	47,270	692	1,225
2011		257		705		5.68	95.3	44,750	918	1,572
2012		275		754		5.86	95.6	44,756	998	1,669
2013		280		767		5.11	95.3	43,850	954	1,411
2014		280		766		7.32	95.7	62,984	757	1,266
2015		277		760		8.82	96.3	75,748	525	1,165
2016		313		857		7.91	96.6	80,351	639	1,250
2017		330		967		8.25	97.4	83,998	602	1,259
2018		352		1,125		9.16	97.4	95,602	505	1,267
2019		344		1,087		9.56	98.2	112,137	464	1,398
2020		255		1,163		10.10	98.4	81,686	534	1,790
2021		382		1,180		9.92	98.4	118,888	514	1,799

Period from and after acquisition of Claude Resources by SSR on 31 May 2016 by SSR

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6 GEOLOGICAL SETTING, MINERALISATION, AND DEPOSIT

6.1 Regional Geology

Northern Saskatchewan forms part of the Churchill Province of the Canadian Shield and has been subdivided into a series of litho-structural crustal units, of which the Seabee Gold Operation (SGO) is located within the Glennie domain of the Proterozoic Trans-Hudson Orogen (Figure 6.1 and Figure 6.2; Corrigan et al. 2007). The Trans-Hudson Orogen marks the collisional suture zone between the Rae-Hearne, Sask and Superior cratons formed during the closure of the Manikewan ocean (Stauffer 1984) and is divided into two distinct zones: namely, the Cree Lake Zone and the Reindeer Zone. The Cree Lake zone is composed of early Proterozoic continental shelf sedimentary rocks that overlie Archean rocks of the Hearne Province to the west. The Reindeer zone is comprised of mid-oceanic ridge basalts, oceanic island-arc basalts, inter-arc volcanogenic sedimentary rocks, and molasse-type sedimentary rocks. Plutonic rocks of various composition and age intrude the supracrustal sequence.

Figure 6.1 Cree Lake Zone and Reindeer Zone of the Trans-Hudson Orogen

Corrigan et al, 2007

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Figure 6.2 Regional Geology of the South-Western Trans-Hudson Orogen

Corrigan et al, 2007

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The Reindeer zone is further subdivided into litho-tectonic domains based on similarities of lithology, metamorphic grade, and structure (Lewry and Sibbald 1977), of which the Glennie domain is one such component. The Glennie domain is wedge shaped and is characterised by arcuate belts of Lower Proterozoic supracrustal rocks separated by granitoid gneisses and granitoid intrusions (Macdonald, 1987). It is bounded on the west by the north–north-east trending Stanley shear zone and bounded on the east by the north–south trending Tabbernor fault zone. To the south, Phanerozoic sedimentary rocks cover the Glennie domain.

Lewry et al. (1990) interpreted the Reindeer zone as a folded stack of nappes and thrust complexes divided by ductile mylonitic zones, emplaced during the terminal collision of the Trans-Hudson Orogen. The interpretation was based on Archean rocks that were found within the Glennie domain and neighbouring Hanson Lake block (Bell and Macdonald, 1982; Chiarenzelli et al., 1987; Craig, 1989) and imply that the Reindeer zone is underlain in part by Archean rocks (Lewry et al. 1990; Bickford et al. 1990). Extensive seismic geophysical studies (White et al., 1994) and samarium-neodymium systematics (Chauvel et al., 1987) support the interpretation.

The SGO is contained within one of the nappe sheets, referred to as the Wapassini Allochthon, and is interpreted as an upper tectonic assemblage separated from a lower sequence (the Iskwatikan Subdomain) by a high strain zone known as the Guncoat Gneisses (Macdonald, 1987). The allochthon was refolded and intruded by later plutons.

6.2 District Geology

The SGO is located within the northern portion of the Pine Lake greenstone belt. The belt has a strike length in excess of 50 km and comprises a variety of geochemically distinct tholeiitic mafic volcanic rocks formed in juvenile island arc settings, along with contemporaneous mafic intrusive rocks, volcaniclastics, sediments and felsic intrusions of varying age, as shown in Figure 6.3. Metamorphic grade across the Pine Lake greenstone belt ranges from upper greenschist to upper amphibolite, with the SGO hosted in the latter. The belt has been complexly folded by at least four major phases of deformation that are observed across the SGO site and elsewhere in the Glennie domain.

The SGO can be subdivided into three main geological domains:

•Seabee mine: The Seabee mine area is hosted within a coarsely layered mafic intrusion dominated by gabbro in the mine sequence.

•Santoy: The Santoy mine area is hosted within a sequence of mafic volcano-sedimentary rocks variably intruded by granodioritic rocks and separated by generally north–south trending thrust faults.

•Porky: The Porky deposit area is a mineralised trend hosted along a 12 km long openly folded unconformity, separating arenaceous sedimentary rocks of the Rae Lake synform to the north from mafic volcanic rocks of the Seabee mine area to the south.

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Figure 6.3 Local Geology Setting

SK GeoAtlas, 2017

6.3 Structural Setting

Coeval folding and thrusting during a protracted period of progressive deformation associated with the collision and amalgamation of several Archean continental fragments resulted in four major phases of deformation on the SGO property and are characterised as follows (SRK 2009):

•D1 (approximately 1,870 million years ago (Ma) to 1,845 Ma): Development of gneissic foliation and intrafolial folds associated with amalgamation of the Glennie and Flin Flon domains.

•D2 (approximately 1,845 Ma to 1,830 Ma): South directed thrusting and roughly east–west folding associated with the collision of the Reindeer zone and Sask craton.

•D3 (approximately 1,830 Ma to 1,800 Ma): West directed thrusting associated with north– north-west trending folding and transposition, and strike-slip reactivation of D2 shear zones controlled by the collision of the Superior and Sask cratons. Peak amphibolite grade metamorphism was reached at approximately 1,810 Ma.

•D4 (approximately 1,830 Ma): Refolding of D3 folds into regional type 1 and type 2 interference patterns associated with the final formation of the Trans-Hudson Orogen.

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SRK (2009) generated an integrated interpretation using published literature, regional mapping data, drilling data, and geophysical data that was collected during Goldak Airborne Surveys’ (Goldak) 2007 (Goldak 2007) aeromagnetic survey over the SGO (see Section 7.1.2.1). The following observations were made (Figure 6.4):

•Minor D1 faults trend north–south in the south-west corner of the interpretation area; Gneissic foliation and intrafolial folds cannot be observed on the scale of interpretation. D1 faults are present where a narrow strip of Pine Lake greenstone is interpreted to make the boundary between the Laonil Lake intrusive complex to the east and granodiorite units to the west. Any larger scale D1 features have been overprinted by subsequent deformation events.

•Regional north–south compression during D2 focussed on main deformation corridors and lithological contacts in the Laonil Lake intrusive complex. The Porky Lake metasedimentary belt was emplaced as late-stage southward thrust sheet(s) on the Pine Lake greenstone belt:

•Early-D2 gold mineralisation in the Seabee deposit is hosted in isoclinally folded quartz veins within D2 reverse shear zones that were reactivated as dextral shear zones during D3. Mapped veins appear offset by late-D2 structures that are subparallel to the Eyahpaise Lake pluton intrusive margin (1,859 Ma), suggesting that gold emplacement commenced prior to 1,859 Ma.

•Late-D2 gold mineralisation in the Porky deposits are associated with the development of a south verging thrust fault which formed late in the D2 phase when the Porky Lake metasedimentary belt was emplaced on the Pine Lake greenstone belt. The hosting fault was subsequently folded along a north–south axis, the Ray Lake synform, during D3 deformation.

•East–west compression during D3 reactivated deformation corridors and D2 structures in the Laonil Lake intrusive complex. Dextral kinematics were observed on west–south-west components, and sinistral kinematics were observed on all other components. Sinistral strike-slip shear zones observed in the central domain of the interpretation area, and north to north-west trending oblique-slip shear zones and folds in the eastern and western domains of the interpretation area. D3 folding affects D2 thrust faults (i.e., Ray Lake synform):

•Gold mineralisation in the Santoy deposits are associated with north–north- west trending D3 reverses and sinistral-reverse shear zones. It is possible the deposits are controlled by fault intersections, enhancing permeability.

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Figure 6.4 Integrated Structural Analysis of the SGO by SRK (2009) Based on Goldak’s 2007 Aeromagnetic Survey

SSR, 2009

6.4 Mineralisation

Gold mineralisation at the (now ceased) Seabee mine is hosted within an extensive network of sub-parallel shear structures, which crosscut the Laonil Lake intrusive complex. Vein mineralogy is dominantly quartz with pyrite, pyrrhotite, and chalcopyrite, and accessory tourmaline and carbonate. Gold occurs primarily as free, finely disseminated flakes and films replacing pyrite or at sulfide boundaries. Higher grade gold values are most often associated within sulfide-rich zones or at vein junctions (Figure 6.5). Silicification is the most common alteration type observed at the Seabee mine.

Gold mineralisation at the Santoy mine is hosted within calc-silicate altered shear structures with diopside-albite ±titanite-bearing quartz veins and occurs in gold-sulfide-chlorite-quartz veins in the shear zones, near or in the granodiorite and granite sills. Diopside-albite calc-silicate alteration facies are the main host to gold mineralisation in the Santoy 8A and Santoy 9A, 9B, and 9C zones. The Gap Hangingwall (GHW) deposit is hosted within a shallowly dipping, north plunging, folded limb of the Lizard Lake Pluton. Mineralisation is concentrated near the fold hinge within cm to m scale quartz veining which strikes roughly north south and dip sub-vertically.

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Figure 6.5 Typical Mineralisation Observed at the SGO

SGO, 2017

In the Porky deposit, the brittle-ductile lode gold system is hosted along a thick corridor of calc-silicate altered mafic volcanic and arenaceous sedimentary rocks that straddle a major unconformity along the southern margin of the Rae Lake synform. Both the Porky Main and Porky West deposits are characterised by the same calc-silicate alteration package, however, the unconformity and arenites host most of the auriferous quartz veins at the Porky West deposit.

Table 6.1 summarises the key stratigraphic and structural elements controlling the mineralisation at each of the SGO deposits.

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Table 6.1 Key Stratigraphic and Structural Elements Controlling Mineralisation at the Seabee, Santoy, and Porky Deposits (SSR, 2017b)

Area		Zone Name		Main Control of Mineralisation		Host Rock		Strike Length<br>(m)		Vertical Extent<br>(m)		Thickness<br><br>(m)		Strike
Seabee		L62		Quartz-tourmaline veins in shear zones		Laonil Lake Intrusive Complex gabbro		150		700		1–11		E
	2 Vein		Quartz-tourmaline veins in shear zones		Laonil Lake Intrusive Complex gabbro		1,800		1,400		2–7		ENE
	5-1 Shear		Quartz-tourmaline veins in shear zones		Laonil Lake Intrusive Complex gabbro		800		1,100		1–11		ENE
Santoy		Zone 7		Quartz veins in diopside-albite (calc- silicate) altered shear zones		Mafic metavolcanic rocks and lesser dioritic to granodioritic sills		330		120		2–10		N
	Zone 8		Quartz veins in diopside-albite (calc-silicate) altered shear zones		Mafic metavolcanic rocks and lesser dioritic to granodioritic sills		600		500		2.5–7		NW
	Zone 8 East		Quartz veins and flooding in sheared and isoclinally folded granodiorite		Granodiorite stock in fold nose near hanging wall contact with mafic metavolcanic rocks		200		250		1.5–15		NNW
	Zone 9		Quartz veins in diopside-albite (calc-silicate) altered shear zones		Mafic metavolcanic rocks and lesser dioritic to granodioritic sills		650		650		2–30		NW
		Gap Hanging-wall		Quartz veins in folded granodiorite intrusion		Lizard Lake Pluton		200		800		1–20		EW
Porky		Porky Main		Quartz veins in diopside-chlorite- actinolite (calc-silicate) altered shear zones		Mafic metavolcanic rocks and to a lesser extent arenaceous sedimentary rocks		280		180		1–4		SSE
	Porky West		Quartz veins in silicified calc-silicate altered shear zones		Arenaceous sedimentary rocks and to a lesser extent mafic metavolcanic rocks		400		250		1.5–12		E

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6.5 Deposit Types

The Seabee mine, Santoy mine, and Porky deposits host mesothermal, quartz-vein hosted lode gold deposits developed in major brittle-ductile to ductile shear systems. The gold mineralisation throughout the SGO exhibits complex geometrical patterns attributed to a combination of structural and/or lithological controls.

Mesothermal gold deposits typically emplaced as a system of en echelon veins, forming tabular veins in competent host rock lithologies, or as stockwork veinlets and stringers in less competent host rock lithologies. Lower grade bulk-tonnage style mineralisation with gold associated with disseminated sulfides may develop in areas peripheral to quartz veins. Mesothermal gold deposits can also be related to broad areas of fracturing, where gold and sulfides are associated with quartz veinlet networks. The quartz veins are typically in sharp contact with the wall-rock and can display a variety of textures including massive, ribboned or banded, and stockworks with anastomosing gashes and dilations, which may subsequently be altered or destroyed during deformation. Gold-quartz veins are found within zones of intense and pervasive carbonate alteration along faults proximal to trans crustal breaks, and often occur at a high angle to the primary collisional fault zone. They are commonly associated with late syn-collisional, structurally controlled intermediate to felsic magmatism, with economic deposits generally hosted by large competent units, such as intrusions or blocks of obducted oceanic crust (Ash and Alldrick, 1996).

Delaney (1992) suggested that lithological heterogeneities between feldspar porphyry dikes and gabbros of the Laonil Lake Intrusive Complex are responsible for the localisation and propagation of the shear zone. At Seabee, the structures trend between 045° to 085°, and dip north near vertically. Three discrete subsets of structures have been recognised trending at 070°, 085°, and 045°, with the 070° structures containing the auriferous veins. At Santoy, the structures trend between 340° to 315°, and dip moderately to the east. Vein geometry within the shear zones are commonly a combination of ‘S’ and ‘Z’ oblique and extensional types, and second order or Riedel shears.

High gold grades occur at the intersection of the primary ‘S’ shears with subordinate shear structures and/or where potassic altered diorite dikes have intruded the Laonil Lake gabbro prior to strain occurrence. It is probable that secondary dikes introduced additional gold to the system, which was later remobilised under strain conditions.

Exploration at the SGO is guided by applying techniques consistent with the identification and discovery of other quartz-vein lode gold systems. Airborne magnetic data is used in surface exploration to identify structural corridors and asymmetrical features, folds and target areas that are known to host gold on the property. This geophysical data is used in conjunction with regional and detailed geological mapping to identify major zones of shearing and alteration, of which calc-silicate alteration has proven to be the most prospective variety on the SGO property.

Geochemical soil sampling is also used as a regional exploration technique to identify gold and trace element vectors associated with Seabee-style gold mineralisation and has successfully identified gold mineralisation at various locations across the property. Once targets have been delineated by the above exploration methods, diamond drilling at wide spacing is used to test the structural systems to allow for SSR’s minimum threshold deposit size to be identified based on observed local grade.

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7 EXPLORATION

7.1 Surficial Exploration

7.1.1 Geochemistry

Historically, several rock and soil sampling programmes have been executed on the Seabee Gold Operation (SGO) property (Figure 7.1 and Figure 7.2).

Placer collected over 1,200 surface rock samples and nearly 7,000 soil samples between 1985 and 1988. The majority of samples were collected from the western portion of the property in the vicinity of Laonil Lake and Pine Lake, and proximal to and north of Porky Lake. Sample spacing was approximately every 20–25 m on 100 m spaced lines.

Claude Resources collected nearly 2,000 surface rock samples and over 7,000 soil samples between 1990 and 2013. Soil samples were primarily collected from the western portion of the property, with additional samples collected in the south-central portion of the property and in the Santoy area. Sample spacing was planned every 20–25 m on 100 m spaced lines. In 1990, rock samples were largely collected around the Laonil Lake, Porky Lake and Pine Lake areas, after which time the focus of exploration shifted to the Santoy area and samples were collected from the south-eastern portion of the SGO property.

Figure 7.1 Historical Rock Samples Collected at the SGO

SGO, 2017

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Figure 7.2 Historical Rock Samples Collected at the SGO

SGO, 2017

Upon its acquisition of the SGO, SSR undertook a review of all exploration activities conducted on the property by previous operators. An exploration programme was subsequently undertaken, including detailed mapping of the Herb West and Santoy Lake areas, as well as the collection of accompanying soil samples to be submitted for gold assay. Limited anomalous occurrences were identified from grab and soil sample results, and no new showings or gold in soil trends were recognised. SSR plans to map additional regions to the north and east within the Herb Lake area as additional shear zones are targeted.

In the Santoy Lake area, mapping extended from Santoy Lake to the west end of the Santoy mine. Soil sampling conducted over the same area resulted in the collection of 501 samples taken every 25 m on lines spaced 200 m apart. No anomalous trends of significance were identified. However, SSR has planned further exploration in prospective areas east and west of the 2016 exploration programme area.

7.1.2 Geophysical Surveys

7.1.2.1 Fixed Wing Aeromagnetic Survey 2007

Goldak performed an aeromagnetic survey over the SGO property on behalf of Claude Resources from 25 February to 15 March 2007 (Goldak, 2007). North–south traverse lines were flown with 100 m spacing and a control line separation of 1,000 m, totalling 2,284 line kilometres of high-resolution magnetic data collected. Nominal terrain clearance was 80 m above ground level.

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In 2009, SRK reviewed the aeromagnetic survey to make an integrated interpretation with the addition of using published literature, regional mapping data, and drilling data (Figure 6.4). The following recommendations were made regarding regional targeting:

•Regional deformation corridors have high prospectivity for gold, as structural complexity in the region over time has enhanced permeability.

•Key locations for gold mineralisation can be identified by understanding the kinematics active during gold mineralisation in combination with the interpreted fault geometry:

•Dilational jogs along D2 and D3 shear zones: shallower dipping segments of D2 and D3 reverse shear zones (similar setting to Santoy 7), left steps along D3 sinistral shear zones, and right steps along D3 dextral shear zones.

•Fault intersections (i.e., deformation corridors).

•Additional parameters that enhance gold mineralisation in the Seabee area include:

•High competency contrast (i.e., variations in lithology).

•Presence of multiple intrusions exploiting similar structural pathways as potential hydrothermal fluids.

•Proximity to the Pine Lake conglomerates, a structurally bound conglomerate package similar to the Abitibi Timiskaming conglomerates.

7.1.2.2 Titan-24 DC / IP and MT Survey 2010

In early 2010, Quantec Geoscience Ltd. (Quantec) were commissioned to perform a Titan-24 direct current / induced polarisation and audio-magnetotelluric ground geophysical survey over the Santoy area on behalf of Claude Resources. The Titan-24 direct current and induced polarisation data were inverted to produce cross-sections of the resistivity and chargeability variations along four survey lines. In its standard configuration, the Titan-24 surveys typically image direct current resistivity and induced polarisation to 500–750 m in sub-vertical tabular geological settings, and up to 50% more for sub-horizontal geological settings. Audio-magnetotelluric inversion depth is generally limited to approximately half the length of the survey line or profile.

Quantec (2013) made the following observations and interpretations based on the 2010 survey results:

•Based on common features observed in the four lines, both the chargeability and resistivity showed weak to strong chargeability responses and low to high resistivity distribution.

•A major difference in the direct current / induced polarisation and audio-magnetotelluric signatures between the north-east, central and south-western regions of the survey lines was observed. The highest conductivity was observed from near surface to approximately 100 m depth in both direct current and audio-magnetotelluric resistivity models. The conductive cap was found above a thick, highly resistive body in the central part of the grid. The central part is relatively more resistive, which potentially depicted the mineralisation of interest having gold traces. Drill data provided by Claude Resources confirmed the presence of gold traces related to high resistivity in audio-magnetotelluric sections and at gradient zone of direct current resistivity sections where resistivity changed in nearly two orders of magnitude.

•It is possible that the direct current and audio-magnetotelluric inversions could be affected by 3D signatures of linear structures which may run parallel and/or sub-parallel to the survey lines. The observed high resistivity contrast in direct current and audio-magnetotelluric inversion models potentially defines the geological structures, lithological units and alteration zones which may be related to gold mineralisation.

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•Low chargeability responses were generally observed from near surface to approximately 100 m depth and associated with the conductive cap. The north-eastern part of the lines represents high chargeability from near surface to a greater depth than the rest of the grid and may be associated with a geological contact and/or fault zone.

•Below the low chargeability top layer, the central part of the grid shows moderate chargeability associated with high resistivity potentially consisting of the mineralisation of interest. Drilling data provided by Claude Resources confirmed the presence of gold traces related to moderate chargeability. The change in chargeability between the north-east and central areas may describe the alteration zone related to gold mineralisation.

•The geological setting of the region giving rise to a variety of geophysical responses for possible mineralisation, and the inversion results of the direct current / induced polarisation and audio-magnetotelluric models along with drilling data, confirmed that the gold deposit in this area is structurally controlled and dominated at gradient zones.

7.1.3 Airborne Magnetic and Radiometric Survey 2016

SSR contracted Precision GeoSurveys Inc. (Precision) to complete a high resolution airborne magnetic and radiometric survey over the most recently staked portion of the SGO land package from 30 August to 4 September 2016 (Precision 2016). The survey block covered an area of 22.9 km x 15.0 km and included 150 survey lines and 25 tie lines that totalled 1,815 line kilometres. Survey lines were spaced 100 m in an east–west orientation and tie lines were spaced 1,000 m in a north–south orientation. Nominal terrain clearance was specified at 75 m.

Selected suspect anomalies were re-flown for confirmation, specifically those found on a single flight line. Lines to be re-flown were a minimum of 2,000 m long, so that survey line re-flights crossed at least two tie lines and tie line re-flights crossed at least five survey lines.

Survey overview maps (flight lines and digital terrain model), magnetic maps (total magnetic intensity, residual magnetic intensity and calculated vertical gradient of the residual magnetic intensity), and radiometric maps were produced by Precision, with the objective of identifying potential new targets for gold mineralisation on the Seabee property.

The magnetic data was collected to better observe the structural nature of the underlying bedrock and, where possible, determine major breaks in the regional stratigraphy along which shear zones can propagate, and the radiometric data was used to determine the relative amounts of uranium, thorium and potassium in the surficial rocks and soils to be used for the mapping of bedrock lithology, alteration and structure. The resultant data were found to be consistent with the structure of the bedrock and major lithological breaks previously interpreted by geological mapping, air photo interpretation and drilling. The data was also consistent with the two-dimensional structural architecture and intensity of previously flown surveys within juxtaposed survey blocks.

7.2 Drilling

Prior to SSR’s acquisition of the SGO, and as at 31 December 2015, a total of 2,037 surface drillholes totalling approximately 389,281 m and 4,818 underground holes totalling approximately 861,514 m had been completed on the property.

For the year ended 31 December 2021, SSR has drilled an additional 287 surface holes totalling approximately 106,916 m and 1,321 underground holes totalling approximately 299,670 m since acquiring the property from Claude Resources Inc.

Table 7.1 summarises the drilling completed on the property. Figure 7.3 displays the surface holes completed on the property. Details regarding the salient drill programmes are discussed in greater detail in the subsections below.

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Table 7.1 Surface and Underground Drilling Completed on the SGO to 31 December 2021

Drilling Programme	Company	No. Surface Drillholes	Surface Metres Drilled	No. Underground Drillholes	Underground Metres Drilled	Total Number of Drillholes	Total Metres
1947–1988	Various (Cominco, Claude Resources, Placer)	278	35,419	77	6,491	355	41,910
1989–2012	Claude Resources	1,742	344,415	4,190	724,858	5,932	1,069,273
2013–2015	Claude Resources	17	9,447	551	130,165	568	139,612
2016	Claude Resources / SSR	51	19,817	306	65,021	357	84,838
2017	SSR	14	10,506	159	61,179	173	71,685
2018	SSR	83	24,389	229	52,500	312	76,889
2019	SSR	44	16,888	174	51,278	218	68,166
2020	SSR	21	9,638	177	30,040	198	39,678
2021	SSR	74	25,678	276	39,652	350	65,330
Total		2,324	496,197	6,139	1,161,184	8,463	1,657,381

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7.2.1 Drilling by Cominco, Claude Resources, and Placer 1947–1988

Cominco identified four gold-bearing zones on the SGO property from 1947 through 1950, after drilling 79 holes totalling 4,414 m, and in 1961 drilled two shallow holes of 41 m as part of an overall review of the known property data. In 1974, Cominco drill tested additional vein structures with 16 holes totalling 458 m, and commenced further exploration from 1982 through 1983 whereby 20 holes were drilled totalling 3,776 m. This drill programme was not completed before Cominco sold the property in 1983.

Upon acquisition of the property, Claude Resources drilled three holes totalling 226 m to corroborate Cominco’s work and property estimates. Pursuant to an option agreement with Claude Resources, Placer executed an extensive surface and underground drilling programme from June 1985 to June 1988, whereby a total of 95 surface holes and 72 underground holes were completed. Placer determined the property did not meet its criteria for development and returned the property to Claude Resources in 1988.

Figure 7.3 Map Showing the Distribution of Surface Drilling in Relation to the Seabee, Santoy and Porky Deposits, and other known Gold Occurrences on the Seabee Property

SGO, 2017

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7.2.2 Drilling by Claude Resources 1989–2015

7.2.2.1 Seabee Area

After obtaining a 100% interest in the Currie Rose property from Currie Rose Resources Inc. in 1994, which consisted of over 4,000 ha surrounding the Seabee mine, Claude Resources conducted a drilling programme to test gold-bearing structures identified the previous year during a prospecting programme. The drill programme consisted of 27 holes totalling 3,458 m. In 1996, drilling defined the 10 zone, identified the previous year and found adjacent to the western boundary of the Seabee mine. A total of 23 holes were drilled totalling 2,567 m. Diamond drilling in 1997 explored the vein extensions of the 10 vein and 2C vein structures with seven holes totalling 1,573 m. The 1999 drill programme focused on an area south-west of the Seabee mine trend and consisted of 7,726 m drilled in 47 holes.

As a follow-up, the majority of holes drilled in 2000 were collared to the west of mining lease ML 5520 in the Bird Lake area, to explore for mineralised structures parallel to the Seabee 2 vein. Targets in the Porky Lake and Pine Lake areas were also tested. Six additional remote targets, namely the Scoop, Porky, Herb, Pine, East, and West Bird Lakes were explored in 2001, with anomalous gold values encountered within variably sheared host rocks.

In 2002, drilling focused on a laterally extensive geochemical soil anomaly on the west shore of Porky Lake, and on a series of quartz-bearing shear structures north and east of the No. 5 ramp access. The drill programme successfully discovered the Porky West zone and produced elevated gold values over narrow widths at the No. 5 ramp access.

Drilling in 2003 in the Porky area discovered the Porky West zone, an arenite-hosted high-grade gold lens. Subsequent drilling in 2004 focused on delineation drilling at the Porky Main and Porky West zones, and exploration drilling on the eastern limb of the Porky Lake anticline targeted the contact between the mafic metavolcanics rocks and feldspathic arenite.

A small diamond drill programme was completed in 2009, which extended the down plunge extent of the Porky West ore shoots.

Evaluation of the Neptune target, located approximately 6 km north of the Seabee mine, was the focus of exploration in 2010, where drill testing included two holes. Exploration efforts in 2011 included a further 28 drillholes to test the 1.8 km strike length of the soil anomaly to vertical depths of up to 250 m, and in 2012, further drilling at the Neptune target confirmed the sporadic nature of the gold-bearing system.

7.2.2.2 Santoy Area

Prospecting and geological mapping in 1998 resulted in the discovery of numerous new veins in the Santoy area. The targets were drill tested in 2002 with encouraging results and became the focus of additional exploration programmes leading to the discovery of the Santoy 7, and Santoy 8 and Santoy 8 East deposits in 2004 and 2005. In 2004, five holes totalling 598 m were drilled at Santoy 6, 48 holes totalling 6,164 m were drilled at Santoy 7, and 21 holes totalling 2,797 m were drilled at Santoy 8. Drilling of the Santoy 8 and Santoy 8 East zones in 2005 was aimed at testing the north–north-west plunge and dip extensions of the mineralised shear structures outlined in previous drill programmes. Sixty-eight holes totalling 15,296 m were drilled, with an additional 20 holes totalling 6,272 m drilled in the summer of 2005. Infill drilling continued in 2007 to collect information for proposed mine plans with 25 m infill data to a depth of 250 m completed on the Santoy 8 and Santoy 8 East deposits. A total of 31,670 m was drilled from 147 holes.

Exploration drilling in 2010 targeted the Santoy Gap area to test the Santoy shear system between the Santoy 7 and Santoy 8 deposits, as well as to continue to investigate the down-plunge continuity of the Santoy 8 and Santoy 8 East deposits. Results from the programme outlined continuity at depth for both the Santoy 8 and Santoy 8 East deposit.

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Drilling defined the Santoy deposit in 2011. Multiple high-grade intervals were intercepted, expanding the strike length and width of the known mineralisation. During 2012, exploration focused on defining the relationship between the Santoy and Santoy 8 deposits to depths up to 750 m. Infill and exploration drilling around the Santoy lens and Santoy Shear zone continued to confirm and expand the Santoy system, and also identified a sub-parallel lens approximately 150 m of the east of the Santoy deposit.

In 2013, surface drilling programmes targeted the down plunge extension of the Santoy and Santoy 8 deposits, resulting in two out of three step-out holes returning high-grade gold intercepts. The Santoy system was extended down plunge to 650 m depth and the Santoy 8 deposit was extended 400 m below the base of the previously estimated Inferred Mineral Resource.

Underground drilling in 2014 focused on defining and expanding the Mineral Reserve and Mineral Resource at the Santoy deposit. Results identified high-grade and promising widths of gold mineralisation hosted within three vein systems, named the Santoy 9A, 9B, and 9C deposits. Additional underground drilling in 2015 focused on the expansion of Mineral Reserve and Mineral Resource at the Santoy deposit, and a 6,000 m drill programme targeted the plunge continuity of the Santoy 8 deposit. Results from the Santoy up-dip drilling demonstrated the potential for expansion of the deposit, and drilling results within, down-dip and down plunge also increased confidence in the continuity of the deposit at depth.

7.2.3 Drilling by Claude Resources and SSR 2016

Drilling in 2016 on the SGO property had the objective of increasing and converting the Mineral Resource to Mineral Reserve.

An underground diamond drilling programme to upgrade the Inferred Mineral Resource and explore the extension of the Santoy 8A and Santoy deposits was completed by SSR. From surface, drilling was conducted to upgrade the up-plunge extension of the Santoy 9A, 9B, and 9C deposits as well as to complete deeper infill drilling on the Santoy 8A Inferred Mineral Resource.

At the Seabee mine, five holes were drilled on the 15 Vein target, an offset mineralised structure along the 19 Shear. At the Carr target, located 4 km along strike to the north of the Santoy mine, SSR drilled nine holes over a 2 km strike length, totalling approximately 2,500 m. At the Herb West target, located 2.2 km west–north-west of the Seabee mine, four holes totalling approximately 1,130 m were completed. Results from drilling the above targets revealed shear-hosted quartz-veining structures with gold-bearing sulfide mineralisation and warranted follow-up drilling.

7.2.4 Drilling by SSR Mining 2017 Onwards

Drilling in 2017 from underground continued to focus primarily on the definition and expansion of the resources on the Santoy 8 and 9 veins. During 2017 the first underground programme designed to test the Gap Hangingwall (GHW) target was implemented. A limited surface programme focused on defining the margins of the Santoy 8 and 9 veins that could not be tested from underground. The Exploration team drilled four surface holes attempting to locate the depth continuity of the Santoy 6 showing without success.

Drilling in 2018 from underground continued to focus primarily on the definition and expansion of the resources on the Santoy 8 and 9 veins. One underground drill was dedicated to exploring the Santoy Hangingwall target. Surface drilling focused on the definition of near surface Santoy 9 veins. A surface based deep exploration programme was completed in an attempt to intersect the 926 zone.

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Drilling in 2019 from underground shifted significantly to focus on the definition and expansion of the GHW target. Limited underground drilling was conducted on the Santoy 8 and 9 veins largely due to a paucity of suitable drill bays. Surface drilling also focused on the definition and expansion of the GHW target; the result of which was a maiden resource for the GHW deposit at year end of 1.15 Mt at 7.5 g/t Au Indicated and 850 Mt at 7.9 g/t Au Inferred for 496 koz gold. The Exploration team conducted a limited programme along strike of the GHW target within the Lizard Lake Pluton following up on the previous year’s prospecting and soil sampling programme with limited success.

Underground drilling in 2020 focused almost exclusively on the definition and expansion of the GHW deposit with limited definition drilling conducted on the Santoy 8 and 9 veins due largely to a paucity of suitable drill bays and size of the GHW. Surface drilling shifted focus to the Santoy Hangingwall target with several smaller programmes also conducted to test near-mine targets that could not be reached from underground platforms. Of the surface targets tested in in 2020 only the Santoy Hangingwall showed promise for developing into a resource. The exploration team conducted limited follow up on their 2019 programme along strike of the GHW with mixed success.

7.3 SSR Drilling Procedures

7.3.1 Underground Drilling Procedures

The most important dataset informing the current Mineral Resource at the SGO is derived from underground drilling. Underground drill layouts are created using Geovia GEMS software three-dimensional software. Three dimensional lines are created between a desired pierce point and a collar location for each planned hole. The resulting azimuths from the developed hole traces are given to the survey department as a digital plan map, which is then uploaded into the Mine Markup tablet. All underground drill layouts are created in mine grid coordinates. The survey crew then goes underground to physically paint the drill lines of all holes on the excavation walls by means of numbered lines, with front sight and back sight marked accordingly. Spads are drilled into the lines with which the line number and azimuth are marked on flagging tape in the event the painted lines become obscured or illegible over time.

Underground drills are equipped with laser sighting systems for accurate alignment on the specified drill line. Dips are set using digital inclinometers magnetically attached on the drill’s feed frame. Completed drillholes are surveyed using a Reflex multi-shot tool and wireless palm unit to measure the azimuth, dip and total magnetic field. Drillholes are surveyed at 10 m intervals from the bottom of the hole to the collar. For holes exceeding 500 m, it is common practice to take single shots every 30–50 m as the hole advances to ensure that deviation is within acceptable ranges. Stored data is transferred to a memory stick from the palm unit and is then uploaded into a programme called S-Process where the data is visually verified, and then transferred into the GEMS MS Access database as a comma-delimited text file (*.csv). Upon completion of each hole, the collar locations and azimuths are recorded by mine surveyors and the data is transferred to the drill geologist as a .csv file for inclusion into the GEMS survey field in the MS Access database. Completed holes are checked against planned hole traces to verify that they are spatially correct in the three-dimensional model.

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Underground drill logging takes place at the drill chamber underground. Data from logging is captured on paper log sheets and include header data containing the drillhole identification number, date, the logging geologist’s name and planned hole directional data. The main body of the log contains row and column fields for depth intervals, lithological descriptions, sample numbers, assay results and rock quality data measurements. Upon completion of logging, the information is manually entered into the GEMS database by a mine geologist. Completed drill logs are placed in a file folder for future verification by the senior mine geologist before inclusion into resource updates. Completed assay data is housed in an excel database owned by the Seabee mine laboratory. The geology department has read only access to this file and can copy and paste results into the GEMS database. Underground chip and muck sample data is recorded in a sample tag book and later manually entered into the Chips MS Access database by the mine geologist, which is updated with assay results as they are made available in the laboratory excel database.

7.3.2 Surface Exploration Drilling Procedures

Upon establishing drill targets, three-dimensional points representing surface drillhole locations are created. Drillhole traces are planned to pierce the target as close to orthogonal as possible to obtain a true thickness of the stratigraphy. After the anticipated hole deviation is accounted for and an optimal trace is obtained, the surface location is inspected to ensure suitability.

In the field, hole collar locations and two front sights are recorded with a handheld GPS prior to data being entered into the MS Access-based Core Logger software. Alternatively, the drill contractor may align the drill using the DeviSight tool which uses GPS to derive an azimuth which is preferable to a compass and front sights due to the elimination of magnetic interference and operator error when aligning two pickets.

Reflex EZ_Shot multi-shot device tests record the hole’s azimuth and dip. Tests are completed at 100 m intervals during down-hole drilling and are collected at 30–100 m intervals upon completion of the hole as rods are being pulled if the desired density of measurements is insufficient with the shots taken while drilling. The data are collected via a handheld device that syncs to the Reflex tool down hole and are recorded onto Reflex paper sheets. The paper sheet and digital data are delivered to the supervising exploration geologist and are downloaded and input into a database to track the hole progression, ensuring that unexpected and/or excessive deviation has not occurred.

Once a hole has been completed an aluminium plug is placed approximately 10 m downhole from the base of the casing and the hole is cemented to the top. The SGO mine survey team then takes a DGPS waypoint of the collar location with the base station for final verification of its location, providing accuracy within 0.3 m of the hole location. Drillholes where this level of accuracy is not required may be surveyed in by handheld GPS unit or using the DeviSight’s GPS coordinates giving an accuracy on the order of +/–3 m. The digital data is sent to the supervising exploration geologist and the final three-dimensional coordinates of the hole are entered into MS Access database and tracking software.

Drill core is transported to the core logging facility, where it is marked and logged. Data from individual drill programmes is captured in an MS Access database, including drillhole collar and header information, detailed descriptions of lithological units, structures, alteration and mineralisation, core recovery and RQD data, and sample information. Photographs of core are taken both wet and dry, and digital copies are archived. Upon receiving laboratory results and confirming quality control results, the entire dataset is combined into a master MS Access database and incorporated into the tracking software. Core boxes are stacked and stored at the SGO core storage yard with metal tags affixed by staples indicating BHID, box number, and interval contained.

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7.3.3 Drill Sampling

7.3.3.1 Sampling by Previous Operators 1949–2009

Generally, historical sampling on the SGO was conducted by a geologist selecting mineralised intervals based on visual inspection of drill core. Selected intervals were split by hydraulic or manual power splitter and sent for analyses at the on-site laboratory or an offsite laboratory.

Information regarding historical sample preparation and analyses is incomplete or unavailable and is therefore not discussed in detail in this Seabee21TRS. Multiple sampling methods are attributed to individual drilling campaigns without differentiation of the method applied to each hole.

Furthermore, drilling prior to 2009 tends only to have dip surveys and no control on azimuth, and is therefore unreliable.

Current Mineral Resource and Mineral Reserve estimates at the SGO are informed almost entirely by drilling post-2009, excluding the Mineral Resources attributed to the Porky West deposit. The historical sample preparation and analyses therefore does not have a significant material impact on the property.

7.3.3.2 Drill Core Sampling by Claude Resources and SSR 2009 to Present

Drill core is logged in detail on site by SSR geologists. Rock quality and core recovery are documented, zones of potential mineralisation are marked for sampling, and three to five samples are marked in both the hangingwall and footwall.

Surface diamond drill core samples are chosen based on geology and average 1.0–1.5 m in width, with 0.3 m width samples taken for geological interpretation purposes. The sampling interval was established by minimum or maximum sampling lengths, and geological and/or structural criteria, and are no less than 0.10 m. Discrete intervals of mineralised or prospective lithologies which measure more than 0.10 m and less than 1.0 m may be sampled as a single sample. Mineralised or prospective lithologies which are greater than 1.0 m in width tend to be broken into one metre sample intervals internal to the interval of interest. Intervals immediately adjacent to mineralised or prospective lithologies are sampled, at a minimum, 1.0 m from the contact with the prospective mineralogy. Sampling of less prospective, or weakly altered lithologies, may be sampled at 1.5–2.0 m intervals at the discretion of the logging geologist.

Intervals deemed unprospective for gold mineralisation by the geologist are sampled using a composite sample, not exceeding 8 m in length. The composite sample consists of no less than one 10 cm piece of core selected per 1.5 m in the total 8 m sample interval. The composite sample is used to ensure that mineralised zones not immediately recognised by the geologist are not missed. If a composite sample grades more than 0.10 g/t Au, then the interval is re-logged and re-sampled at a 1 m sample interval to determine the source of the anomalous gold grade. Field geologists are trained to sample additional intervals that may have associated gold mineralisation, such as zones of increased sulfide mineral content or quartz veining not previously associated with a known mineralised zone. Sample intervals are recorded in a MS Access database, and photographs of each core box are taken. Certified reference material, blanks or duplicate samples are inserted into the sample stream at regular intervals of at least 1-in-20 samples.

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After the drill core is logged and marked for assay it is transferred to the core splitting facility, where the selected intervals are sawed lengthwise. The half core to be analysed is double-bagged, sealed, and labelled with coded security tags, while the other half remains in the core box as a record. In the case of duplicate samples or re-sampling, core is sawn in quarters and a quarter core is retained as a record. Some core intervals are destroyed in metallurgical testing and are marked by survey stakes with metal labels in the core boxes from which the interval is removed from. Samples to be sent for analyses are placed in white rice bags, weighed and closed with a uniquely coded security zip tie. Sample submittal forms are sent to the appropriate laboratory indicating the number of samples, weight and security tag numbers of each sample in the shipment. This data is verified by the laboratory when the shipment is received, and any broken tags or sample bags that appear to have been tampered with are reported.

Underground drill core is logged by geologists in the underground drill chamber. Sample intervals are selected by the logging geologist and measure no less than 0.10 m. Discrete intervals of mineralised or prospective lithologies which measure more than 0.10 m and less than 1.0 m may be taken as a single sample. Mineralised or prospective lithologies that are greater than 1.0 m in width are typically divided into 1 m sample intervals. Intervals immediately adjacent to mineralised or prospective lithologies are sampled, at a minimum, 1.0 m from the contact with the prospective mineralogy. Less-prospective, or weakly altered lithologies may be sampled at 1.5–2.0 m intervals at the discretion of the logging geologist. No samples are taken of core considered by the geologist to be unmineralised. Sample intervals are recorded on paper logs and later transcribed by hand into a GEMS project database. Certified reference material standards (CRM) are inserted at a rate of 1-in-20 samples.

Once the intervals to be sampled are selected, the whole core is placed in a sample bag with a uniquely numbered identification tag and delivered to the Seabee laboratory for analyses. Unsampled core is dumped near the drill chamber and used as fill in the mine.

Unauthorised personnel are not permitted access to the drill machines or the core logging and core splitting facilities.

7.3.3.3 Underground Chip and Muck Sampling

Chip samples are collected by a geologist at the working face; the hangingwall to footwall is sampled, with intervals divided based on lithological boundaries and not exceeding 1.5 m in width. Wall rock is also included in this sample type, as it is primarily used as a daily estimate of grade being delivered to the mill. Muck samples are obtained by the geologist when they are unable to reach the working face in a heading. These samples consist of grabs of muck on the floor of the drift, with no less than three muck samples taken at a face unless extenuating circumstances requires fewer samples. Chip samples retain their specific width weighting, while muck samples are assigned a proxy interval based on the number of samples collected and the width of the sill from which the samples are collected. The samples are bagged, tagged with a unique identifying number and transported to the Seabee laboratory for analyses following the methodology described in the previous section. Assay values are tracked in an MS Access database.

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7.4 Density

Density data was collected from NQ diameter drill core during the 2011 Santoy drilling programme by the SGO exploration department. Half core was weighed within mineralised zones, while whole core was weighed within waste domains. A total of 433 density measurements were collected from 45 different holes. The results were tabulated, sorted, and averaged by lithology. Initially weight percent estimates of the various ore zones were calculated based on drill core and underground observations. Assigned densities are reviewed annually by comparing to collected daily density determinations carried out on mill feed samples. Analyses were performed on site by water displacement using the following methodology:

•Place a dry glass vessel on a balance and zero the weight;

•Collect a 20–25 cm piece of half core or whole core from the interval of interest and place into the vessel;

•Record the weight of the core, and zero the balance;

•Fill the vessel to marked line with cold water; and

•Suspend core in water and weigh the vessel with the water and core.

The difference between the original water weight and the second reading is equal to the volume of water displaced by the core, from which the density was calculated using the original weight of the core sample. From this data, an average density value was calculated based on lithology.

Since mid-2014, the Seabee mill has been performing a daily density determination from an approximately 5 kg 24-hour composite sample collected from the belt. The samples are analysed on site by water displacement using the following methodology:

•Riffle composite sample down to an approximately 1 kg representative sample;

•Place a dry flask on a 200 g balance and zero the weight;

•Add sample to the flask (greater than 55 g);

•Record the weight of the sample, and zero the balance;

•Fill the flask to marked line with cold water and ensure outside of flask is dry; and

•Place the flask back on the balance and record the weight.

Two 200 ml flasks have been labelled by SSR staff with the water weight when filled to a specified line to be used for the original water weight. The difference between the original water weight and the second reading is equal to the volume of water displaced by the sample, from which the density is calculated using the original weight of the dry sample.

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8 SAMPLE PREPARATION, ANALYSES, AND SECURITY

8.1 Historical Samples

The drillhole sampling, sample preparation, analyses, and security procedures applied prior to 1989 have not been documented in detail.

8.2 Diamond Core Samples (1989 to Present)

Drill core is monitored by SSR staff from the time it is taken out of the ground until it is split and the samples are delivered to the laboratory. Unauthorised personnel are not permitted access to the drill machines or the core logging and splitting facility. Samples that are split for assaying are double bagged within the splitting facility and identified with a coded security tag. Upon receipt of samples at the laboratory, any sample tags that are broken or any sample bags that appear to have been tampered with are reported to SSR.

All underground samples are assayed at the non-accredited Seabee Gold Operation (SGO) laboratory. Samples are dried for 30–60 minutes, crushed to 10 mesh, and riffle split using a Jones splitter until only 200 g of material remains. The samples are then pulverised in a ring and puck pulveriser until greater than 80% passes through a 200 mesh screen. Thirty grams of pulp material is then analysed for gold by fire assay with gravimetric finish using a 0.01 g/t Au detection limit.

Most surface drilling samples are assayed at TSL Laboratories Inc. (TSL) in Saskatoon, Saskatchewan. TSL is independent of SSR. The laboratory was ISO/IEC 17025 accredited until 18 April 2017 and has since withdrawn from the Standard Council of Canada’s system.

Upon receipt of samples, TSL attaches a bar code label to the original sample bag, and the label is scanned to record the sample weight, date, time, equipment used and operator name, allowing for complete traceability of each sample during the laboratory process. Samples are crushed to 70% passing 10 mesh in two stages. The crushed reject is homogenised by passing it once through a Jones riffle splitter down to 250 g and then recombining the two halves, from which 250 g are split using the same riffle splitter. The split is then ring pulverised to 95% passing 200 mesh. Samples are analysed for gold by 30 g fire assay with gravimetric finish using a 0.03 g/t Au detection limit. Pulps and rejects are stored in containers on the TSL laboratory property.

TSL employs comprehensive quality assurance and quality control protocol and control charts for standards assayed at the laboratory show routine performance within two standard deviations of the certified value. The relative precision for gold meets contract specifications and established limits.

8.3 Chip and Muck Samples

Chip and muck samples are bagged, tagged with a unique identification number and transported to the SGO laboratory for analysis following the same methodology as described in Section 8.2.

8.4 Quality Assurance and Quality Control Programmes

In 2006, the SGO geology department introduced an analytical quality assurance and quality control (QA/QC) programme to verify the accuracy of its internal, non-accredited assay laboratory. The programme has since been adopted and modified by SSR and involves the insertion of certified reference material (CRM) standards, duplicate assays, and monthly umpire check assays at an external certified laboratory.

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A Rocklabs Ltd. (Rocklabs) CRM is inserted by a mine geologist at a frequency of one per 20 samples, regardless of the sample type. Three distinct CRM samples are typically cycled through the process; one low-grade, one average grade, and one high grade. The mine geologist records the identification numbers of the CRM samples introduced into the assay stream and checks them as a pass or fail upon receipt of laboratory results. Assay batches with failed CRM results are re-analysed. CRM results are recorded digitally in a spreadsheet provided by Rocklabs to track the pass and fail rates of each of the various reference materials used. The results are compiled in a monthly report and shared with the relevant departments involved in the process.

On a monthly basis, an average of 20 pulp samples are submitted for external analyses by TSL in Saskatoon, Saskatchewan. One CRM is included in each batch of external check samples, and a sieve analysis is performed on one of the pulps to determine percentages passing through –150 and –200 mesh. Results from the analyses at TSL are compared to the on-site results and included in a monthly report.

A blank sample of a coarse-grained quartz-rich rock is inserted after every sample containing visible gold, and pulp duplicates are run every tenth sample by the laboratory. According to SSR, blanks were used and recorded from 2010 to 2014.

SSR reviews the results from the above control samples to accept the data from each individual batch or to reject the data and request a re-run. A batch is rejected if the result for the standard exceeds the tolerance of the 95% confidence level stated on the standard’s certificate. The failure trigger for pulp duplicates is less defined due to the lode-gold nature of the mineralisation; however, batches are considered for re-run when duplicate assay values are greater than ±10%. With respect to coarse-grained blanks, sample batches are rejected if the result is greater than three times the detection limit of the laboratory.

8.5 Conclusions and Recommendations

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9 DATA VERIFICATION

9.1 Verifications by SSR

All exploration and production procedures undertaken by SSR follow detailed procedures and exploration and production data are verified prior to consideration for geological modelling and Mineral Resource estimation. Experienced mine geologists implement industry standard measures to ensure the reliability and trustworthiness of data.

SSR closely monitors analytical quality control data, and upon receipt of results from the lab confirm that sample batches have either passed or failed. Quality control failures are investigated, and failing batches are requested for re-assaying. In addition, monthly check assays are sent for external analysis at TSL and compared to the on-site results. Monthly reports are compiled outlining the performance of analytical quality control data and distributed amongst departments involved in the process.

In 2016, SSR commissioned a review of the exploration and mine geology department databases at the (SGO). The mine databases encompassed the period 2004–2017, whereas the exploration review involved the 2016 database only.

In early 2016, 585 pulp duplicates from the mine database were evaluated from randomly chosen samples, representative of the Santoy deposit. The following assay audit observations were made (Konst, 2016a):

•Of the matched pairs, 60 outliers (10%) were identified to exhibit significant nugget effect. Outliers were defined as matched pairs with a grade difference over 0.1 g/t Au and greater than 100% precision, and those with a grade difference over 0.5 g/t Au and greater than 25% precision.

•A total of 102 of the second pulp analyses returned higher assay values, 116 returned lower assay values, and 367 returned the same value. Of the matched pairs, the original gold analyses had a mean gold grade of 2.19 g/t Au while the second pulp returned a mean value of 2.51 g/t Au, a 13.4% difference.

•A total of 223 matched pairs above the lower reporting limit were considered suitable for precision evaluation. Calculated precision, including outliers, was 24% at a cut-off grade of 3.0 g/t Au and 23% at a cut-off grade of 5.0 g/t Au.

•The evaluation indicated likely issues with gold grain size for the analytical method used. The more erratic higher grade matched pairs represented 27% of the pulp duplicates reporting above the lower detection limit and represented approximately 35% of the gold contained within the matched pair sample set. Improvements to the analytical method, in the form of screen-metallic assays, was suggested.

A total of 54 screen metallic assay results were chosen, based on grade, from the 585 pulp duplicates for evaluation. Konst (2016b) concluded the following from his investigation:

•Evaluation of the screen fire assay results confirmed the presence of significant coarse gold in the selected samples and highlighted its potential impact on grade estimation. The percentage of contained metal present as coarse gold could not be quantified as details were lacking regarding the screen fire assay determinations.

•Samples were selected based on having original average grades greater than 5.0 g/t Au, forming a selection bias. Samples that returned low initial grades, but that may have contained unassayed nuggets of coarse gold that would have been picked up in a screen fire assay, were excluded. This bias exaggerates the apparent impact that coarse gold may have on the deposit as a whole, and a direct comparison of mean gold grades suggested that the original gold grades were overestimated by 23%.

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•Based on pulp duplicates alone, the analytical precision well above detection was 15%. Subtracting it from the pulp-duplicate and screen fire assay precision of 31% indicated that the overall bias between the two datasets was closer to 16%, exaggerating grade by 16%. This, however, is not indicative of the exploration assay data because nugget samples, which reported below their true grade, were not selected for analysis in the screen fire assay study.

•A comprehensive screen fire assay programme was recommended for zones of interest to provide an accurate determination of gold content in deposits of this nature.

•Gold grain size analyses would assist in determining screen size for future screen fire assays and given the coarse nature of gold in the deposit, it was recommended to use a rigorous multi-subsample “no-roll” method of selecting aliquots from pulps and minus fractions for assay.

An audit of drilling and assay data indicated poor precision was noted (23% at 5.0 g/t Au and a detection limit of 0.1 g/t Au) but was assumed to have negligible impact on Mineral Resource estimates due to the abrupt nature of the mineralisation boundaries. Significant re-tooling steps were not deemed necessary, however low-cost steps to improve assay precision, and additional recommendations aimed at improving overall assay and survey accuracy, process efficiencies, and auditability were made.

A portion of the database was compared to the source information to understand the nature and frequency of database errors. Disagreement between surveyed collar azimuths and downhole magnetic surveys of six drillholes, and significant disagreements between high-grade assays and re-runs were the most notable issues identified and are described as follows (Konst, 2016c):

•Variations in azimuth, ranging from 2° to 16°, were observed when corrected FlexIT survey comma delimited files were compared to GEMS database exports of downhole surveys, implying a possible counter-clockwise rotation of the local magnetic field due to local concentrations of magnetic minerals. A non-magnetic survey was recommended where historical drillholes could be re-surveyed.

•A total of 204 assay results from mine geology drill core and quality assurance / quality control samples were randomly selected from mine lab worksheets dated February, March, May, and June 2016 and compared to an August 2016 GEMS database export of assay results. Two high-grade assay re-runs returned nil values, which is what resides in the current database. The errors affect only 1% of the test data, but amount to a simple average grade for the GEMS dataset that is 20% lower than the original laboratory results. A lack of documentation made it unclear as to which results, or combination of results, should be deemed correct and included in the database, and it was recommended that assays greater than 60 g/t Au be reviewed to ensure all relevant assay data is included in the final assay database.

•An addendum to the original Seabee mine site drilling and assay audit involved the inclusion of exploration collar survey data from six randomly selected Santoy drillholes. Variations between the final database and original azimuth ranging from 7.6° to –2.8° reinforced the recommendation that non-magnetic surveys be executed where historical holes are available for re-survey.

Konst (2016d) also conducted a review of the sampling, preparation, and analytical quality assurance of the 2016 Seabee exploration programme and made the following recommendations:

•Blanks were submitted as pre-prepared pulps. It was recommended that barren half-core be used instead and inserted following mineralised samples to properly test for contamination during the sample preparation process.

•CRM analysed at TSL were submitted at a frequency of 1-in-40, however samples were assayed in batches of 20, meaning that only half of the analytical batches were controlled for accuracy. The insertion rate of quality control standards was recommended to be increased to 1-in-20 to test all analytical batches.

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•Improvements in grind quality control at the Seabee mine was highlighted, as TSL’s analytical precision was observed to be significantly better.

•A lack of sufficient data to quantify prep precision and sampling precision, and how it varies with grade, was identified. Crush duplicates to quantify prep precision were recommended to be incorporated into the Seabee quality assurance protocol, and half-core field duplicates over quarter-core were recommended to provide a true measure of sampling precision.

•A comparison of analytical methods indicated that results by fire assay with a gravimetric finish be given priority over fire assay with an atomic absorption spectrometry finish.

•TSL’s greater than 5.0 g/t Au protocol for selecting screen metal assay samples has a selection bias. A site driven protocol such as selecting zones containing visible gold was recommended.

An evaluation of 240 umpire pulp duplicates provided as matched pairs of Seabee mine data from January to November 2016 and January 2017 was performed. The matched pairs were created by taking a second random selection of pulp material from Seabee mine pulp samples and sending them to TSL for check analysis. A total of 238 matched pairs returned results above the reported lower detection limit of 0.03 g/t Au and were therefore suitable for precision analysis. The following observations and recommendations were made by Konst (2017):

•Only one (9%) of the 11 samples sieve tested passed the 95% passing 150 mesh pulverisation specifications. The remaining 10 reported over 80% passing 150 mesh. It was recommended that the mine laboratory increase its efforts to meet grind specifications.

•The 2016 calculated analytical precision was 32% at a cut-off grade of 3.0 g/t Au. The mine versus TSL precision was confirmed by TSL versus TSL pulp duplicate precision, indicating precision issues are not attributable to the Seabee mine laboratory performance, but that gold grain size distribution presents significant challenges for the analysis method used.

•Screen metallic assays previously reported on indicated the average proportion of coarse gold was 32% but could range up to 72% of the total grade. There was no direct correlation between coarse gold and grade, and other geologic controls influenced the distribution of ‘nugget’ gold.

Mr. Konst was subsequently contracted by SSR to perform routine reviews of the monthly quality assurance and quality control results of the SGO.

9.2 Verifications by OreWin

9.2.1 Site Visit

In accordance with S-K 1300 guidelines, OreWin visited the SGO on 6 February 2020, accompanied by representatives of SSR. The OreWin team included Sharron Sylvester BSc (Geol), RPGeo AIG (10125), employed by OreWin Pty Ltd as Technical Director and Graeme Baker, BEng (Mining ), Fellow AusIMM (200051), employed by OreWin Pty Ltd as Principal Mining Consultant.

The site visits took place during active drilling and production activities. All aspects that could materially impact the integrity of the data informing the Mineral Resource estimate (core logging, sampling, analytical results, and database management) were reviewed with SSR staff. OreWin spoke with mine staff to ascertain exploration and production procedures and protocols. OreWin observed core from six drillholes and confirmed that the logging information accurately reflects actual core. The lithology contacts checked by OreWin match the information reported in the core logs. OreWin toured the underground operations at Santoy and assessed the attributes of the shear-hosted gold-sulfide-chlorite-quartz veins.

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9.2.2 Verifications of Analytical Quality Control Data

To assess the accuracy and precision of analytical quality control data, OreWin routinely analyses such data. Analytical quality control data typically comprises analyses from standard reference material, blank samples, and a variety of duplicate data. Analyses of data from standard reference material and blank samples typically involve time series plots to identify extreme values (outliers) or trends that may indicate issues with the overall data quality. To assess the repeatability of assay data, several tests can be performed, of which most rely on certain statistical tools. OreWin routinely plots and assesses the following charts for duplicate data:

•Bias charts

•Quantile-quantile (Q-Q) plots

•Mean versus half relative deviation (HRD) plots

•Mean versus half absolute relative deviation plot

•Ranked half absolute relative deviation (HARD) plot

9.2.3 Discussion

OreWin analysed the available analytical quality control data of the SGO to confirm that the analytical results are reliable for informing Mineral Resource estimates. All data were provided in Microsoft Excel spreadsheets from SSR, and OreWin aggregated the assay results for the external quality control samples for further analysis. Control samples (blanks and CRM) were summarised on time series plots to highlight the performance of the control samples. Field duplicates and umpire laboratory pulp duplicates were analysed using bias charts, quantile-quantile, and relative precision plots.

The analytical quality control data produced between 2010 and early 2017 are summarised in Table 9.1. The data produced on the SGO represents approximately 4.1% of the total number of samples.

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Table 9.1 Summary of Analytical QA/QC Data

Sample	ALS	(%)	TSL	(%)	Seabee	(%)	Total	(%)	Comment
Sample count	181,288		71,325		59,607		312,220
Blanks	310	0.17	155	0.22	849	1.42	1,314	0.42
QC samples	624	0.34	257	0.36	4,139	6.94	5,020	1.61
CRMs
- SE29	16	0.01	17	0.02			33	0.01	0.597 g/t Au
- SE44					98	0.16	98	0.03	0.660 g/t Au
- SF57			11	0.02	139	0.23	150	0.05	0.848 g/t Au
- SF85					131	0.22	131	0.04	0.848 g/t Au
- SG40	174	0.10	41	0.06	95	0.16	310	0.10	0.976 g/t Au
- SG56					191	0.32	191	0.06	1.027 g/t Au
- SH35	57	0.03	48	0.07			105	0.03	1.323 g/t Au
- SH24	18	0.01	23	0.03	14	0.02	55	0.02	1.326 g/t Au
- SH82					43	0.07	43	0.01	1.333 g/t Au
- SH41	157	0.09	27	0.04			184	0.06	1.344 g/t Au
- SH69					652	1.09	652	0.21	1.346 g/t Au
- SH65					54	0.09	54	0.02	1.348 g/t Au
- SJ63					685	1.15	685	0.22	2.632 g/t Au
- SJ53					96	0.16	96	0.03	2.637 g/t Au
- SJ80					160	0.27	160	0.05	2.656 g/t Au
- SK62					98	0.16	98	0.03	4.075 g/t Au
- SL46	56	0.03	44	0.06	97	0.16	197	0.06	5.867 g/t Au
- SL51	79	0.04			92	0.15	171	0.05	5.909 g/t Au
- SL61	35	0.02	30	0.04	301	0.50	366	0.12	5.931 g/t Au

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- SL76					131	0.22	131	0.04	5.960 g/t Au
- SN16					90	0.15	90	0.03	8.367 g/t Au
- SN60					58	0.10	58	0.02	8.595 g/t Au
- SN50					97	0.16	97	0.03	8.685 g/t Au
- SP59					354	0.59	354	0.11	18.12 g/t Au
- SP73					121	0.20	121	0.04	18.17 g/t Au
- SQ36	19	0.01	15	0.02			34	0.01	30.04 g/t Au
- SQ48	13	0.01	1	<0.01	342	0.57	356	0.11	30.25 g/t Au
Field duplicates	186	0.10	73	0.10			259	0.08
Check assays			902	1.26			902	0.29
Total QC Samples	1,120	0.62%	1,387	1.94%	4,988	8.37%	12,695	4.07%

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Blank samples analysed at the Seabee mine laboratory, TSL, and historically at ALS Limited (ALS), indicated acceptable performance. Several samples, however, yielded values above the warning limit (defined as ten times the lower detection limit), though, this occurred 5% of the time or less at each laboratory. Further examination identified a number of blank samples analysed at the Seabee mine laboratory between 2010 and 2011 that displayed anomalously high gold grades and indicated potential contamination during the sample preparation process or possible mislabelling of blank material. After 2011, the abundance of failed blanks appeared to be rectified, with all blank samples assaying at or below the warning limit. Post-2013, however, blank material has not been submitted to the Seabee mine laboratory. OreWin strongly recommends that blank material, such as barren half core, be inserted routinely into the sample stream to monitor any potential contamination during sample preparation.

SSR uses a series of CRM (standards) which are submitted with mine geology samples at the Seabee mine laboratory, and with exploration samples at TSL, and historically at ALS. Standards submitted to the Seabee mine laboratory and TSL largely performed within expected ranges, and mean grades are similar to expected values. Several significant outliers, however, have been observed, which are likely attributed to the mislabelling of other standards used at the time or from the possible mislabelling of blank material. Standards submitted to ALS demonstrate an overall worse performance than those submitted to TSL; however, due to the historical nature of the samples, the cause of the deficiency remains unknown. OreWin recommends that SSR continue to monitor the performance of standards and investigate and identify the cause of any significant outliers.

Monthly umpire check assaying is performed at TSL of pulp duplicate samples processed at the Seabee mine laboratory. HARD plots suggest that approximately 60% of umpire samples have HARD below 10%, indicating that the umpire laboratory had difficulty consistently reproducing pulp assay results from Seabee mine laboratory.

9.3 QP Opinion

In the opinion of the QP the data is adequate for the purposes used in the Seabee21TRS. No material sample bias was identified during the review of the drill data and assays. Observation of the drill core during the site visit and inspection and validation of the data collected indicate that the drill data is adequate for the estimation of Mineral Resources.

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10 MINERAL PROCESSING AND METALLURGICAL TESTING

10.1 Style of Mineralisation

10.2 Metallurgical Investigations

10.2.1 Metallurgical Testwork

With the introduction of Santoy ore to the process plant metallurgical testing of drill composites has been undertaken. The composites selected from the diamond drilling represent the footwall, centre and hanging wall of the stacked vein zones.

The results of the testwork indicate the following key metallurgical parameters:

•Diagnostic leach testing of the Master Composite indicated that 99% of the gold was extractable by cyanide leach, indicating the material is free milling.

•In the Master Composite approximately 55% of the gold grains were >100 micron in size. Indicating the gold is gravity recoverable.

•High gravity recoverable gold of up to 91% to gravity concentrate at a 0.18% mass pull

•High cyanide gold recovery of gravity tailings at 95%.

•The overall gold recovery, by gravity and cyanide leach, was 95% to 99% for the size samples tested.

•SSR is not aware of any processing factors or deleterious elements that could impact potential economic extraction.

10.2.2 Process Plant Improvements

The Figure 10.1 shows the average annual milled tonnes per day. The current Seabee process plant capacity is nominally 1,320 tpd or 1,240 tpd annual average.

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Figure 10.1 Historical Annual Milled Daily Throughput

SSR, 2021

While throughput has increased since commencement of operations as shown in Figure 10.1, metallurgical performance has improved with improvement in metallurgical control with a consistent trend in reduction of tailings losses, even with increasing head grade, as shown in Figure 10.2.

Figure 10.2 Historical Annual Head and Tailings Grade Trend

SSR, 2021

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A number of further capacity improvement projects are being investigated including:

•Optimisation of the gravity gold Acacia recovery circuit.

•Optimisation of grinding cyclone performance.

•Improvements in thickener performance and flocculant usage.

•Improvements in carbon management, with recovery of fine carbon and carbon activity improvement.

These programmes are expected to provide further improvements in throughput, gold recovery and reduction in operating costs.

10.3 Recovery Estimates

Figure 10.3, indicates the Mill feed grade and recovery for 2017 to 2020.

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Figure 10.3 Mill Feed Grade and Gold Recovery 2017–2020

SSR, 2021

The future recovery estimates of 98% are based on the recent mill performance with mill recoveries of more than 96.5%.

10.4 QP Opinion

In the opinion of the QP the data is adequate for the purposes used in the Seabee21TRS and the analytical procedures used in the analysis are of conventional industry practice.

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11 MINERAL RESOURCES ESTIMATES

The Mineral Resources for the Seabee Gold Operation (SGO) comprise the Santoy 8, 9, GHW, and the Porky West deposits. The deposits are mined by underground mining methods.

The Mineral Resources estimates for SGO were completed by the SSR technical department on site. The Seabee21TRS QPs have reviewed and accepted this information for use in the Seabee21TRS.

This section summarises the Mineral Resource estimation methods and the key assumptions and parameters.

11.1 Resource Modelling Methods

Cell modelling techniques were used for Mineral Resource evaluation for all deposit areas. The resource models used to report Mineral resource estimates were created using data to 31 December 2020.

Geovia GEMS software was used to construct the geological solids, conduct geostatistical analysis and variography, construct the cell model, estimate metal grades, and to report the Mineral Resource. Cell modelling methodologies have been adapted and refined from previous audits in 2011, 2014, and 2016.

The Mineral Resource evaluation methodology involves the following procedures:

•Database compilation and verification.

•Construction of wireframe models for the boundaries of the gold vein mineralisation.

•Data conditioning (compositing and capping) for geostatistical analysis and variography.

•Cell modelling and grade interpolation.

•Definition of Mineral Resource classification domains and validation.

•Assessment of “reasonable prospects for eventual economic extraction”.

•Preparation of the Mineral Resource Statement.

11.1.1 Santoy Mine

The Santoy mine is comprised of three zones: Santoy 8, Santoy 9, and Gap Hangingwall (GHW). Mineral Resources are reported for all three areas. Mining and drilling at Santoy 8 has defined sub-parallel ore bodies dipping from 40°–60° east and plunging to the north. Similarly, mining and drilling at Santoy 9 has defined sub-parallel veins dipping from 45°–55° east and plunging to the north. Drilling and mining at the GHW has identified a body of north–south oriented quartz veins occupying the hinge of a folded limb of the Lizard Lake Pluton. The mineralised envelope plunges to the north and the dip of the orebody ranges from <40° to 50°. Both the lateral and horizontal limits of the Santoy 8 and Santoy 9 veins have been defined through underground development and core drilling from surface and underground. The central area of the GHW has seen underground development on two levels. The horizontal and vertical extremities of the orebody are currently only defined by surface and underground diamond drilling.

The Santoy mine drill database contains 2,030 underground diamond drillholes for 394,345 m, and 774 surface diamond drillholes for 248,048 m.

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Mineral Resources for all veins from the Santoy mine are estimated using cell modelling methods. The Santoy 8 veins are: 8A, 8B, 8C, 8D, 8F and 8G. The Santoy 9 veins are: 9A, 9B, and 9C. GHW is reported as a single entity. Drilling data are used to generate 10–30 m spaced vein sectional and/or plan polylines. Vein polylines are linked to create vein solids that are used to code Mineral Resource cells.

The individual assays are composited to 1.0 m length for all veins. Residual composites less than 10% of the composite length are excluded. Grade capping is applied on composites in each vein separately. Several capping assessments were undertaken, incorporating the use of histograms, cumulative frequency curves, and probability plots. Statistical impacts are also verified with cap percentiles, coefficient of variation, and changes in mean values. Capping at Santoy 8 ranges between 15–110 g/t Au, Santoy 9 ranges between 91–110 g/t Au, and GHW is capped at 45 g/t Au (Table 11.1).

Table 11.1 Capping Values at Santoy Mine

Vein	Capping <br>(g/t Au)
Santoy 8A	110
Santoy 8B	25
Santoy 8C	29
Santoy 8D	29
Santoy 8F	45
Santoy 8G	15
Santoy 9A	110
Santoy 9B	91
Santoy 9C	110
Gap Hangingwall	45

A combination of linear semi-variography and variography is performed on the capped composited data to determine the variograms, search ellipses, and estimation parameters. A standard cell size of 3 m x 3 m x 3 m is used. Ordinary kriging is used to interpolate gold grades in the cell model. For the GHW the broadest interpolation is inverse distance squared (ID2).

The density assigned to veins is 2.75 t/m3 at Santoy 8 and Santoy 9 and GHW ore is assigned a density of 2.65 t/m3. Waste density is set to 2.91 t/m3. Density values are based on the average density of samples measured using the water displacement method.

11.1.2 Porky Deposit Area

The Porky deposit area is comprised of two zones: Porky West and Porky Main. Drilling and mining at Porky West has defined sub-vertical structures dipping approximately 65° to the south-west. Drilling at Porky Main defined shear zones plunging at about 45° to the south-east.

The Porky West drill database contains 89 surface diamond holes for 17,647 m drilled between 2003 and 2009. In addition, 166 underground chip and muck sample traverses (1,291 samples) were completed. The polygonal Mineral Resource estimation at Porky West was completed in 2009 using GEMS software and verified recently by the mine geology team. A capping value of 15 g/t Au was determined using the 95th percentile. Density of 2.70 t/m3 was used based on testwork completed at the Seabee assay lab. Due to the 65° dip of the orebody, the GEMS polygonal resources were estimated using an inclined longitudinal section method.

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Porky Main was estimated using polygonal methods by Claude Resources in 2005. However, SSR has not been able to verify the results of this polygonal estimate, therefore Porky Main is not included in the Mineral Resource Statement contained herein but may be included in the future pending additional drilling and modelling.

11.2 Cell Model Validation

SSR used a variety of methods to validate the Mineral Resources determined by cell modelling.

Validation of the high-grade capping thresholds was performed by an independent selection of capping values. Capping for each domain was based on probability plots and a proprietary statistical utility.

For all domains, SSR validated the cell model using a visual comparison of model estimates and the drillhole composites for each domain on sections and plans. The grades can be seen to follow the orientation of the search ellipses. Visual validation of model grades in addition to reconciliation data, as described in Section 11.6, have been the primary methods of cell model validation.

11.3 Mineral Resource Classification

Industry best practices suggest that Mineral Resource classification should consider the confidence in the geological continuity of the mineralised structures, the quality and quantity of exploration data supporting the estimates, and the geostatistical confidence in the tonnage and grade estimates. Appropriate classification criteria should aim at integrating these concepts to delineate regular areas at similar resource classification.

SSR is satisfied that the geological modelling honours the current geological information and knowledge. The location of the samples and the assay data are sufficiently reliable to support Mineral Resource evaluation. The sampling information was acquired primarily by closely spaced surface and underground core drilling and supported by underground development and chip sampling.

SSR considers that the gold mineralised zones show good geological continuity, respecting the direction of maximum continuity, and defined by an adequate drill spacing with reliable sampling information allowing classified within the meaning of the S-K 1300 Regulations for Mineral Resources and Mineral Reserves. Mineral Resources are reported within wireframed classification domains at a specified cut-off grade determined annually. The classification parameters used to define classification domains are detailed in Table 11.2.

Generally, for gold mineralisation exhibiting good geological continuity, SSR considers that zones can be classified as Measured if one (or more) of the following criteria is applicable:

•The zone is sampled in two-dimensions by mine development within a maximum of 1 or 2 sublevel spacing.

•The zone is sampled in one dimension by mine development and informed by core drilling at a drill spacing of less than 25 m while respecting the direction of maximum continuity.

•Despite no adjacent sampled mine development, the drill spacing is less than 15 m while respecting the direction of maximum continuity.

Similarly, SSR considers that gold mineralised zones can be classified as Indicated if the zone is sampled in one dimension by mine development and informed by core drilling at a drill spacing of less than 35 m while respecting the direction of maximum continuity; or drilled at a spacing of 25 m or less with no adjoining underground development.

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Table 11.2 Parameters for Mineral Resource Classification

Classification				Areas – Underground Development
	Two Dimensions		One Dimension		None
Measured		Distance from Development		1 or 2 sublevel spacing		Projected no more than the spacing of 2 sublevels		–
	Drill Spacing		–		Closely spaced drilling on the same structure (~<25m)		Drill spacing of ~<15m
Indicated		Distance from Development		–		Projected no more than the spacing of 4 sublevels		–
	Drill Spacing		–		Closely spaced drilling on the same structure (~<35m)		Drill spacing of ~<25m
Inferred		Distance from Development		–		–		–
	Drill Spacing		–		Closely spaced drilling on the same structure (~<75m)		Taken to the extents of the inferred search ellipse while being subject to geological interpretation

Conversely, gold mineralised zones sampled in one direction of mine development and informed by drill spacing of less than 75 m or estimated at the extent of the search ellipse can appropriately classified in the Inferred category because the confidence in the estimate is insufficient to allow for the meaningful application of technical and economic parameters or to enable an evaluation of economic viability.

11.4 Reasonable Prospects for Eventual Economic Extraction

11.5 Mineral Resource Statement

The Mineral Resources estimates for SGO were completed by the SSR technical department on site. The Seabee21TRS QP has reviewed and accepted this information for use in the Seabee21TRS. The Seabee21TRS QP reviewed the assumptions, parameters, and methods used to prepare the Mineral Resource Statement and is of the opinion that the Mineral Resource is estimated and prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300).

The Mineral Resources are estimated based on cell models representative of the mineralised veins and using an assumed gold price of $1,750/oz.

The gold price of $1,750/oz was selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The metal price is representative of the range of price estimates publicly reported for Mineral Resource cut-offs. The Mineral Resources are assumed to be mined by underground stoping.

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In determining the cut-off grade, the reasonable prospects for eventual economic extraction requirement generally implies that the quantity and grade estimates meet certain economic thresholds considering an underground stoping extraction scenario and on-site processing. Using this operating scenario, the cut-off grade is estimated to be 2.07 g/t Au.

Mineral Resources are reported exclusive of Mineral Reserves and have been summarised by project and resource classification in Table 11.3.

Table 11.4 shows the cut-off values and metallurgical recoveries associated with the Mineral Resources.

Table 11.3 Summary of Seabee21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price

Area		Mineral Resources Classification
	Measured				Indicated				Measured + Indicated				Inferred
	Tonnage<br>(kt)		Grade (Au g/t)		Tonnage<br>(kt)		Grade (Au g/t)		Tonnage<br>(kt)		Grade (Au g/t)		Tonnage<br>(kt)		Grade (Au g/t)
Santoy Mine		71		19.75		745		12.74		816		13.35		2,238		6.43
Porky West		–		–		52		5.03		52		5.03		516		4.42
Total SGO		71		19.75		797		12.23		869		12.85		2,754		6.05

1. Mineral Resources are reported based on 31 December 2021 as-mined survey data.

2. Mineral Resources are reported exclusive of Mineral Reserves

3. Mineral Resources are shown on a 100% basis

4. The Mineral Resources estimates are based on a 2.07 g/t Au cut-off with a gold price assumption of $1,750/oz.

5. Santoy Mine includes Santoy 8, Santoy 9, and GHW lodes.

6. The Mineral Resources in the Seabee21TRS were assessed for reasonable prospects for eventual economic extraction by reporting only material that fell within conceptual underground shapes.

7. SSR has 100% ownership of the Project.

8. The point of reference for Mineral Resources is the point of feed into the processing facility.

9. Tonnage is metric tonnes and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

Table 11.4 Summary of Cut-off Values and Metallurgical Recoveries of Seabee21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021) Based on $1,750/oz Gold Price

Mineral Resources <br>Classification	Tonnage<br><br>(kt)	Grade <br><br>(Au g/t)	Contained Gold<br><br>(koz)	Cut-off Value<br><br>(Au g/t)	Metallurgical Recovery<br><br>(%)
Measured	71	19.75	45	2.07	98
Indicated	797	12.23	313	2.07	98
Measured + Indicated	869	12.85	359	2.07	98
Inferred	2,754	6.05	536	2.07	98

1. Mineral Resources are reported based on 31 December 2021as-mined survey data.

2. Mineral Resources are reported exclusive of Mineral Reserves

3. Mineral Resources are shown on a 100% basis

4. The Mineral Resources estimates are based on a 2.07 g/t Au cut-off with a gold price assumption of $1,750/oz.

5. Santoy Mine includes Santoy 8, Santoy 9, and GHW lodes.

6. The Mineral Resources in the Seabee21TRS were assessed for reasonable prospects for eventual economic extraction by reporting only material that fell within conceptual underground shapes.

7. SSR has 100% ownership of the Project.

8. The point of reference for Mineral Resources is the point of feed into the processing facility.

9. Tonnage is metric tonnes, ounces represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

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11.6 Reconciliation

The SGO routinely compares the Mineral Resource and Mineral Reserve models with production results. As an example, the yearly grade reconciliation between the Mineral Resource model and the actual mined grade from the Santoy underground workings for the period 2020 to 2021, is presented in Table 11.5. The reconciliation between the Mineral Resource model and recovered grades is reasonable. This demonstrates that the Mineral Resource model adequately predicts grades achieved during mining.

Table 11.5 Annual Grade Reconciliation at Santoy for 2020 and 2021

Period	M&I Mineral Resources Estimate Grade<br>(g/t Au)	Mine Grade<br><br>(g/t Au)	Variance
2020	10.61	10.40	–2%
2021	10.38	10.11	–3%

11.7 Comparison with Previous Estimates

The Mineral Resources have been compared to the previous Mineral Resources. Comparison of the 2021 Mineral Resource with the 2020 Mineral Resources shows a net decrease in contained gold of 36 koz (–2.4%).

Key changes in the Mineral Resources (contained metal) have resulted from:

•Santoy mine depletion (–8.3%)

•Santoy 8 conversion (+1.5%)

•Santoy 9 conversion (+0.9%)

•Gap HW conversion (+5.5%)

•Porky West remodel (–2.1%)

11.8 QP Opinion

The Seabee21TRS QP has not identified any relevant technical and/or economic factors that require resolution with regards to the Mineral Resources estimate.

The Seabee21TRS QP reviewed the assumptions, parameters, and methods used to prepare the Mineral Resources Statement and is of the opinion that the Mineral Resources are estimated and prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300).

11.9 Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects

The Mineral Resources reported in the Seabee21TRS are suitable for reporting as Mineral Resources using the Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

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12 MINERAL RESERVES ESTIMATES

This section summarises the key assumptions, parameters, and methods used in the preparation of the Mineral Reserves Statement for the Seabee Gold Operation (SGO).

The Mineral Reserves estimate was completed by the SSR technical department on site at the SGO. The QP reviewed the assumptions, parameters, and methods used to prepare the Mineral Reserve Statement and is of the opinion that the Mineral Reserve is estimated and prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300).

Access underground at the Santoy mine is provided from surface at the Santoy portal via a main ramp, with sublevels spaced between 17–20 m vertically. Mining is carried out using sublevel open stoping mining methods with backfill.

Stopes are filled with a combination of rock fill (RF) and cemented rock fill (CRF), mined in a bottom-up mining sequence. Sill pillars are mined on retreat once the stopes below and above have been mined (stopes above filled with CRF and allowed to cure).

The estimated cut-off grade for Mineral Reserves was based on a $1,600/oz gold price and current operating costs and metallurgical performance. Table 12.1 details the basic parameters used for Mineral Reserves definition.

Table 12.1 Mineral Reserves Input Parameters

Item	Unit	Rate
Minimum Mining Width	m	1.8
Hangingwall Dilution	m	0.18
Footwall Dilution	m	0.18
Minimum Dip	degrees	45
Maximum Stope Length	m	20
Mining Recovery	%	94
Process Recovery	%	98
Gold Price	$/oz	$1,600

The gold price of $1,600/oz was selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The metal price is representative of the range of price estimates publicly reported for Mineral Reserve cut-offs.

Dilution and mining recovery factors were derived from ongoing stope reconciliations using actual mucking and cavity monitor survey data.

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A cut-off grade of 2.52 g/t Au was used to estimate the Mineral Reserves. The cut-off grade was determined based on the following:

•Gold price of $1,600/oz

•Exchange rate of C$1.26:US$1.00

•Average milling recovery of 98%

•Royalty of 3.0%

•Payable factor of 99.5%

•Refinery charge of $3.09/oz

•Operating cost of $128/t

12.1 Mineral Reserves Statement

The SGO Mineral Reserves estimate was completed by the SSR technical department on site. The Seabee21TRS QP has reviewed and accepted this information for use in the Seabee21TRS. The Seabee21TRS QP reviewed the assumptions, parameters, and methods used to prepare the Mineral Reserves Statement and is of the opinion that the Mineral Reserves are estimated and prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300).

The Mineral Reserves Statement is reported in Table 1.3 and Table 1.4. The reference point at which the Mineral Reserves are identified is where ore is delivered to the processing plant (i.e., mill feed). The QPs are unaware of any environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant issues that may materially affect the Mineral Reserves estimate. However, the Mineral Reserves may be affected by further infill and exploration drilling that may result in increases or decreases in subsequent Mineral Resources and Mineral Reserves estimates. The Mineral Reserves may also be affected by subsequent assessments of mining, environmental, processing, permitting, taxation, socio-economic, and other factors. The effective date of the Mineral Reserves Statement is 31 December 2021.

12.2 Comparison with Previous Estimates

The 2021 Mineral Reserves represent a net increase of 86 koz (18%) total contained gold ounces as compared with the 2020 Mineral Reserves. Although mining depletion has occurred in the Santoy 8A and 9A mining zones, the 2021 Mineral Reserves have increased the conversion of the Santoy Mineral Resources in the GHW zone into a Mineral Reserves. An increase in the gold commodity price has also resulted in a decrease in the Mineral Reserves cut-off grade.

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Table 12.2 Summary of Seabee21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,600/oz Gold Price

Area		Mineral Reserves Classification
	Proven				Probable				Total
	Tonnage<br>(kt)		Grade (Au g/t)		Tonnage<br>(kt)		Grade (Au g/t)		Tonnage<br>(kt)		Grade (Au g/t)
Santoy Mine		304		9.16		2,379		6.40		2,684		6.72

1. Mineral Reserves are reported based on 31 December 2021as-mined survey data.

2. The Mineral Reserves estimate is based on metal price assumptions of $1,600 gold.

3. The Mineral Reserves estimate is reported at a cut-off grade of 2.52 g/t Au.

4. Economic analysis for the Mineral Reserves has been prepared using long-term metal prices of $1,600/oz of gold.

5. No mining dilution is applied to the grade of the Mineral Reserves. Dilution intrinsic to the Mineral Reserves estimate is considered sufficient to represent the mining selectivity considered.

6. Processing recoveries vary based on the feed grade. The average recovery is estimated to be 98%.

7. SSR has 100% ownership of the Project.

8. Santoy Mine includes Santoy 8, Santoy 9, and Gap Hangingwall lodes.

9. Metals shown in this table are the contained metals in ore mined and processed.

10. The point of reference for Mineral Resources is the point of feed into the processing facility.

11. Tonnage is metric tonnes and g/t represents grams per metric tonne.

12. Totals may vary due to rounding.

Table 12.3 Summary of Cut-off Values and Metallurgical Recoveries, of Seabee21TRS Mineral Reserves Estimate (as at 31 December 2021) Based on $1,600/oz Gold Price

Mineral Reserves Classification	Tonnage<br><br>(kt)	Grade <br><br>(Au g/t)	Contained Gold<br><br>(koz)	Cut-off <br>Value<br><br>(Au g/t)	Metallurgical Recovery<br><br>(%)
Proven Mineral Reserves	304	9.16	90	2.52	98
Probable Mineral Reserves	2,379	6.40	490	2.52	98
Total Mineral Reserves	2,684	6.72	580	2.52	98

1. Mineral Reserves are reported based on 31 December 2021as-mined survey data.

2. The Mineral Reserves estimate is based on metal price assumptions of $1,600 gold.

3. The Mineral Reserves estimate is reported at a cut-off grade of 2.52 g/t Au.

4. Economic analysis for the Mineral Reserves has been prepared using long-term metal prices of $1,600/oz of gold.

5. No mining dilution is applied to the grade of the Mineral Reserves. Dilution intrinsic to the Mineral Reserves estimate is considered sufficient to represent the mining selectivity considered.

6. Processing recoveries vary based on the feed grade. The average recovery is estimated to be 98%.

7. SSR has 100% ownership of the Project.

8. Metals shown in this table are the contained metals in ore mined and processed.

9. The point of reference for Mineral Reserves is the point of feed into the processing facility.

10. Tonnage is metric tonnes, ounces represent troy ounces, and g/t represents grams per metric tonne.

11. Totals may vary due to rounding.

12.3 Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects

The Mineral Reserves reported in the Seabee21TRS are suitable for reporting as Mineral Reserves using the Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

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13 MINING METHODS

13.1 Introduction

This section summarises the key components of the mine plan that form the basis of extracting the Mineral Reserves. Actual data and current operating practice are referenced heavily as the mine plan is based on the successful continuation of current practice. Mining at Seabee is now complete therefore all mining references in this Seabee21TRS relate to the Santoy mine.

The mine plan was initially completed by the SSR technical department on site at the SGO and included plans for the overall extraction of the Mineral Resources. OreWin has since reviewed the mine plan and made the appropriate modifications to include extraction of only the Mineral Reserves.

The LOM plan of the Mineral Reserves at the SGO, commencing 1 January 2022, includes 2.68 Mt at an average grade of 6.7 g/t Au. The Mineral Reserves estimate includes dilution from hangingwall and footwall overbreak based on ongoing stope reconciliation. A total of 580 koz of gold will be delivered to the mill.

Stopes are filled with a combination of rock fill (RF) and cemented rock Fill (CRF), mined in a bottom-up mining sequence. Sill pillars are mined on retreat once the stopes below and above have been mined (stopes above filled with CRF and allowed to cure).

Mining factors are derived from ongoing analysis of site performance data.

Ore is trucked to surface and dumped on a surface mine ore pad. Ore is then loaded into surface trucks for a 14 km haul to the run-of-mine (ROM) stockpile at the mill, located at the old Seabee mine. A longitudinal section of the existing Santoy mine is provided in Figure 13.1 and Figure 13.2 and Figure 13.3 show the Santoy LOM design.

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Figure 13.1 Santoy Mine – Existing Development

OreWin, 2021

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Figure 13.2 Santoy Mine – 2021 Life-of-Mine Development Design

OreWin, 2021

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Figure 13.3 Santoy Mine – 2021 Life-of-Mine Stope Designs

OreWin, 2021

13.2 Mining Methods

The primary mining method at the Santoy mine is longitudinal retreat (longhole open stoping). Avoca mining (which is similar in its process) is also used when access from both sides of the sill is available. The minimum mining width is 1.8 m. Planned stopes typically range in width from 3.9–9.0 m. The length of the stopes varies based on deposit geometry and geotechnical guidance. Typical stope extents range from 20–30 m along strike, with a maximum stope strike length of 40 m. External dilution is included based on ongoing stope reconciliations using actual mucking and cavity monitor survey data.

Level spacing varies from 17–20 m (vertical), floor-to-floor. The sill drifts on the levels are connected to a ramp to permit access for the rubber-tired mobile equipment fleet.

Longhole drills are used to drill down from the top level to breakthrough into the bottom level of the stope.

For localised areas with minimal strike length, Alimak mining methods, or captive longhole mining methods, are used to reduce lateral development costs. Access to the captive stopes is provided via an Alimak raise. Where the stope extends less than 15 m from the Alimak raise, production drilling is completed from the Alimak raise climber. For larger stopes, sub-drifts are driven to permit access for a longhole drill.

On completion of mining, stopes are backfilled with development waste rock. A cemented waste rock rib pillar is placed when mining occurs adjacent to backfill stopes. Based on stope sequence and long-term access requirements, some stopes can be left open without backfill.

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13.3 Primary Access

Primary access is provided via a main ramp from the surface to the deepest levels. The main ramp begins at the Santoy portal, which is located at the top of the Santoy 8 deposit. Depending on the geometry of the deposit, the main ramp splays into secondary ramps to access longitudinally separated sections of the mine. The ramps are driven 5 m wide x 5 m high to permit access for the 45 t haulage trucks. Extensions to the existing ramps and additional ramps will be driven to provide access to planned mining levels.

13.4 Level Design

Mining levels are driven from access crosscut drifts from the ramps. A standoff distance of at least 25 m from the ramp to the orebody is used when accessing from the footwall side, and 20 m when accessing from the hangingwall side. Typical level infrastructure excavations include a sump and a truck dump / remuck. Some levels include a drift to access and transfer ventilation from the levels above or below. Access to the longhole stopes is accomplished with sill drifts driven from the access crosscuts. The sill drifts are used for deposit definition, production mucking, and production drilling. The sill drifts are driven 4.2 m high with the width varying to suit the width of the ore (minimum 4.6 m), to a maximum of 8 m.

For a typical Alimak stope, level development consists of a haulage drift, a sill drift, and an Alimak chamber on the lower level and an access drift on the upper level.

The permanent and temporary excavation dimensions at the Santoy mine are provided in Table 13.1. Temporary excavations are openings that are typically only used for less than two years before they are shut down or backfilled.

Table 13.1 Excavation Dimensions

Category	Width<br><br>(m)	Height<br><br>(m)
Permanent Excavations
Ramp	5.0	5.2
Remuck	5.0	5.2
Truck Dump	5.0	6.7
Haulage	4.6	4.2
Sump	4.0	4.0
Refuge Station	6.0	5.0
Safety Bay	1.5	2.4
Ventilation Access	4.6	4.2
Ventilation Raise – Alimak	3.0	3.0
Ventilation Raise – Longhole	4.0	4.0
Temporary Excavations
Sill	8.0 * max.	4.2
Access	4.6	4.2
Alimak Chamber	4.2	4.2
Alimak Nest	4.2	6.5
Alimak Sublevel	7.0* max.	3.0

* Width varying to suit the width of the mineralised zone

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13.5 Material Handling

Ore and waste are hauled via 45 t haulage trucks. Ore is transported to surface where it is dumped and transferred via a wheeled loader into 40 t articulated dump trucks. The ore is then hauled 14 km to the mill ROM stockpile located near the old Seabee mine. Waste rock remains underground for deposition into mined stopes where possible, otherwise stockpiled on surface for later back haul as backfill.

13.6 Ventilation

The Santoy primary ventilation circuit currently provides 170 m3 per second (360,000 ft3 per minute (CFM)) downcasting through two ventilation raises located centrally at the Gap Main fresh air raise (FAR) and the Santoy 8 Main FAR, and exhaust via the main ramp and the Santoy 8 East return air raise (RAR). The Gap Main FAR provides fresh air via two fans in parallel with a total power of 597 kw (800 hp), while the Santoy 8 Main FAR provides fresh air via a single fan with a total power of 149 kw (200 hp).

From the Gap Main FAR fans, roughly 113 m3 per second (240,000 CFM) is sent down the raise until it reaches 31L vent drift where roughly 42 m3 per second (90,000 CFM) is split off and sent down to the bottom of the Gap decline with the help of a 74 kw (100 hp) booster fan located in 31L vent drift. The remaining 56 m3 per second (120,000 CFM) continues down a system of raises on the 41 decline until it reaches the bottom of the face. There is a 149 kw (200 hp) booster fan located on 47L of the 41 decline to assist with the movement of air to this area.

The Santoy 8 Main FAR pushes roughly 56 m3 per second (120,000 CFM) from surface to the bottom of 48 decline through a system of raises. At the bottom of this system are two 74 kw (100 hp) booster fans in parallel that assist with airflow. From here, 38 m3 per second (80,000 CFM) moves up the 49 incline and provides ventilation at the face, while the remaining 19 m3 per second (40,000 CFM) is sent up the 48 decline and returns to the main exhaust system.

The Gap Hangingwall (GHW) development is currently ventilated using auxiliary fans (located at 46L on the Gap decline) and ducting pulling from the 90,000 + 40,000 CFM provided by the Gap decline and the 8 Main FAR respectively. General ventilation plans have been established for the future LOM at Santoy, with ventilation raises being developed at each working level as mining continues. No additional significant capital expenditures have been identified in the current LOM Ventilation plans.

Fresh air is heated by existing propane heaters during the colder winter months.

13.7 Backfill

Rock fill (RF) and cemented rock fill (CRF) using mine waste rock is used for backfill at the Santoy mine. Waste rock is stockpiled on surface temporarily when open stopes are unavailable for deposition.

CRF is used, but not limited to, the creation of sill pillars at the start of a mining front. This is implemented to create a solid barrier between mining fronts. The CRF consists of run-of-mine waste mixed with a cement binder. Santoy mine uses 5% binder in CRF backfilling for sill pillar creation and as low as 3% binder is used for less critical locations. The cement for the backfill comes from a bagged dry-mix product that is turned into useable wet cement near the workplace using a transportable mixer. The bagged dry-mix is stored on surface and brought underground as required.

Based on the underground development design and schedule, there will be sufficient backfill material available from development waste rock.

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13.8 Dewatering

Main dewatering is accomplished via main sump / pumping station on 28L and 30L at the Santoy deposit. The water is pumped to the Santoy 8 deposit underground settling sump and then to the surface mine water management pond located near the Santoy portal.

The Santoy mine dewatering requirements are summarised in Table 13.2 and are based on actual ground water inflows and mining activities.

Table 13.2 Santoy Mine Dewatering Requirements

Source		Dewatering Requirement
	m3/day		US gallons/min
Ground Water		280		50
Mining Activities		150		28
Total		430		78

13.9 Hydrology Considerations

Water inflow is well understood at the SGO based on actual data and is not expected to change during the LOM. The current dewatering infrastructure system adequately manages water inflows and the system will continue to be expanded as the footprint of the Santoy mine expands.

13.10 Geotechnical Considerations

13.10.1 Rock Mass Quality and Rock Properties

The rock mass at the Santoy mine is generally classified as good with a rock mass rating (Bieniawski, 1976) (RMR76) of 71%–79%. There are some areas classified as fair, with a RMR76 of 52%–57%.

Rock property testing has not been performed at the Santoy mine, but rock property testing performed for the Seabee mine provides analogous results (Table 13.3, Table 13.4, and Table 13.5).

13.10.2 Stress Regime and Most Likely Mode of Failure

Stress monitoring and in situ stress measurements have not been conducted at the Santoy mine. It is assumed, based on typical Precambrian Canadian Shield conditions (Herget 1988) that the horizontal to vertical stress ratio is two and that the major principal stress direction is horizontal and parallel to the strike of the orebody.

The most likely mode of failure at the Santoy mine is either structural or rock mass driven failure. In areas where the RMR is 71%–79%, the dominant mode of failure will be structural. In areas where the RMR is 52%–57%, the dominant mode of failure will be wedge failure. Gravity is the driving force for failure as high stress with seismic activity and rock bursting is not a concern due to the shallow depth of mining.

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Table 13.3 Summary of Testing Results for the Hangingwall Structure at Seabee Mine

Zone / Box	Conversion Factor, K	UCS Point Load<br>(MPa)	UCS<br>(MPa)	Tensile Strength<br>(MPa)	Static E<br>(GPa)	Static v	Dynamic E<br>(GPa)	Dynamic v	Φ	C<br>(MPa)
HW / U11-037	18	100 ± 19 (10)	102 ± 53 (6)		43 ± 9 (2)	0.29 ± 0.06 (2)	56 ± 10 (10)	0.23 ± 0.03 (10)
HW / U11-357	18	109 ± 29 (9)	100 ± 5 (4)	13 ± 2.4 (10)	43 ± 12 (3)	0.19 ± 0.08 (3)	51 ± 3 (4)	0.28 ± 0.01 (4)
Average*	18	105 ± 25	102 ± 41	13 ± 2.4	43 ± 11	0.23 ± 0.09	54 ± 9	0.24 ± 0.03	45°	12

Table 13.4 Summary of Testing Results for the Footwall Structure at Seabee Mine

Zone / Box	Conversion Factor, K	UCS Point Load<br>(MPa)	UCS<br>(MPa)	Tensile Strength (MPa)	Static E<br>(GPa)	Static v	Dynamic E (GPa)	Dynamic v	Φ	C<br>(MPa)
HW / U11-037	22	169 ± 45 (10)	171 ± 48 (6)		75 ± 12 (2)	0.20 ± 0.05 (2)	62 ± 7 (10)	0.17 ± 0.06 (10)
HW / U11-357	11	90 ± 22 (11)	90 ± 35 (4)	12.8 ± 2.6 (10)	78 ± 21 (4)	0.21 ± 0.11 (4)	76 ± 15 (4)	0.22 ± 0.05 (4)
Average*	17	138 ± 35	139 ± 59	12.8 ± 2.6	77 ± 18	0.21 ± 0.09	67 ± 11	0.18 ± 0.06	48°	30

* Averages are calculated from all test results

Table 13.5 Summary of Testing Results for the Orezone Structure at Seabee Mine

Zone / Box	Conversion Factor, K	UCS Point Load<br>(MPa)	UCS<br>(MPa)	Tensile Strength<br>(MPa)	Static E<br>(GPa)	Static v	Dynamic E<br>(GPa)	Dynamic v	Φ	C<br>(MPa)
OZ / U11-037				17.6 ± 2.0 (10)
OZ / U11-357	8	71 ± 26 (20)	70 ± 18 (10)		35 E 18 (5)	0.17 ± 0.05 (5)	55 ± 9 (14)	0.21 ± 0.04 (14)	57°	23

(x) Number of tests completed

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13.10.3 Specific Geotechnical Risk

The geotechnical risks at the Santoy mine are structural and rock mass driven failure. Based on geotechnical underground mapping, the Santoy mine has three primary joint sets that contribute to potential structural failure (Figure 13.4):

•JS1 – 58°/358°

•JS2 – 80°/267°

•JS3 – 13°/195°

In the areas where the rock mass is fair the failure mode will likely be wedge failure due to gravity in sills that are greater than 8 m in width. In areas of with a good rock mass or under 8 m span, the failure mode will likely be structural.

13.10.4 Current Mitigation Measures Used to Minimise the Geotechnical Risk Support System

At the Santoy mine there are currently several ground support systems in place that are selected depending on the width of the excavation and its application:

•Inclines / declines: 2.4 m threaded-both-end mechanical rock bolts on a 1.2 m x 1.2 m pattern in the back, 10 mm x 10 mm 6-gauge screen on the back and walls, and 1.8 m split sets on 1.2 m x 1.2 m pattern in the walls.

•Intersections 6–9 m wide: 2.4 m #6 rebar on a 1.2 m x 1.2 m pattern in the back, 10 mm x 10 mm 6-gauge screen on the back and walls, and 1.8 m split sets on 1.2 m x 1.2 m pattern in the walls.

•Sills less than 6 m wide: 1.8 m threaded-both-end mechanical rock bolts on a 1.2 m x 1.2 m pattern in the back, 10 mm x 10 mm 6-gauge screen on the back and walls, and 1.8 m split sets on 1.2 m x 1.2 m pattern in the walls.

•Sills 6–7 m wide: 2.4 m threaded-both-end mechanical rock bolts on a 1.2 m x 1.2 m pattern in the back, 10 mm x 10 mm 6-gauge screen on the back and walls, and 1.8 m split sets on 1.2 m x 1.2 m pattern in the walls.

•Sills greater than 7 m wide: designed cable bolt plan that is dependent on site investigation.

13.10.4.1 Barrier Pillar

Currently the only barrier pillar at Santoy is the 37L to 38L CRF pillar. Barrier pillars are located at the bottom of the first stoping block, and in the future, there will be uphole stoping directly underneath the CRF pillar. CRF barrier pillars are composed of 5% binder mixed with development waste rock.

13.10.4.2 Extraction Sequence

The extraction sequence includes stopes being extracted from the bottom of the stoping block to the top of the stoping block, and from the extremities of the level towards the access near the centre.

13.10.4.3 Backfill

At the Santoy mine, two types of backfill are used: unconsolidated RF, and CRF. CRF is used when mining occurs directly adjacent to the backfill stopes. All other stopes that are backfilled, are filled with unconsolidated waste rock from development elsewhere in the mine.

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13.10.5 Geotechnical Reports Review

A review of previously completed geotechnical studies for SGO has been undertaken. The purpose of the review is to confirm the studies that have been completed to date are appropriate and identify any gaps or areas of residual concern. The following six geotechnical reports were reviewed:

•2015 – Stantec Study

•2017 – Pakalnis & Associates Study

•2017 – SRK NI 43-101 Technical Report

•2018 – Northern Rock Mining Solutions

•2020 – Northern Rock Mining Solutions

•2021 – SSR Internal Report

•2021 – SSR Ground Control Management Plan (GCMP)

Stantec Geomechanical Overview of Stope and Pillar Stability – Santoy Orebody (March 2015)

The Stantec 2015 Study provided a review of stope and pillar design for the Santoy orebody. Designs are based on rock mechanics utilising empirical design methods using available core logging and very limited intact strength testing to define geotechnical rock mass ratings (Q’ and RMR89). Q’ was assessed as ranging from 10–33 (‘Good’ rock) and RMR89 ranging from 51–77 (‘Fair’ to ‘Good’ rock).

Mapping from development within the area indicated three defect sets: foliation/bedding dipping moderately to the east of south (54°/160°), dominant joint set dipping steeply to the northeast (74°/059°) and a minor joint set dipping moderately to the west (46°/286°).

The Mathews Potvin Stability Graph Method was utilised to assess stope dimensions and indicating hangingwall dimensions not to exceed 25 m down dip and open strike length of 20 m. Mathews Potvin approach to stope design requires appropriate understanding of the structural pattern and at Seabee the appropriate dip of foliation / bedding is critical.

MAP3d numerical analyses was utilised to assess stability of interstitial pillars (i.e., pillars left between parallel ore zones) and indicating backfill of stope is required to mitigate interstitial pillar failure coupled with bottom up sequence and lag mining in hangingwall by one stope.

No support designs were provided for development drives, albeit there was comment that for drifts wider than 5 m, longer support such as cables may be required.

Pakalnis & Associates Report on Site Visit – Santoy Mine (June 2017)

Pakalnis & Associates undertook a site visit in June 2017 and reviewed aspects of development support for the Santoy mine. The key driver in engaging the review was two falls of ground (FOG) on 34L, one in January 2017 and the other in April 2017. Key findings of the report were that the standard support using either 1.8 m or 2.4 m long anchors on a 1.2 m x 1.2 m pattern was not adequate where wide spans were utilised in areas where rock mass quality (RMR) is reduced with the presence of “south dipping 30° structure”. Indications that “unstable unless supported by dead-weight requirements”. The “south dipping 30° structure” noted in both FOG are not present in the structural pattern noted in the Stantec 2015 Study.

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SRK NI 43-101 Technical Report for the Seabee Gold Operation (October 2017)

The SRK 2017 report was a NI 43-101 Technical Report for the Seabee operation and the review focused on Section 16.10 of SRK 2017 report regarding Geotechnical Considerations. Key findings of the SRK Study comprise:

•RMR76 typically ‘Good’ ranging from 71–79 with some areas classed as Fair ranging from 52–57.

•Structural pattern noted as: foliation/bedding dipping moderately to the north (58°/358°), dominant joint set dipping steeply to the west (80°/267°) and a minor joint set shallow dipping to the south (13°/195°).

•In areas of ‘Fair’ rock mass quality and for spans greater than 8 m failure mode is likely to be wedges. In areas of ‘Good’ rock mass quality or under 8 m span the failure mode will likely be structural.

Observations and comments of the review are as follows:

•The difference between RMR76 and RMR89 relates to slightly different ratings in the Bieniawski system over time and typically the latter provides values 5 points higher.

•The structural pattern noted by SRK is distinctly different to that noted in the Stantec 2015 Study but largely identical to that later presented in 2021 SSR Ground Control Management Plan. Two key aspects are indicated. Firstly, stope design needs appropriate consideration of local foliation/bedding dips. Secondly, the SRK structural pattern includes the flat south dipping structures, which was a critical issue in the FOG’s noted in the Pakalnis & Associates 2017 Study.

•The SRK Study does not provide an update on stope design parameters or ground support in development and simply states the support types in use for the latter.

NRMS Ramp Inspection and Related Mining Geotechnics (November 2018)

North Rock Mining Solutions (NRMS) provides a review of the Santoy operation. Key aspects of note in the review include:

•Rock mass quality ratings in keeping with the Stantec 2015 Study.

•NRMS indicates support in the declines somewhat conservative.

•Wider spans (>8 m to 10 m) have experienced brow and back failures although a conservative design is suggested based on the rock mass rating and tight cable spacing utilised (nominally 1.5 x 1.0 to 0.75 m).

NRMS provides several potential causes for the failures in what would be expected to be conservative designs. The two most significant issues raised by NRMS to explain the failures include: large discrete structures (which require dead-weight consideration in support design) and undercutting of the hangingwall though inappropriate consideration in the design (this would comprise the hangingwall design being steeper than the local geological bedding / foliation).

NRMS Q4 Mining-Geotechnical Site Visit Summary Notes (September 2020)

In 2020, Northern Rock Mining Solutions (NRMS) provided a further review of the Santoy operation. Part of this review focused on the GHW zone. It was noted that the variability in the bedding/foliation, whilst typically dipping east of north, shows local areas with dip direction to the west and with dips ranging from 35º to 85°. NRMS state: “excavation stability during mining operations largely governed by the presence, detection, and subsequent handling of stope-scale geologic ‘contact’ structures located in the HW, and less frequently within the FW”.

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SSR Ground Control Analysis (June 2021)

In 2021 SGO provided a report on the GHW deposit focused on stope stability and ground control. The report was based on a combination of site-specific drilling, mapping and laboratory testing. Key aspects of report include:

•Presented mapping data: foliation / bedding dipping steeply to the north (74°/004°), dominant joint set dipping steeply to the south-east (78°/134°) and a minor joint set near horizontal. SGO notes the structural pattern has a significant role in determination of rock mass characterisation.

•Logging data indicates geotechnical rock mass ratings of Q ranging from 10–40 (‘Good’ rock) and RMR ranging from 80–85 (‘Very Good’ rock).

•Implications of these rock mass ratings indicating wide spans (up to 12 m) before systematic bolting required and with potential to increase strike length of stopes of up to 30 m.

•The study looked at the option of using transverse stopes and with uncemented backfill of stopes to limit failure of proposed narrow transverse pillars.

Of note in the study is the approach to ground control support at Santoy since the 2017 FOG incidents and rationale as to why potential wider spans can be considered with the current systematic support of 2.4 m long rebar on a 1.5 m x1.5 m dice pattern.

SSR notes: “QA/QC programme to ensure frequent and continuous evaluation of the rock mass and joint sets is available to show that the large wedges predicted by the deadweight analysis cannot form”.

SSR Ground Control Management Plan (2021)

The SSR 2021 Ground Control Management Plan (GCMP) is considered comprehensive and provides an appropriate overview of the support requirements and stope designs at Seabee and appropriate components of a management plan of a principal hazard.

Geotechnical Studies Review Comments

The following comments represent perceived gaps:

•Whilst latter reporting from NRMS and internal SSR reporting indicate potential for wider spans and less support. While the data supports this being feasible in principle, this is not advisable at a practicable level without appropriate mapping, rock mass evaluation, and design assessment.

•The structural pattern / data is a key driver in the stope designs and support in wider spans as indicated in the Pakalnis & Associates 2017 Study. Whilst the SRK 2017 Study and SSR 2021 GCMP indicate an identical structural pattern, the Stantec 2015 Study and the SGO / SSR 2021 Study indicate distinctly different structural patterns and highlight potential for local variation.

•There is concern for the QA/QC being implemented as it is required to confirm viability of wider spans and less support suggested in the latter geotechnical reporting. The structural data shown in Figure 13.4, for the SRK 2017 Study (320 data points) compared to that presented in Figure 13.5, for the SSR 2021 GCMP (322 data points) is largely identical. This suggests only 12 additional mapping points have been collected in three years, whilst this may not be the case, it somewhat confirms the concern.

•The SSR 2021 GCMP comments on use of a cemented rockfill sill pillar as a solid barrier between mining fronts. However, there is no guidance on the vertical separation between mining fronts in any of the above technical reports that have been reviewed.

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Figure 13.4 2017 Structural Data

SRK, 2017

Figure 13.5 2021 Structural Data

SSR, 2021

13.11 Mine Schedule

The Mineral Reserve life of mine extends to the first quarter of 2029 at a production rate of 425 ktpa. The Mineral Reserve production profile tonnage and recovered gold ounces are shown in Figure 13.6 and Figure 13.7.

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Figure 13.6 Production Plan Tonnage

OreWin, 2021

Figure 13.7 Processing Schedule

OreWin, 2021

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Figure 13.8 Production Plan Recovered Gold Ounces

OreWin, 2021

The underground development requirements to realise the LOM production plan are summarised in Table 13.6. Annual waste rock generation and backfill requirements are also included in Table 13.6.

Table 13.6 Development, Waste Rock, and Backfill Summary

Item	Unit	2022	2023	2024	2025	2026	2027	2028	Total
Development
Capital Lateral	m	4,176	3,673	3,431	2,308	1,672	469	–	15,729
Operating Lateral	m	2,508	2,794	1,903	1,761	756	148	–	9,871
Capital Alimak	m	51	–	–	–	57	–	–	109
Operating Alimak	m	186	116	151	40	379	106	–	978
Total	m	6,922	6,583	5,485	4,108	2,865	724	–	26,686
Waste Rock Generated	kt	506	389	493	393	258	109	–	2,149
Backfill Requirement	kt	1,564	278	308	269	305	326	78	1,564

Sum of individual values may not match total due to rounding

13.12 Mobile Equipment

The existing equipment fleet will fulfil the peak requirements of the schedule, capital allowances have been made for the rebuild or replacement of equipment as required over the LOM.

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14 PROCESSING AND RECOVERY METHODS

14.1 General

The Seabee deposit was processed for 25 years in the mill constructed immediately adjacent to the Seabee shaft.

The remote location of the mine in northern Saskatchewan is sustained by air transport for the workforce and winter road access for supplies. The operation was initially developed and operated on diesel power and later connected to Saskatchewan grid power in 1992. The initial capacity was 500 tpd, which was later expanded to 1,000 tpd with the addition of a third grinding mill. The mill flow sheet as shown in Figure 14.1 is a conventional crushing and grinding circuit employing gravity gold recovery and cyanide leaching with carbon-in-pulp (CIP) for recovery and production of doré gold on site.

Table 14.1 shows the main operating statistics for the Seabee mill over the last ten years, which was the main reference in planning the future operations on other deposits in the area as well as the Santoy deposit.

The mill maintains a high availability and routinely averages more than 94% operating time with the average monthly rate from 2014 to the present being 94.6%. Currently, an addition to the gravity recovery circuit is being installed that will increase the gravity gold recovery and reduce the limitations of the main cyanide leach circuit.

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Table 14.1 Seabee Mill Production Statistics 2006–2021

Item	Unit	2006	2007	2008	2009	2010	2011	2012	2013
Production	t	246,000	227,700	228,400	247,641	203,958	257,181	275,235	280,054
Daily Rate	tpd	674	624	626	678	559	705	754	767
Mill Head Grade	g/t	6.16	6.35	6.46	6.17	7.55	5.68	5.86	5.11
Recovery	%	93.6	95.4	95.8	95.3	95.5	95.3	95.6	95.3
Gold Produced	oz	46,300	44,323	45,466	46,827	47,270	44,750	44,756	43,850
Item	Unit	2014	2015	2016	2017	2018	2019	2020	2021
Production	t	279,597	277,386	312,679	330,415	352,000	344,040	255,172	382,478
Daily Rate	tpd	756	760	857	967	1,125	1,087	1,163	1,180
Mill Head Grade	g/t	7.32	8.82	7.91	8.25	9.16	9.56	10.10	9.92
Recovery	%	95.7	96.3	96.6	97.4	97.4	98.2	98.4	98.4
Gold Produced	oz	62,984	75,748	80,351	85,395	100,953	110,864	81,540	120,030

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Figure 14.1 Seabee Mill Flow Sheet

SSR, 2021

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14.2 Crushing

The run of mine ore is crushed at the mil. The circuit consists of primary crushing with a jaw crusher followed by secondary crushing with a cone crusher in closed circuit with a triple deck screen. The product from the crushing circuit, at minus 20 mm, is conveyed to the ore storage bin, which has a live capacity of 400 t. To increase the storage capacity between the crusher and the grinding circuit, and allow for crusher breakdowns or scheduled maintenance, fine crushed ore is stockpile and fed into the circuit through the original crushing feed point.

14.3 Grinding

The grinding circuit consists of a ball mill 2.9 m in diameter and 3.7 m long serving as the primary grinding mill. Two secondary mills, 2.7 m in diameter and 2.6 m long, complete the grinding to 80% passing 200 mesh. The grinding mills operate in closed circuit with hydrocyclones. The ground product is thickened to 48% solids in a 12 m thickener prior to entering the leach circuit. Lime is introduced to the grinding circuit to maintain the pH and free cyanide levels for optimum leach conditions.

14.4 Gravity Recovery

A portion of the cyclone underflow on the primary grinding mill is directed to a Knelson concentrator for further concentration on a vibrating table. The gravity concentrate, averaging approximately 65% of the total gold recovered, is refined with the gold recovered in the hydrometallurgical circuit.

The gravity recovery circuit consists of two Knelson concentrators and an Acacia reactor, which recovers the gravity gold in a separate intense cyanide leach and electrowinning circuit. Installation of this equipment was complete later in 2017. Further optimisation of the Acacia circuit is being undertaken to further improve gravity gold recovery.

14.5 Cyanide Leaching

The leach circuit consists of five agitated leach tanks: one of which is 14.6 m in diameter and 14.6 m in height, and four of which are 8.8 m in diameter and 8.8 m in height. Air injection is maintained in all tanks as well as cyanide addition to the initial tank to maintain the free cyanide level to complete gold dissolution. At the nominal mill capacity, the circuit provides 39 hours of leach time.

14.6 Carbon-in-Pulp

The carbon absorption circuit consists of eight tanks that are 3.4 m in diameter and 4.6 m in height equipped with launder screens to maintain the activated carbon captive in the tanks. The carbon circuit typically has about 17.2 t of activated carbon distributed in the tanks. The CIP tankage provides about three hours of retention time with the gold loaded carbon routinely advanced to the strip circuit.

14.7 Carbon Elution and Electrowinning

The loaded carbon is stripped at atmospheric pressure with a heated solution of caustic and iso-propyl alcohol over an average of three days. Gold is collected on stainless steel cathodes in a single electrowinning cell.

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14.8 Gold Refining

The gold recovered by electrowinning from the CIP circuit and the gold recovered by gravity is periodically refined in a gas-fired furnace and poured in doré gold bars on site.

14.9 Carbon Regeneration

To maintain the activity level of the carbon inventory, the Seabee mill has a carbon regeneration process. Prior to elution, the carbon is washed in hydrochloric acid for removal of calcium and other acid soluble impurities. Following elution, the carbon is subjected to heat treatment and attrition in a rotary kiln and screened to remove fines prior to recycle to the CIP circuit.

14.10 Mill Tailings

All tailings solutions in excess of the mill recycle water that are released to the environment are treated with cyanide destruction to maintain the water quality below release quality standards.

The mill operates primarily on recycled water with 96% of the mill water requirements recycled within the grinding circuit and from reclaim water from the tailings management area.

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15 INFRASTRUCTURE

The major infrastructure at the Seabee mine is shown in Figure 15.1, Figure 15.2, and Figure 15.3 and includes:

•Roads and airstrip.

•Mill buildings and related services facilities including maintenance and truck shops, assay lab, crushing plant, shops and storage buildings, and miscellaneous infrastructure.

•Shaft and headframe.

•Portal.

•Ventilation raises.

•2B mine water management ponds.

•Administrative buildings.

•Water supply and distribution.

•Waste management.

•Fuel storage.

•On-site explosive storage.

•Powerhouse and electrical distribution system.

•Ore stockpile.

•TMFs and water management.

•East Lake water treatment plant.

•Camp accommodation.

•Winter road portages.

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Figure 15.1 Seabee Gold Operation Major Infrastructure

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Figure 15.2 Seabee Gold Operation Mill Site Infrastructure

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Figure 15.3 Seabee Gold Operation Tailings Management Facility Infrastructure

The major infrastructure at the Santoy mine is shown in Figure 15.4 and includes:

•Roads

•Administrative and shop buildings

•Powerhouse and electrical distribution system

•Portal

•Vent raises

•Ore stockpiles

•Waste rock pile

•Settling ponds

•Water treatment plant

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Figure 15.4 Santoy Mine Major Infrastructure

15.1 Site Access Roads

As previously stated, the site can be accessed by an 8 km winter road, which begins at Highway 102 near the community of Brabant Lake, Saskatchewan and consists of 12 portages spanning 11 lakes. The majority of annual supplies are transported to site via the winter road, typically throughout the months of February and March, and until mid-April depending on ice quality.

The two mines are connected via a 14 km haul road, called the Santoy Road. This access road is a one-way road that is operated using radio callouts every 1 km and has specific travel convoy times throughout the day. There are also several miscellaneous roads throughout both the Seabee mine and Santoy mine sites that provide access to infrastructure.

15.2 Product Loadout

The product from the processing facility (doré bars) is transported by air to a third-party refinery.

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15.3 Utilities

The current camp facilities at the Seabee mine and Santoy mine can accommodate up to 251 employees.

15.3.1 Water

Potable water is obtainable locally through SSR’s potable water system at both the Seabee and Santoy mine sites. The site currently uses a slow sand filter system. To better meet the current and future site water needs, a new ultrafiltration potable water system has been installed on site and will be commissioned in Spring of 2022.

15.3.2 Sewage Disposal

At the Seabee mine, sewage is treated in the mill and discharged with the tailings to either the East Lake tailings management facility (TMF) or Triangle Lake TMF.

The septic system at the Santoy mine is a mound system, which is pumped every second day by a vacuum truck to prevent leakage from the system.

15.3.3 Power

Electrical power is provided by a transmission line to the mine by the provincial power authority, Saskatchewan Power Corporation. The mine is connected to a 138 KV hydroelectric power line from Island Falls.

The total power usage for SGO is approximately 8.9 MV amperes and the electrical distribution system has an installed capacity of 10.0 MV.

15.3.4 Fuel Storage

Fuel farms and propane tanks are located at both the Seabee mine and Santoy mine sites.

15.3.5 Explosives Storage

A magazine and an explosives storage area are located at the Santoy 7 deposit servicing the Santoy mine, with a secondary magazine and explosive storage area used previously for the old Seabee mine site situated just off the Porky access road, approximately 1.3 km north-east from the Seabee mill area. Both of these areas have been designed and prepared in accordance with the Mines Regulations (The Mines Regulations 2018, Saskatchewan Employment Act).

15.4 Tailings Management Facilities

There are currently two tailings management facilities (TMF) that are being utilised by the Seabee mill: the East Lake TMF and the Triangle Lake TMF, as shown in Figure 15.3. Tailings deposition alternates between the two TMFs where winter deposition occurs in the Triangle Lake TMF and summer deposition is in the East Lake TMF. The current remaining storage capacities of both TMFs, based on an average production rate at 1,200 tpd, will potentially be reached in late-2030.

Maximum capacities also allow that 200,000 m3 of water are treated and discharged from the tailings management facilities each year. To ensure the treatment volumes are attained, a new water treatment plant at East Lake TMF was constructed in 2017.

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Work is currently underway investigating options for extending the life of the TMFs to accommodate any further extensions of the SGO life.

15.4.1 East Lake Tailings Management Facility

East Lake was a natural lake that was converted to a TMF when the Seabee mine was initially developed in 1991. East Lake was partially dewatered prior to tailings deposition, which provided containment for the first six years of operation. Subsequently, vertical concrete dams lined with high density polyethylene (HDPE) were constructed along the topographic lows along the east and south flanks of the TMFto provide additional storage capacity up to mid-2004. At this time, tailings deposition was relocated to the newly constructed Triangle Lake TMF. To accommodate an increased mine life, further expansion of the East Lake TMF was implemented in 2015. The expansion consists of a 6 m high expansion dike that is comprised of waste rock. Stage 1 construction of the expansion dike (Crest elevation 463 m) was completed in 2016 and additional raises have lifted the dike to its current elevation of 465 m.

The existing tailings line is a 6” diameter HDPE pipe that is approximately 2 km in length and stretches from the mill to the East Lake TMF. Spigot locations at the TMFvary over time.

Supernatant water during tailings deposition in the East Lake TMF is regulated by a pump station situated at the north-east corner of the facility. The pond level is maintained below the maximum operating level by pumping and discharging supernatant to either the Back Pond or to the Triangle Lake TMF. There are also three freshwater diversion pumps situated along the western flank of the East Lake TMF that capture and divert water towards Laonil Lake.

15.4.2 Triangle Lake Tailings Management Facility

Similar to the East Lake TMF, the Triangle Lake TMF was a natural lake that was converted to a TMF. To provide initial containment, a North dam was constructed along the northern shoreline of the TMF and tailings deposition commenced in 2004. In 2007, the north dam was raised and the south dam was constructed along the southern shoreline of the TMF. Both dams were vertical concrete structures lined with HDPE.

As part of the combined East Lake TMF and Triangle Lake TMF expansion to accommodate an increased mine life, the design of the Triangle Lake TMF was modified so that both structures would be raised with mine rock and lined with non-woven geotextile and HDPE liner. The expansion of the TMF was staged, which also included construction of two saddle dikes: saddle dikes W2 and W2A, situated east of the North dam. The design of the saddle dikes was consistent with the raise to the North dam (i.e., rockfill construction with non-woven geotextile and HDPE liner). In the final stage of construction, an emergency spillway was situated at the west abutment of the South dam, accommodating the design storm event for the TMF.

Further wall lifts have been completed on the Triangle Lake TMF and it is currently constructed to its final permitted elevation at 466 m, which will accommodate tailings until late 2030.

Construction of a seepage collection system commenced in the summer of 2014 along the downstream toe of the North dam to collect and manage seepage.

There is a 6″ diameter HDPE pipe that connects to the tailings line at the East Lake TMF and extends approximately 1.2 km to either the North or South dams at the Triangle Lake TMF. Spigot locations at the TMF vary over time.

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Water from the East Lake TMF is immediately discharged to the Triangle Lake TMF and thus the water repository and overall water management is accommodated and regulated at the Triangle Lake TMF. Reclaimed supernatant from the Triangle Lake TMF is discharged into the Back Pond, which serves as a lift station, where supernatant is either pumped to the East Lake water treatment plant for treatment or to the Seabee mill as reclaim. Two freshwater diversion pumps are situated along the eastern flank of the TMF that capture and divert water towards Laonil Lake.

15.4.2.1 Tailings Dam Geotechnical Report Review

A review of previous Geotechnical studies was conducted. The purpose of the review is to confirm the completed studies are appropriate and identify any gaps or areas of residual concern. The following eight reports formed the basis of the review:

•2001 – KHS EIS Study

•2016 – SRK study to evaluate tailings alternatives

•2016 – SRK Annual review

•2017 – SRK Dam Safety Review

•2018 – SRK Design study for Triangle Lake TMF Expansion

•2018 – SRK annual review

•2019 – SRK Dam Break Analysis

•2019 – Newfields independent review

Whilst the above does not comprise the full extent of reporting the above reports represent key information to allow an appropriate geotechnical review to be conducted.

Below is a summary of key points from the geotechnical reports review related to the tailings dam:

•There are seven dams noted in the reporting. Of the seven dams, up to seventeen structures/elements are noted and have been reviewed.

•The extent of ongoing review is comprehensive with Golder having been involved between 2007 and 2013, SRK commencing studies in 2012 and with SRK completing all studies since 2013.

•The SRK studies have involved annual reviews and Dam Safety reviews. Whilst all the historical reporting has not been reviewed, the available SRK reporting since 2016 is comprehensive and involving appropriate rigour and detail.

•The SRK studies are comprehensive and have involved numerous SRK staff at various technical levels to indicate appropriate internal review and maintain a high standard

•A third-party review by Newfields of SRK’s design for Triangle Lake TMF Expansion stated “well done and meets required standards”.

In view of the latter comments in the bullet points above, there are no perceived gaps in the technical studies with regard to the tailings dams at Seabee and with appropriate dam management being practiced.

15.5 Waste Rock Structures

Access roads, the airstrip, dams, dikes, laydown areas, and general site areas were constructed using waste rock, which was characterised as non-acid generating.

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15.6 Rock Quarry

In order to sustain waste rock requirements for construction, SSR developed a rock quarry at the SGO. The location of the quarry is adjacent to the existing Triangle Lake TMF. To date, the main consumption of the waste rock has been for the expansions of both TMFs and for the Santoy road upgrade / maintenance.

15.7 Water Facilities

The Santoy mine has one water management structure, which is the Santoy 8 deposit water management pond. The old Seabee mine has two water management structures: the East Lake water treatment plant and the 2B mine water settling ponds.

15.7.1.1 Santoy Mine

Mine water from the Santoy underground mine is discharged into the north-west corner of the Santoy 8 deposit water management pond where it is then pumped to a water treatment plant. The water is treated by a moving bed bioreactor unit to reduce ammonia concentrations. The treated water is pumped into settling Pond 1 where biomass from the process settles out and from there water flows to settling Pond 2 via an overflow spillway. The water is discharged from settling Pond 2 through a culvert and into the north-east corner of the mine water management pond for final settling. Final discharge to the environment is done via a pump situated at the south end of the mine water management pond. Approximately 100,000 m3 of water from the underground mine is treated and discharged annually.

Settling Ponds 1 and 2 have a perimeter of approximately 105 m and 90 m, respectively, and a maximum height of approximately 3.3 m and 3.8 m, respectively. The ponds are lined with 60 mil HDPE and have a combined total storage volume of approximately 2,250 m3.

The mine water management pond is contained by a main dike situated at the south end of the facility and a north saddle dike located at the north-west flank. Both structures are comprised of waste rock with slopes graded at 2.0H:1V. The upstream slopes are lined with 60 mil HDPE, which are keyed into a low permeable till foundation. The main dike and north saddle dike are approximately 180 m and 120 m in length, respectively and have a maximum height of approximately 7 m and 3.5 m, respectively. The storage volume of the mine water management pond is approximately 40,000 m3.

15.7.1.2 Seabee Mine

The East Lake water treatment plant and associated settling ponds 1 and 2 are used to treat and settle the supernatant water from the East Lake TMF and Triangle Lake TMF. Supernatant is transferred from the Back Pond at the East Lake TMF to the water treatment plant where it is initially treated with lime, ferric sulfate, and peroxide. Subsequently, the treated water is discharged to settling pond 1, which overflows to settling pond 2. From here the treated water is pumped to East Pond where it is monitored prior to the final discharge to the environment. Settling ponds 1 and 2 have a perimeter of approximately 190 m and 100 m, respectively, and a depth of 2.5 m and 6 m, respectively. The ponds are lined with 60 mil HDPE and have a combined storage capacity of approximately 13,000 m3. Approximately 80,000–100,000 m3 are treated and discharged to the environment annually, which correlates to a treatment rate of approximately 835 m3 per day, based on a four-month treatment period.

As previously stated, a water treatment plant was constructed in 2017. The water treatment plant has capacity to treat up to 3,400 m3 per day, removing cyanide, ammonia, and copper from the TMF supernatant. In general, the treatment process consists of a pre-treatment step for removal of copper and cyanide followed by a moving bed bioreactor unit for removal of ammonia.

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16 MARKET STUDIES

16.1 Marketing and Metal Prices

The metal prices used in this Seabee21TRS are based on an SSR internal assessment of recent market prices, long-term forward curve prices, and consensus amongst analysts regarding price estimates. The metal prices selected for the Seabee Gold Operation (SGO), have taken into account the current project life. For the economic analysis in the Seabee21TRS, the metal prices shown in Table 16.1 were used.

Table 16.1 Seabee21TRS Economic Analysis Gold Price Assumptions

Commodity	Unit	2022	2023	2024	2025	Long- Term
Gold	$/oz	1,800	1,740	1,710	1,670	1,600

SGO currently produces doré bars. The doré refining terms are typical and consistent with standard industry practices and similar to contracts for the refining of doré elsewhere.

The doré is transported by secure freight to a refinery, refined into gold bullion and sold by SSR to banks that specialise in the purchase and sale of gold bullion.

No external consultants or market studies were directly relied on to assist with the sales terms and commodity price projections used in the Seabee21TRS. The QP for this Section 16 agrees with the assumptions and projections presented.

16.2 Contracts

There are a number of acceptable refineries with capacity to refine doré. Currently, SSR is in a non-exclusive contractual relationship with Asahi Refining Canada Ltd. (Asahi). The terms of this contract with Asahi are within industry norms. The cost for transport and refining of the doré is in accordance with industry standards.

16.3 QP Opinion

Data and assumptions for macroeconomic trends, taxes, royalties, interest rates, marketing information and plans, legal matters such as statutory and regulatory interpretations affecting the mine plan, and environmental matters are outside the expertise of the QP and are within the control of the registrant (see Section 25).

The Seabee21TRS QP considers it reasonable to rely on SSR because SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

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17 ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS

17.1 Regulatory Setting

The environmental assessment and permitting framework for mining in Canada is well established. Proposed projects are screened both federally and provincially to determine whether an environmental assessment under federal, provincial, or both levels of legislation is required. Following the assessment decision, the project advances to a licensing and permitting phase.

In the event the project’s environmental assessment is successful and all necessary licences and permits are granted, the project is then regulated through all phases (construction, operation, closure, and post closure) by both federal and provincial departments and agencies.

17.2 Federal Environmental Assessment Process

In the spring of 2012, the Canadian Environmental Assessment Act (1992) was amended and replaced by the Canadian Environmental Assessment Act (2012) (CEAA 2012). Two significant results of this amendment were the re-definition of what triggers a federal environmental assessment and the introduction of legislated time periods within a federal environmental assessment, if required.

Under CEAA 2012, an environmental assessment focuses on potential adverse environmental effects that are within federal jurisdiction including:

•Fish and fish habitat.

•Other aquatic species.

•Migratory birds.

•Federal lands.

•Effects that cross provincial or international boundaries.

•Effects that impact on aboriginal peoples, such as their use of lands and resources for traditional purposes.

•Changes to the environment that are directly linked to or necessarily incidental to any federal decisions about a project.

Under the CEAA 2012, there are two main scenarios in which a federal environmental assessment could be required:

1. A proposed project will require an environmental assessment if the project is described in the Regulations Designating Physical Activities, CEAA 2012.

2. Section 14(2) of CEAA 2012 allows the federal Minister of Environment to (by order) designate a physical activity that is not prescribed by regulation if, in the Minister’s opinion, either the carrying out of that physical activity may cause adverse environmental effects or public concerns related to those effects may warrant the designation.

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17.3 Provincial Environmental Assessment Process

The provincial environmental assessment process begins with the submission of a technical proposal to the Saskatchewan Environmental Assessment Branch (EAB) of the Ministry of Environment (MOE) to determine if the project is considered a ‘development’. The MOE will coordinate an inter-ministry review of the technical proposal and the environmental impact statement using a standing panel of representatives from provincial departments and agencies, which is known as the Saskatchewan Environmental Assessment Review Panel (SEARP).

The Saskatchewan Environmental Assessment Act (SEAA 2013) states:

A ‘development’ means any project, operation or activity or any alteration or expansion of any project, operation, or activity, which is likely to:

•have an effect on any unique, rare, or endangered feature of the environment;

•substantially utilise any provincial resource and in so doing pre-empt the use, or potential use, of that resource for any other purpose;

•cause the emission of any pollutants or create by-products, residual or waste products which require handling and disposal in a manner that is not regulated by another act or regulation;

•cause widespread public concern because of potential environmental changes;

•involve a new technology that is concerned with resource utilisation and that may induce significant environmental change; or

•have a significant effect on the environment or necessitate a further development which is likely to have a significant effect on the environment.

17.4 SGO Environmental Assessments

The SGO has been in production since 1991. As part of the initial environmental assessment, approvals and the subsequent expansions at the operation, the existing environment was characterised in three environmental assessments, in accordance with the SEAA 2013. The initial environmental assessment focused on the original Seabee mine and mill and was completed in 1990 (Beak, 1990). The second environmental assessment was necessary to assess the potential environmental impacts associated with the construction and operation of the Triangle Lake TMF and was completed in 2001 (KHS, 2001). The third environmental assessment was necessary to assess the potential environmental impacts associated with the development of the Santoy mine and was completed in 2009 (Golder 2009). For each of these assessments, baseline data was collected, and the potential environmental impacts associated with the proposed project were assessed. In all three environmental assessments, no significant potential environmental impacts were identified that could not be mitigated through the implementation of management plans. Subsequently, Ministerial Approvals to proceed to construction and operation were granted for each of the three environmental assessments.

The Triangle Lake TMF, as well as the Santoy mine projects, were screened by the Canadian Environmental Assessment Agency in 2001 and 2009, respectively. The SGO has never required a federal environmental assessment.

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17.5 Environmental Permits/Authorisations

Following a successful environmental assessment, the SGO is required to obtain a number of federal and provincial permits / approvals / licences. These permits outline the environmental operating specifications and reporting requirements of the operation. Although all regulatory permits and approvals carry the same level of importance, the Provincial Approval to Operate (PO) is the primary regulatory approval required to operate a gold mine in Saskatchewan. The PO is issued in accordance with numerous provincial legislation and regulations governing Saskatchewan’s mining industry.

Since its inception, the SGO has operated under the terms and conditions of a PO, issued by the MOE. As discussed in Section 3.6, the operation’s current Approval to Operate number PO19-193, was issued in October 2019 and expires in September 2022. This approval outlines all monitoring and reporting requirements for all operations, including:

•Surface and groundwater in immediate and surrounding areas

•Sediment quality of surrounding lakes

•Aquatic biota in surrounding lakes

•Facilities and areas requiring daily, weekly, and monthly inspections

•Regular acid rock drainage/metal leaching testing

•Annual geotechnical inspection by a Professional Geotechnical Engineer

•Development and regular updates to a variety of management plans

The SGO is in compliance with the terms and conditions of this approval.

17.6 Environmental Considerations

Additional environmental baseline information was gathered to augment the existing environmental baseline database. These completed studies included:

•Heritage Resource Impact Assessment (CanNorth, 2016a)

•Vegetation Inventory Study (CanNorth, 2016b)

•Seabee Mine Quarry Rock ML/ARD Assessment (SRK, 2016a)

Following the completion of the above studies and the integration of those results with the existing baseline database developed for the operation as a result of its three previous environmental assessments, a self-screening of the proposed quarry was completed. The quarry project did not require a formal environmental assessment and the quarry has been established to provide waste rock for TMF expansions and other site projects.

Solid non-hazardous waste generated at the site is disposed of in the approved landfill. In accordance with the SGO’s Approval to Operate, hazardous wastes are stored in approved facilities at the site until the winter, when these materials are transported off site for disposal at approved hazardous waste disposal facilities. In addition, recyclable materials such as scrap metal are stored in segregated piles on an approved lay down area, and later transferred off site as backhaul material on emptied supply trucks via the winter road.

Sodium cyanide, ferric sulfate, lime, hydrogen peroxide, diesel, gasoline, propane, and all other consumables are transported to the site via truck over the winter road, which is generally operational from the end of January through to the end of March each year. All consumables are transported to the site in accordance with the Transport Canada Transportation of Dangerous Goods Regulations and stored in approved bulk storage facilities in accordance with the SGO’s Approval to Operate and Saskatchewan’s Hazardous Substances and Waste Dangerous Goods Regulations.

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SSR has characterised mine rock and tailings for the potential of acid rock drainage/metal leaching at the SGO since 2012. The results of these analyses are reported to the MOE as part of the operation’s annual reporting commitments. Similar programmes will be refined and periodically carried out as operations continue. To date, the findings indicate that the mine rock is non-acid generating. All ores mined at the SGO have a low sulfide content, which is consistent with most vein hosted gold deposits. The current data set shows the Santoy ores carry a lower sulfide content than the ores of the now-ceased Seabee mine. From a geochemical perspective, this means the tailings with the higher sulfur content are located in the lower elevations of the tailings facilities, which are typically saturated or partially saturated. These tailings are then covered stratigraphically by the Santoy tailings through continued operation. The Santoy tailings display the lowest sulfur content (less than 1%) and an equivalent balance of carbonate content, meaning that the residual sulfur content after the carbonate is consumed in the neutralisation process, would not likely support acidic drainage from the upper-most layers of tailings in both facilities. Thus, tailings found in the unsaturated zones of the facilities that will be more readily oxidised are the most geochemically stable tailings. Following 25 years of operation, the site continues to display no evidence of acid drainage.

The geochemical characterisation to date, combined with the tailings operational plan, which ensures that at closure, the unsaturated zone consists of low sulfur bearings tailings, supports the current closure plans for these facilities.

17.7 Mine Closure

In accordance with Saskatchewan’s Mineral Industry Environmental Protection Regulations (1996), the SGO has, since 1996, submitted to the MOE a decommissioning and reclamation plan (closure plan) and cost estimate to implement this plan every five years or when required by the Ministry. In accordance with these regulations and the site’s Approval to Operate, this closure plan is required to be revised and submitted for review and approval at least every five years or as requested by the MOE. The most recent closure plan (SGO Preliminary Decommissioning and Reclamation Plan, 2016 Update) was submitted in January 2017 and accepted by the Government of Saskatchewan in July 2020.

The closure plan meets the following objectives:

•Complies with previous environmental assessment and existing commitments as outlined in the SGO’s Approval to Operate.

•Meets the MOE’s final mine closure objectives as outlined in the Guidelines for Northern Mine Decommissioning and Reclamation (SMOE 2008), specifically:

•Leaves all disturbed areas safe for traditional land uses and in an ecological condition that is consistent with the surrounding physical and biological environment.

•Leaves the site in a state that requires minimal or no maintenance.

•Eliminates potential short and long-term health, safety and environmental risks associated with any aspect of the site.

•Ensures long term physical stability of all landforms and containment structures, in accordance with the Canadian Dam Association Guidelines.

The total estimated cost to implement the closure plan through an independent contractor is approximately C$12M.

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SSR, in accordance with the Mineral Industry Environmental Protection Regulations, is responsible to post financial assurance equalling the closure cost estimate with the Government of Saskatchewan. An update to the closure estimate is currently underway, to cover the approved expanded TMF.

In accordance with the EAB guidance, effluent discharges from the site during the implementation of closure activities will meet Saskatchewan Effluent Quality Limits. Final decommissioning and reclamation water quality objectives for the site, which are determined jointly by the operator and the MOE, will be met at the site prior to the Ministry’s acceptance of the property into its Institutional Control Program. The previously approved closure plan and its current reiteration, which is under final review, assume these final water quality objectives will meet Saskatchewan Environmental Quality Standards for Surface Water guidelines which are more stringent than Saskatchewan’s Effluent Quality Limits.

The total estimated cost to implement the closure plan (under existing site conditions as of December 2020) through an independent contractor is approximately C$12M. The closure cost estimate allows for the full life cycle of mine closure, which includes the following three phases: 1) a decommissioning and reclamation phase to complete the closure activities; 2) a transitional phase to allow for the monitoring of all decommissioning and reclamation activities, ensuring that all closure criteria have been met; and 3) an institutional control phase. Saskatchewan’s Institutional Control Programme requires funds to be set aside for maintenance and monitoring during a 70-year period and requires additional funds to manage the maintenance that may occur as a result of unforeseen events. SSR, in accordance with the Mineral Industry Environmental Protection Regulations, is responsible to post financial assurance equalling the closure cost estimate with the Government of Saskatchewan, covering the three phases of mine closure.

The proposed closure activities for the main components of the SGO, as described in the SGO Preliminary Decommissioning and Reclamation Plan, 2016 Update are summarised below.

17.7.1.1 Mill, Headframe and Supporting Infrastructure

All infrastructures will go through a systematic process of decontamination of potentially hazardous wastes. All assets will be removed and staged on site for transport off site. The remaining structures will be demolished with the use of heavy equipment and recyclable metal will be segregated and stored for transport off site. The soils, if present under and around the foundations, will be characterised for potential contamination of hydrocarbons or metals. If contamination is identified the extent will be delineated and removed for disposal or onsite remediation in accordance with the applicable regulations. All non-recyclable demolition debris will be buried or disposed of in a designated area on site.

17.7.1.2 Tailings Management Facilities and Water Treatment Sludge

Each TMF will be decommissioned and reclaimed using a dry cover, graded towards a spillway, located at the south end of each of the tailings facilities. A 0.3 m cover of erosion-resistant mine rock will be placed on the tailings to form the final cover. This rock cover will mimic the grading of the underlying tailings and eliminate the migration of windblown tailings from the facilities. Dams are constructed of erosion-resistant rock fill and no further closure activities are proposed. The dams are designed and operated in accordance with the guidelines of the Canadian Dam Association, which are reviewed and approved by the MOE. All closure activities associated with the containment structures will comply with the guidelines of the Canadian Dam Association (CDA, 2013).

The plan and costing allows for water treatment to occur until such time as the quality of any remaining ponded water meets site specific water quality objectives.

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Water treatment sludges at the mine are relatively small in volume. Following the decommissioning and reclamation of the water treatment plant, the sludges will be covered in place with a till cover or a combination of a liner / till / sand / mine rock cover.

17.7.1.3 Mine Rock and Ore Stockpiles

No mine rock associated with the SGO is characterised as potentially acid generating, and therefore the closure objective is to ensure long-term physical stability of the piles. The largest single source of mine rock in a central location forms the foundation of the airstrip. All of this material will be used as the construction material for the tailings facility covers. A portion of the remaining mine rock will be used as cover material for the clean demolition debris and backfill material for the existing portals and mine openings, where appropriate. Any remaining mine rock not used as construction material in the decommissioning and reclamation activities, will be contoured to a 3:1 slope and allowed to naturally revegetate.

Prior to the completion of operations, all ore stockpiles will be processed.

17.7.1.4 Contaminated Soils

In the event hydrocarbon contaminated material is identified, the material will be excavated and land farmed in a designated area. Any liquid product produced from the land farm will be transferred into drums and sent offsite for disposal in a licenced facility or used in the waste oil burner.

Due to the low sulfide nature of the orebody, and the clean characterisation of the mine rock, soils containing metals that exceed Canadian Council of Ministers of the Environment Canadian Environmental Quality Guidelines are not expected to be encountered; however, should they be, the material will be hauled to and disposed of within the tailings facility.

17.7.1.5 Non-Hazardous Waste Landfill

The current operating procedures for the landfill call for progressive reclamation. Following placement of refuse, it is covered with mine rock. At closure, all slopes of the covered landfill will be contoured to a minimum of a 3:1 slope.

17.7.1.6 Water Treatment Plants

The Santoy mine will be allowed to flood naturally following operations, and therefore the Santoy water treatment plant will be decommissioned. Its components will be either transported off site as assets or disposed of on site as non-hazardous waste.

The East Lake water treatment plants will remain operational throughout the decommissioning and reclamation activities until such time as further water treatment is not required. Following the need for water treatment, the plants will be dismantled and removed from site.

17.7.1.7 Mines

Following completion of production, all rolling stock will be removed from the underground, stored at the staging area, and prepared for transportation off site for either resale or salvage. The underground workings will be inspected and all hazardous wastes and dangerous goods will be transferred to the surface and ultimately off site for disposal at an approved facility. Following this recovery of assets and decontamination, the mines will be allowed to flood naturally.

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There are 12 vertical to sub-vertical vent raises and one shaft associated with the SGO . Each of these openings will also be fitted with an engineered concrete reinforced cap keyed into bedrock, in accordance with accepted industry practices. The sub-horizontal openings (five portals) will be backfilled with approximately 15 m of waste rock. The waste rock will be extended past the portal entrance and will be contoured to a slope of 3:1.

A final evaluation of all crown pillars will be completed as part of the engineering of the final closure plan. Crown pillars determined to pose a higher risk of failure will be collapsed as part of the decommissioning process. There are currently 17 crown pillars that do not pose a long-term risk of failure, and six crown pillars, which may require collapse and backfilling as part of the decommissioning and reclamation activities.

17.7.1.8 Miscellaneous Infrastructure

All roads, parking areas, lay down areas, settling ponds, winter road portages, and footprint of the air strip will be scarified to support revegetation following the removal of all culverts, power lines, pipelines, and other miscellaneous infrastructure. This infrastructure will be disposed of as part of the major infrastructure decommissioning and reclamation plan.

17.7.1.9 Revegetation

The site will be revegetated in accordance with MOE’s Guidelines for Northern Mine Decommissioning and Reclamation through a combination of natural and active revegetation.

17.8 Social and Community Impact

The SGO is within the Treaty 10 area and borders the Pelican Narrows and Brabant Lake community areas of influence (SMOE 2003). These communities were consulted during the completion of previous environmental assessments in support of the project throughout its operating history. The socio-economic study area for the Santoy mine environmental impact statement (the most recent environmental assessment completed in 2009) included La Ronge, Air Ronge, Kitsakie IR 156B, Lac La Ronge IR 156, Nemeiben River IR 156C, Stanley Mission IR 157, Grandmother’s Bay IR 219, Brabant Lake, Pelican Narrows IR 184B, Pelican Narrows, Sandy Lake, Southend IR 200, and Deschambault Lake IR 203.

In accordance with the terms and conditions of the operation’s Surface Lease Agreement, continual effort has been made at the SGO to engage the nearby communities in order to maximise northern employment opportunities as well as the local purchase of goods and services to support the mine. As of the end of 2021, approximately 19% of the nearly 360 employees at the SGO are northern Saskatchewan residents. The operation continues to honour its social commitments outlined in the project’s surface lease agreement.

Since SSR’s purchase of the SGO, a concerted effort has been made to maintain and strengthen the relationship with the surrounding communities, including the Lac La Ronge Indian Band and the Peter Ballantyne Cree Nation.

In addition, stakeholder engagement plans have been developed to support the proposed quarry. Engagement activities defined in these plans are currently underway.

17.9 Safety

The management of safety and health at the SGO reflects the effective management of risk. The mine’s safety and health strategy is two-fold: to ensure full compliance with the Saskatchewan Mine Act regulations; and to minimise residual risk in relation to regulatory compliance through a risk-centred safety and health management system.

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SGO is committed to continuous improvement in all functions and especially in Safety Health and Environment. In 2021, SGO more than halved its TRIFR (Total Recordable Injury Frequency Rate – per million work hours) from 23.0 in 2020 to 10.8 in 2021. SGO has set a target to halve the TRIFR again in 2022.

Mining-related hazards are inventoried and characterised in terms of their risk, i.e., development of a comprehensive risk registry. Controls, in the form of appropriate engineering and mine design, fixed and mobile equipment optimisation, work processes, training and competency verification, and others, are implemented in relation to risks with proportional emphasis on catastrophic risk. Special emphasis is given to risks such as geotechnical, mine design and operational risk.

In addition to the central risk management framework, the mine employs a wide variety of policies, processes and procedures that populate the safety and health management system including, but not limited to, safety committees, daily workplace audits, safety communication, proper use of protective equipment, job hazard analysis and standard operating equipment, contractor management, a focus on behaviour modification and human error, and incident investigation and root cause analysis, among others.

In instances where changes to risk management practices occur as a result of changes to mine equipment, practices, geotechnical information as well as other change criteria, the mine undertakes a change management review to ensure that those changes do not result in an increase in potential risk. Where change does result in additional risk, relevant control measures are modified.

While the SGO’s approach to risk management is primarily focused on the prevention of incidents, and has substantially reduced safety incidents, the operation also maintains a properly staffed, trained, and provisioned mine rescue team that is prepared to address any foreseeable emergency that might occur underground or on surface. Dedication, and diligent preparation and training have resulted in provincial recognition for the mine’s rescue team and system.

SGO’s safety and health management system, like all effective management systems, undergoes review of continuous improvement involving performance metrics and other training and leading key performance indicators. However, SSR also recognises that the system is only as effective as the organisational culture and the degree to which the system is adopted by its members as common practice. Accordingly, there is also recognition that the behaviour of leaders at the mine has a substantial impact on the mine’s operational culture. As such, the mine emphasises culture assessment and enhancement through leadership development.

17.10 QP Opinion

Following a review of the information supplied, the opinion of the QPs is that it is reasonable to rely on the information provided by SSR as outlined above for use in the Seabee21TRS because a significant environmental and social analysis has been conducted for the project over an extended period, the project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

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18 CAPITAL AND OPERATING COSTS

This section summarises the costs used by the QPs to validate the economics of the Mineral Reserve estimate for the SGO. The cost estimate was prepared by the SSR technical department at both the SGO site and Saskatoon office. The QPs reviewed the assumptions, parameters, and methods used to prepare the cost estimate and is of the opinion that they are sufficient for the purposes of validating the economics of the Mineral Reserves.

The cost estimates were completed in C$ and converted to US$ at an exchange rate of C$1.26:US$1.00.

Cash costs and all-in sustaining costs (AISC) per payable ounce of gold sold are non-GAAP financial measures. Please see “Cautionary Note Regarding Forward-Looking Statements” in this Seabee21TRS.

18.1 Capital Costs

The estimated capital costs required to achieve the Mineral Reserve LOM are summarised in Table 18.1. The capital costs were estimated from historical construction costs and equipment purchase prices, actual development costs, as well as results from study work completed by OreWin and third-party consulting firms. Where costs were not available for some minor components, an experience-based allowance was included.

Table 18.2 represents the categorised capital costs estimated as of the beginning of 2022. As the project has been in operation for a number of years, the level of project definition for the capital cost estimate is very high. Given some new capital projects are planned, a contingency of 10% was included for capital costs outside of mine development from 2023 onwards. The QP considers the capital estimate to be in the accuracy range of +/–15%.The sustaining capital costs include:

•Surface infrastructure construction such as upgrades to the camp and kitchen, IT upgrades, and asset integrity costs.

•Mill improvements and replacement of major components.

•TMF construction costs.

•Mobile equipment such as new and replacement purchases and major rebuilds.

Table 18.1 Capital Costs Estimate

Cost Component	$M
Capital Development	85
Sustaining Capital	71
Capital Cost Before Contingency	156
Contingency	6
Total Capital Cost	162

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18.2 Operating Costs

The operating costs were estimated based on the actual operating expenditures at the SGO in 2021. The costs were estimated by process / activity with fixed and variable components.

The operating expenses estimated to validate the positive cash flow for the Mineral Reserve LOM are summarised in Table 18.2. The mining expense includes all labour, supplies / consumables, and equipment maintenance to complete mining related processes / activities, less exploration diamond drilling and capital excavations and construction. The milling expense includes all labour and supplies / consumables to complete milling related processes / activities. The administrative expense includes all labour, supplies / consumables, and equipment maintenance to complete administrative, finance, human resources, environmental, safety, supply chain, site services, camp and kitchen, and travel related processes / activities.

As the project has been in operation for a number of years, the level of project definition for the operating cost estimates is very high. Given the available project performance data and the high level of project definition, no contingency was included in the operating cost estimate. The QP considers the operating cost estimate to be in the accuracy range of +/–15%.

Table 18.2 LOM Average Operating Costs Estimate

Cost Component	$/t milled
Mining	46
Surface Haulage	6
Milling (incl. Fixed Plant)	35
G&A	68
Total Operating Cost	155

Sum of individual values may not match total due to rounding

The estimated total cash costs for the first two years of production is $538 per payable ounce of gold, with a LOM average of $735. The AISC, which includes infrastructure capital and capital development, is $868 per payable ounce of gold for the first two years of production, with a LOM average of $1,021.

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19 ECONOMIC ANALYSIS

19.1 Economic Assumptions

The modelling and taxation assumptions used in the Seabee21TRS are discussed in detail below.

All monetary figures expressed in this report are in US Dollars ($) unless otherwise stated. The SGO financial model is presented in 2021 constant US dollars, cash flows are assumed to occur evenly during each year and a mid-year discounting approach is taken.

19.1.1 Pricing and Discount Rate Assumptions

The gold prices used for the economic analysis are shown in Table 19.1. Gold provides the only revenue included in the analysis.

Table 19.1 Seabee21TRS Economic Analysis Gold Price Assumptions

Commodity	Unit	2022	2023	2024	2025	Long- Term
Gold	$/oz	1,800	1,740	1,710	1,670	1,600

Other key assumptions in the economic modelling relating to product pricing are tabulated in Table 19.2. A discount rate of 5% is used for calculating net present value (NPV).

Table 19.2 Seabee21TRS Key Economic Assumptions

Model Assumption	Unit	Value
Refinery Charge	$/oz gold	0.45
Gold Payability	%	99.5
Tax Rate	%	25.9

The estimates of cash flows have been prepared on a real basis as 1 January 2022 and a mid-year discounting is used to calculate NPV.

In the analysis, carry balances such as tax and working capital calculations are based on nominal dollars and outputs are then deflated for use in the integrated cash flow calculation.

19.1.2 QP Opinion on Inputs

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19.2 Overview and Results

The projected financial results include:

•After-tax NPV at a 5% real discount rate is $249M

•Mine life of six years

As SGO is an existing operation with a forecast of positive cash flows, internal rate of return, and payback calculations were not required.

There are no credits from metals other than gold included in the cash cost.

The key results of the Seabee21TRS are summarised in Table 19.3.

Table 19.3 Seabee21TRS Results Summary

Description	Unit	Seabee21TRS
Ore Processed
Ore Tonnes Treated	kt	2,684
Au Feed Grade	g/t	6.72
Gold Recovery	%	98.0
Metal Produced
Gold	koz	568
Key Cost Results
Site Operating Costs	$/t milled	155
Mine Site Cash Cost	$/oz payable gold	734
Royalties and Refining	$/oz payable gold	0.5
Total Cash Costs (CC)	$/oz payable gold	735
All-in Sustaining Costs (AISC)	$/oz payable gold	1,021
Average Gold Price	$/oz payable gold	1,701
NPV	$M	249
Discount Rate	%	5
Project Life	years	6

19.2.1 Production and Cost Summary

The process production forecasts are shown in Table 19.4 and forecast ore tonnes mined are shown in Figure 19.1. The processing tonnes and metal production are summarised in Figure 19.2 and Figure 19.3 respectively.

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Table 19.4 Production Statistics

Item	Unit	Total LOM	2-Year Annual Average	LOM Annual Average
Ore Processed
Ore Tonnes Treated	kt	2,684	424	424
Au Feed Grade	g/t	6.72	9.34	6.72
Gold Recovery	%	98.0	98.0	98.0
Metal Produced
Gold	koz	568	125	90

Figure 19.1 Production Plan Tonnage

OreWin, 2021

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Figure 19.2 Processing Schedule

OreWin, 2021

Figure 19.3 Production Plan Recovered Gold Ounces

OreWin, 2021

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19.2.2 Project Financial Analysis

The estimated Mine Site Cash Cost is shown in Table 19.5. These estimated costs include only direct operating costs of the mine site, namely:

•Mining

•Processing

•Tailings

•General and administrative (G&A) costs

•Government fees and charges (excluding corporate taxation)

The projected financial results include:

•After-tax net present value (NPV) at an 5% real discount rate is $249M.

•Mine life of six years.

Table 19.5 Cash Costs

Description	2-Year Average<br>($/oz)	LOM Average<br>($/oz)
Mine Site Cash Cost	538	734
Royalties and Refining Charges	0.5	0.5
Total Cash Costs (CC)	538	735
All-in Sustaining Costs (AISC)	868	1,021

The estimated revenues and operating costs have been presented in Table 19.6, along with the estimated net sales revenue value attributable to each key period of operation.

The gold prices used for the economic analysis are shown in Table 19.1. Gold provides the only revenue included in the analysis.

The gold price used in this Seabee21TRS is based on an SSR internal assessment of recent market prices, long-term forward curve prices, and consensus amongst analysts regarding price estimates. The gold price selected for SGO have taken into account the current project life.

The estimated total Project direct capital costs are shown in Table 19.7.

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Table 19.6 Operating Costs and Revenues

Description	Total<br>($M)	2-Year Average<br>($/t Milled)	LOM Average<br>($/t Milled)
Revenue
Gross Sales Revenue	961	518	358
Less Realisation Costs
Treatment and Refining Charges	0.3	0.1	0.1
Total Realisation Costs	0.3	0.1	0.1
Net Sales Revenue	961	518	358
Less Site Operating Costs
Mining	124	49	46
Surface Haul	15	6	6
Milling (incl. Fixed Plant)	95	35	35
G&A	181	68	68
Total Operating Costs	415	157	155
Operating Margin	546	361	204

Table 19.7 Total Project Capital Cost

Item	Total<br>($M)
Mine Development	85
Sustaining	71
Capital Cost Before Contingency	156
Contingency – 10%	6
Capital Cost After Contingency	162

Capital includes only direct project costs and does not include non-cash shareholder interest, management payments, foreign exchange gains or losses, foreign exchange movements, tax pre-payments, or exploration phase expenditure.

The projected financial results for undiscounted and discounted cash flows at a range of discount rates are shown in Table 19.8.

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Table 19.8 Financial Results

Discount Rate		NPV (M)
	Before-Tax
Undiscounted		372	274
2%		358	263
5%		338	249
10%		309	228
12%		299	221
15%		285	211
18%		273	201
20%		265	196

All values are in US Dollars.

The results of NPV sensitivity analysis to a range of changes in gold price and discount rates are shown in Table 19.9. NPV sensitivity analysis for changes to operating and capital costs are shown in Table 19.10. The estimated cumulative cash flow is depicted in Figure 19.4 and a complete cash flow is provided in Table 19.11.

Table 19.9 After-Tax NPV5% Sensitivity to Gold Price and Discount Rates

Discount Rate		Gold Price(/oz)
	–400			–200		–100		–		+100		+200		+300		+400
Undiscounted		106	148		190		232		274		316		358		400		442
2%		104	144		184		224		263		303		343		383		422
5%		101	138		175		212		249		286		323		360		396
10%		96	129		162		195		228		261		294		327		359
12%		94	126		158		189		221		252		284		315		347

All values are in US Dollars.

Table shows NPV5% $M

Table 19.10 After-Tax NPV5% Sensitivity to Operating and Capital Cost Changes

Item		Changes to Cost<br>(%)
	–30%		–20%		–10%		–5%		–		+5%		+10%		+20%		+30%
Operating Cost		328		302		275		262		249		236		223		196		169
Capital Cost		280		269		259		254		249		244		239		230		221

Table shows NPV5% $M

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Figure 19.4 Cumulative Cash Flow

OreWin, 2021

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Table 19.11 Estimated Cash Flow

Description	Unit	Total	2022	2023	2024	2025	2026	2027	2028	2029	2030
Total Gross Revenue	$M	961.4	218.8	220.4	158.8	115.1	111.6	104.9	31.7	–	–
Total Realisation Costs	$M	0.3	0.1	0.1	0.0	0.0	0.0	0.0	0.0	–	–
Net Revenue	$M	961.1	218.7	220.4	158.8	115.1	111.6	104.8	31.7	–	–
Site Operating Costs
Mining	$M	123.9	20.8	20.8	20.4	19.9	19.3	17.4	5.4	–	–
Milling (incl. Fixed Plant)	$M	109.8	17.3	17.4	17.4	17.4	17.4	17.4	5.7	–	–
G&A	$M	181.3	28.6	28.6	28.6	28.6	28.6	28.6	9.5	–	–
Total Operating Costs	$M	415.0	66.7	66.7	66.3	65.9	65.3	63.4	20.6	–	–
Operating Surplus / (Deficit)	$M	546.2	152.0	153.6	92.5	49.2	46.3	41.4	11.2	–	–
Capital Costs
Mine Development	$M	84.8	25.3	23.8	17.3	7.8	8.2	2.4	–	–	–
Sustaining	$M	83.0	19.8	11.9	11.1	12.6	8.4	6.2	6.8	6.0	–
Contingency	$M	6.1	–	1.1	1.0	1.3	0.8	0.6	0.7	0.6	–
Total Capital	$M	173.8	45.1	36.9	29.4	21.6	17.4	9.3	7.5	6.6	–
Pre-tax Cash flow	$M	372.4	106.9	116.8	63.1	27.6	28.8	32.2	3.6	–6.6	–
Tax Payable	$M	98.3	27.7	30.3	16.3	7.1	7.5	8.3	0.9	–	–
After-tax Cash Flow	$M	274.1	79.2	86.5	46.7	20.4	21.3	23.8	2.7	–6.6	–

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20 ADJACENT PROPERTIES

On 2 December 2021, SSR announced the proposed acquisition of Taiga Gold Corp. for $21M (“Taiga”) for C$0.265 per Taiga share, implying an equity value of $21M. The transaction consolidated a 100% interest in the Fisher property contiguous to the SGO, eliminated an existing 2.5% net smelter return (NSR) royalty on the Fisher property, and added five new properties covering 30,480 ha to complement SSR’s existing exploration platform in the underexplored and highly geologically prospective Province of Saskatchewan. The transaction, which is subject to Taiga shareholder approval, court and regulatory approvals, and customary closing conditions, is expected to close in the first half of 2022.

Since optioning the Fisher Property in 2016, SSR has fulfilled all the minimum work and payment requirements to trigger the current 80/20 joint venture. During this time, SSR has completed extensive systematic exploration including prospecting, soil geochemical sampling, detailed geological mapping, geophysical surveys and 36,897 m of diamond drilling in 95 holes. SSR expenditure to date on the Fisher property totals more than $11M. In addition, SSR has made cash payments to Taiga and predecessor Eagle Plains Resources of more than $2.9M as outlined in the original option agreement.

The acquisition of Taiga would provide SSR with 100% ownership of the Fisher property (33,171 ha) as well as an additional five new properties (34,569 ha) providing new exploration targets south from the SGO to SSR’s 100%-owned Amisk property (Figure 20.1). The deal would also unencumber the Fisher property through the elimination of a 2.5% NSR royalty covering much of the Fisher property.

Based on the technical work completed to-date, the Fisher property appears integral to the future life of mine plan at the SGO. Drilling completed over the past four years indicates strong exploration potential across the Fisher property with several large target areas yet to be tested. Initial due-diligence work completed for the additional five properties also indicates high exploration potential evidenced by excellent historical results that have yet to be properly tested using modern exploration techniques in a robust gold-price environment.

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Figure 20.1 Location of the Six Properties included in the Taiga Gold Transaction with Respect to the Seabee Gold Operation

SSR, 2022

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21 OTHER RELEVANT DATA AND INFORMATION

There is no other relevant data or information.

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22 INTERPRETATION AND CONCLUSIONS

22.1 Mineral Resources

Mineral Resources in the Seabee21TRS are reported in accordance with subpart 1300 of US Regulation S-K Mining Property Disclosure Rules (S-K 1300).

Areas of uncertainty that may materially impact the Mineral Resource estimates include:

•Assumptions used to generate the data for consideration of reasonable prospects of eventual economic extraction.

•Gap Hangingwall (GHW) mining recovery could be lower, and dilution increased. Early stoping in GHW should be used to confirm mining method parameters for the GHW zone in terms of costs, dilution, and mining recovery. Early development will also provide access to data and metallurgical samples at a bulk scale that cannot be collected at the scale of a drill sample.

•Commodity prices and exchange rates.

•Cut-off grades.

22.2 Mineral Reserves Estimation

Mineral Reserves in the Seabee21TRS are reported in accordance with subpart 1300 of US Regulation S-K Mining Property Disclosure Rules (S-K 1300).

Areas of uncertainty that may impact the Mineral Reserve estimate include:

•Any changes to the resource model as a result of further definition drilling at the site.

•Changes to mining conditions that have an impact to operating costs, production rates or mining recovery factors.

•Commodity prices and exchange rates.

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23 RECOMMENDATIONS

The QPs are not aware of any significant risks and uncertainties that could be expected to affect the reliability or confidence in the information discussed herein.

23.1 Further Assessment

The key areas for further studies / work are:

•Ongoing drilling to expand the Mineral Resource aimed to increase mine life and grade in years 2024 and beyond, as SGO has managed to do for many years.

•Ongoing geotechnical drilling and logging will be required to increase the confidence in geotechnical data as the project develops.

•While the structural analysis provides an impression of the major joint sets across the project area, further geotechnical scanline mapping should be conducted regularly as mining commences to allow for the identification of low angle joints in the hanging wall, localised joint sets and for potential wedges and instabilities.

•Update the Santoy geotechnical model to include the expanded GHW mining zone.

•Implementation of Operational Excellence projects identified based on SSR’s recent operational review may present incremental improvements to production and operating costs.

•Continue with ongoing review of capital and operating cost estimates and performance and productivity tracking.

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24 REFERENCES

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[SMOE] Saskatchewan Ministry of Environment, 2016. Approval to Operate a Pollutant Control Facilities Approval No. Po16-002.

SRK Consulting (Canada) Inc., 2017. Seabee Mine, Tailings Alternatives Assessment (Draft) ICC042.006, March 2016

SRK Consulting (Canada) Inc., 2016. Annual Geotechnical, Inspection of Tailings Facilities and Water Management Ponds ICC042.013, March 2017.

SRK Consulting (Canada) Inc., 2018. Triangle Lake TMF Expansion Detailed Design ICC042.025, July 2018.

SRK Consulting (Canada) Inc., 2018 Annual Geotechnical Inspection of Tailings Facilities and Water Management Ponds ICC042.025, July 2018.

SRK Consulting (Canada) Inc., 2018. Triangle Lake and East Lake Tailings Management Facilities – Dam Breach Analysis – DRAFT ICC042.029.500, August 2019.

SRK Consulting (Canada) Inc., 2019. Triangle Lake and East Lake Tailings Management Facilities – Dam Breach Analysis – DRAFT ICC042.029.500, August 2019.

SRK Consulting (Canada) Inc., 2009. Structural Interpretation of Aeromagnetic Data. Seabee Gold Project Saskatchewan, Canada. Internal report prepared for Claude Resources Inc.

SRK Consulting (Canada) Inc., 2016a. Seabee Mine Quarry Rock ML/ARD Assessment. SRK Consulting (Canada) Inc. 2016b. Seabee Mine Tailings Alternatives Assessment. Claude Resources Inc.

SRK Consulting (Canada) Inc., 2020. Seabee Tailings Operation, Maintenance and Surveillance (OSM) Manual.

SRK Consulting (Canada) Inc., 2017b. Seabee Gold Operation Preliminary Decommissioning and Reclamation Plan, 2016 Update – Final.

SRK Consulting (Canada) Inc., 2017. NI 43-101 Technical Report for the Seabee Gold Operation, Saskatchewan, Canada, 5CS010.001, October 2017. (SGOTR17)

SSR Mining Inc., 2021. Seabee Gold Operation, Standards and Guidelines - Ground Control SGO-1700, May 2021.

Stantec Mining, 2015. Geomechanical Overview of Stope and Pillar Stability for the Santoy Gap Orebody Project no. 169514558, March 2015.

Stauffer, M.R., 1984. Manikewan: and early Proterozoic ocean in central Canada, its igneous history and orogenic closure; Precambrian Research, v. 25, pp. 257–281.

21014Seabee21TRS220916A.docx Page 133 of 135

The Mines Regulations, 2003. Chapter 0-1.1 Reg 2 (effective July 16, 2003).

White, D.J., Lucas, S.D., Hajnal, A., Green, A.G., Lewry, J.F., Weber, W., Bailes, A.H., Syme, E.C., and Ashton, K., 1994. Paleo-Proterozoic Thick-Skinned Tectonics: Lithoprobe Seismic Reflection Results from the Eastern Trans-Hudson Orogen; Canadian Journal of Earth Sciences, v. 31, pp. 458–469.

Wood, C.R., 2016. Structural study of the auriferous Santoy shear zone, northeastern Glennie domain, Saskatchewan; Unpublished masters thesis, University of Regina, Regina, Saskatchewan.

21014Seabee21TRS220916A.docx Page 134 of 135

25 RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT

The Seabee21TRS QPs have relied on the following information provided by SSR in preparing the findings and conclusions in this Technical Report Summary regarding the following aspects of modifying factors:

•Macroeconomic trends, taxes, royalties, data, and assumptions, and interest rates.

•This has been used in Section 19 as described in this section. The QPs have relied exclusively on SSR for this information.

•Marketing information and plans within the control of the registrant.

•This has been used in Sections 16 and 19 as described in those sections

•Legal matters outside the expertise of the qualified person, such as statutory and regulatory interpretations affecting the mine plan.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR. The QPs have relied exclusively on SSR for this information.

•Environmental matters outside the expertise of the qualified person.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR

•Accommodations the registrant commits or plans to provide to local individuals or groups in connection with its mine plans.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR

•Governmental factors outside the expertise of the qualified person.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR

Following a review of the information supplied, the opinion of the QPs is, that it is reasonable to rely on the information provided by SSR as outlined above for use in the Seabee21TRS because a significant environmental and social analysis has been conducted for the project over an extended period, the project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

21014Seabee21TRS220916A.docx Page 135 of 135

Document

Exhibit 96.4

Explanatory Note

SSR Mining Inc. (the “Company”) previously filed the Puna 2021 Technical Report Summary (the “Puna21TRS”), with an effective date of December 31, 2021 and issued on February 23, 2022, as Exhibit 96.4 to its Annual Report on Form 10-K for the year ended December 31, 2021, as amended (the “Form 10-K/A”). The Puna21TRS has been amended to reflect certain revisions in compliance with Subpart 1300 of Regulation S-K, which revisions consist of adding confirmatory statements and other modifications that SSR does not consider material. The amended Puna21TRS has been reissued as of September 29, 2022 and is presented with an effective date of December 31, 2021. The information in this amended Puna21TRS has not been updated to reflect events, information or developments occurring after the effective date.

The Contained Metal amounts for lead and zinc are reported in Table 1.1, Table 1.3, Table 1.5, and Table 11.11 of the Puna21TRS in thousands of pounds (klb) for internal consistency. In the Company’s filings with the Securities and Exchange Commission (the “SEC”), including in its Form 10-K/A, the same information is reported in millions of pounds (Mlb). Any inconsistencies between the amounts reported in the Puna21TRS and the amounts reported in other SEC filings of the Company are generally a result of rounding through the conversion.

The Form 10-K/A reported the inferred contained metal amount of zinc incorrectly on page 37 due to a typographical error. The amount presented in the Puna21TRS in Table 1.3 and Table 11.11 (177,394 klb) is consistent with the information correctly reported elsewhere in the Form 10-K/A.

The Form 10-K/A reported the probable zinc reserve for Chinchillas (Stockpile) incorrectly on page 36 due to a typographical error. The amount presented in the Puna21TRS in Table 1.5 and Table 12.1 (2,056 klb) is consistent with the information correctly reported elsewhere in the Form 10-K/A.

The Form 10-K/A reported the indicated and measured and indicated contained metal amounts of lead in respect of the Pirquitas property incorrectly on page 37 due to immaterial rounding errors. The amounts presented in the Puna21TRS in Table 1.3 and Table 11.11 (895 klb and 1,240 klb, respectively) are correct and consistent with other information regarding lead amounts correctly reported elsewhere in the Form 10-K/A.

This page does not constitute a part of the amended Puna21TRS.

Title Page

Project Name:	Puna Operations
Title:	Puna 2021 Technical Report Summary
Location:	Jujuy, Argentina
Effective Date of Technical Report Summary:	31 December 2021
Effective Date of Mineral Resources:	31 December 2021
Effective Date of Mineral Reserves:	31 December 2021

Qualified Persons (QPs):

•Gregory Gibson, BSc (Mining Engineering), MSc (Mining & Earth Systems Engineering), SME Registered Member (4134135), employed by SSR Mining Inc. as Vice President of Operations - Americas, was responsible for the preparation of Sections 1 to 25.

•Bernard Peters, BEng (Mining), FAusIMM (201743), employed by OreWin Pty Ltd as Technical Director – Mining, was responsible for the overall preparation, the Mineral Reserves estimates, and Sections 1 to 5; Section 10; and Sections 12 to 25.

•Sharron Sylvester, BSc (Geol), RPGeo AIG (10125), employed by OreWin Pty Ltd as Technical Director – Geology, was responsible for the preparation of the Mineral Resources, Sections 1 to 3; Sections 6 to 9; Section 11; and Sections 22 to 25.

OreWin Pty Ltd ACN 165 722 574

140 South Terrace Adelaide 5000

P +61 8 8210 5600 E orewin@orewin.com W orewin.comi

Signature Page

Project Name:	Puna Operations
Title:	Puna 2021 Technical Report Summary
Location:	Jujuy, Argentina
Date of Signing:	29 September 2022
Effective Date of Technical Report Summary:	31 December 2021

/s/ Gregory Gibson

Gregory Gibson, Vice President Operations - Americas, SSR Mining Inc., BSc (Mining Engineering), MSc (Mining & Earth Systems Engineering), SME Registered Member (4134135)

/s/ Bernard Peters

Bernard Peters, Technical Director – Mining, OreWin Pty Ltd, BEng (Mining), FAusIMM (201743)

/s/ Sharron Sylvester

Sharron Sylvester, Technical Director – Geology, OreWin Pty Ltd, BSc (Geol), RPGeo AIG (10125)

21015Puna21TRS220927Rev0.docx ii

TABLE OF CONTENTS

1EXECUTIVE SUMMARY	1
1.1Introduction	1
1.2Land Tenure and Ownership	1
1.3Property Description and Location	2
1.3.1Property Description and Location – Chinchillas	2
1.3.2Property Description and Location – Pirquitas	2
1.4Geological Setting and Mineralisation	2
1.4.1Geological Setting and Mineralisation – Chinchillas	3
1.4.2Geological Setting and Mineralisation – Pirquitas	3
1.5Exploration	3
1.5.1Exploration Activities – Chinchillas	3
1.5.2Exploration Activities – Pirquitas	4
1.6Sample Preparation, Analyses, and Security	4
1.7Data Verification	4
1.8Mineral Resources Estimates	4
1.8.1Mineral Resources Estimate – Chinchillas Property	4
1.8.2Mineral Resources Estimate – Pirquitas Deposit	6
1.9Mineral Reserves Estimates	8
1.9.1Mineral Reserves Estimate – Chinchillas	8
1.10Metallurgy and Processing	9
1.11Environment, Communities, and Permitting	10
1.12Production	10
1.13Capital and Operating Costs	13
1.14Economic Analysis	13
1.15Interpretation and Conclusions	17

21015Puna21TRS220927Rev0.docx iii

1.15.1Mineral Resources	17
1.15.2Mineral Reserves	18
1.16Recommendations	18
1.16.1Further Assessment	18
2INTRODUCTION	19
2.1Terms of Reference	19
2.2Qualified Persons	20
2.3Qualified Persons Property Inspections	20
2.4Units and Currency	20
2.5Effective Dates	20
3PROPERTY DESCRIPTION	21
3.1Location	21
3.2Ownership	22
3.3Mineral Tenure	23
3.3.1Chinchillas Mineral Tenure	23
3.3.2Pirquitas Mineral Tenure	25
3.4Other Significant Factors and Risks	27
4ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	28
4.1Accessibility	28
4.2Physiography, Climate, and Vegetation	28
4.3Local Resources and Infrastructure	29
4.3.1Chinchillas Infrastructure	29
4.3.2Pirquitas Infrastructure	30
5HISTORY	31
5.1Chinchillas History	31

21015Puna21TRS220927Rev0.docx iv

5.2Pirquitas History	32
6GEOLOGICAL SETTING, MINERALISATION, AND DEPOSIT	33
6.1Regional Geology	33
6.1.1The Sub-Andean Belt	33
6.1.2The Eastern Cordillera	33
6.1.3Puna Belt	34
6.2Deposit Types	37
6.3District Geology	39
6.3.1Chinchillas Geology	39
6.3.2Pirquitas Geology	48
7EXPLORATION	54
7.1Surficial Exploration	54
7.2Drilling	54
7.2.1Chinchillas Summary	54
7.2.2Pirquitas Summary	55
7.3Drill Core Handling Protocol	57
8SAMPLE PREPARATION, ANALYSES, AND SECURITY	58
8.1Sample Preparation, Analyses, and Security – Chinchillas	58
8.1.1Sampling Method – Chinchillas	58
8.1.2Sample Custody and Security – Chinchillas	58
8.1.3Sample Preparation – Chinchillas	58
8.1.4Sample Analysis – Chinchillas	58
8.1.5Density – Chinchillas	59
8.1.6Quality Assurance and Quality Control – Chinchillas	59
8.1.7Conclusions and Recommendations – Chinchillas	65
8.2Sample Preparation, Analyses, and Security – Pirquitas	65

21015Puna21TRS220927Rev0.docx v

8.2.1Sampling Method – Pirquitas	65
8.2.2Sample Custody and Security – Pirquitas	66
8.2.3Sample Preparation – Pirquitas	66
8.2.4Sample Analysis – Pirquitas	67
8.2.5Density – Pirquitas	68
8.2.6Quality Assurance and Quality Control – Pirquitas	68
8.2.7Conclusions and Recommendations – Pirquitas	68
9DATA VERIFICATION	69
9.1Database Validation	69
9.1.1Collar Coordinate Validation	69
9.1.2Downhole Survey Validation	69
9.1.3Assay Verification	70
9.2QA/QC Protocol	71
9.3Geological Data Verification and Interpretation	71
9.4Assay Database Verification	71
9.5QP Opinion	71
10MINERAL PROCESSING AND METALLURGICAL TESTING	72
10.1Chinchillas	72
10.1.1Initial Testwork 2013	72
10.1.2Second Phase Testwork 2014	72
10.1.3Third Phase Testwork 2016	73
10.2Socavon / Chinchilla	77
10.2.1Testwork 2018	77
10.3Metallurgical Performance Estimates	78
10.4Recommendations for Additional Testwork	79
10.5QP Opinion	79

21015Puna21TRS220927Rev0.docx vi

11MINERAL RESOURCES ESTIMATES	80
11.1Mineral Resources Estimate – Chinchillas	80
11.1.1Available Data – Chinchillas	80
11.1.2Compositing – Chinchillas	81
11.1.3Exploratory Data Analysis – Chinchillas	81
11.1.4Evaluation of Outlier Grades – Chinchillas	83
11.1.5Geological Model – Chinchillas	83
11.1.6Grade Estimation – Chinchillas	89
11.1.7Density – Chinchillas	89
11.1.8Validation – Chinchillas	90
11.1.9Classification – Chinchillas	91
11.2Pirquitas	92
11.2.1Available Data – Pirquitas	92
11.2.2Exploratory Data Analysis – Pirquitas	93
11.2.3Domain Interpretations – Pirquitas	94
11.2.4Compositing – Pirquitas	98
11.2.5Evaluation of Outlier Grades – Pirquitas	98
11.2.6Continuity Analysis – Pirquitas	98
11.2.7Grades Estimation – Pirquitas	98
11.3Mineral Resources Estimate	102
11.3.1Chinchillas Mineral Resources	103
11.3.2Pirquitas Mineral Resources	105
11.3.3Mineral Resources Statement	106
11.4Comparison with Previous Mineral Resources Estimates	109
11.4.1Chinchillas Comparison – 2021 vs. 2020	109
11.4.2Pirquitas Comparison – 2021 vs. 2011	109

21015Puna21TRS220927Rev0.docx vii

11.5Conclusions and Recommendations	109
11.6QP Opinion	109
11.7Canadian National Instrument 43-101 Standards of Disclosure forMineral Projects	109
12MINERAL RESERVES ESTIMATES	110
12.1Summary	110
12.2Mineral Reserves Statement	110
12.3Factors that Affect the Mineral Reserves Estimates	110
12.4Comparison with Previous Mineral Reserves Estimates	112
12.5Canadian National Instrument 43-101 Standards of Disclosure forMineral Projects	112
13MINING METHODS	113
13.1Geotechnical Review	113
13.1.1Knight Piesold 2015 Site Investigation	113
13.1.2Knight Piesold PFS Pit Slope Design	115
13.1.32021 Pit Design Slope Criteria	117
13.1.4Knight Piesold Hydrogeologic Conceptual Model and PreliminaryPit Inflow Estimates	117
13.1.5NI 43-101 Technical Report	117
13.1.6Summary of Geotechnical Studies Review	118
13.2Mining	119
13.3Mine Design	119
13.4Rock Storage Facilities	121
13.5Mining Equipment and Personnel	121
13.6Production Scheduling	122
13.7General Mine Site Layout	122
14PROCESSING AND RECOVERY METHODS	124

21015Puna21TRS220927Rev0.docx viii

14.1Process Overview for Chinchillas	124
14.1.1Stockpiling and Crushing	125
14.1.2Grinding	125
14.1.3Lead / Silver Flotation	125
14.1.4Zinc Flotation	126
14.1.5Concentrate Handling	127
14.1.6Tailings Handling	127
14.2Process Plant Performance	128
15INFRASTRUCTURE	129
15.1Ore Haulage	129
15.2Gas Pipeline and Power Supply	130
15.3Water Supply	130
15.4Tailings	130
15.5Communications Systems	131
15.6Camp, Office, and Chinchillas Infrastructure	132
15.7Mine Short-Term / Long-Term Ore Stockpiles	132
15.8Rock Storage Facilities	133
15.9Other Pirquitas Infrastructure	133
16MARKET STUDIES	134
16.1Marketing and Metal Prices	134
16.2Concentrate Terms	134
16.3QP Opinion	136
17ENVIRONMENTAL STUDIES, PERMITTING AND PLANS, NEGOTIATIONS, ORAGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS	137
17.1Chinchillas	137
17.1.1Surface Hydrology and Water Quality	137

21015Puna21TRS220927Rev0.docx ix

17.1.2Hydrogeology	138
17.1.3Geochemistry	139
17.1.4Water Management	141
17.1.5Flora and Fauna	143
17.1.6Protected Areas	146
17.2Social and Community Engagement	148
17.2.1Local Communities	148
17.2.2Archaeology	148
17.3Project Permitting	149
17.4Mine Closure	151
17.4.1Closure Activities	151
17.5Pirquitas Mine	151
17.5.1In-Pit Tailings Disposal	151
17.5.2Pirquitas Pit	152
17.5.3Environmental and Social description and Closure	152
17.5.4Closure	155
17.6QP Opinion	156
18CAPITAL AND OPERATING COSTS	157
18.1Capital Costs	157
18.2Operating Costs	157
19ECONOMIC ANALYSIS	159
19.1Economic Assumptions	159
19.1.1Pricing and Discount Rate Assumptions	159
19.1.2QP Opinion on Inputs	160
19.2Overview and Results	160
19.2.1Production and Cost Summary	162

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20ADJACENT PROPERTIES	168
21OTHER RELEVANT DATA AND INFORMATION	169
22INTERPRETATION AND CONCLUSIONS	170
23RECOMMENDATIONS	171
24REFERENCES	172
25RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT	175

TABLES

Table 1.1Summary of Chinchillas Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021)Based on $22.00/oz Silver, $0.95/lb Lead, and $1.15/lb Zinc	5
Table 1.2Summary of Metallurgical Recoveries and Cut-off Values of ChinchillasMineral Resources Estimate Exclusive of Mineral Reserves(as at 31 December 2021)Based on $22.00/oz Silver, $0.95/lb Lead, and $1.15/lb Zinc	6
Table 1.3Summary of Pirquitas Mineral Resources Estimate Exclusive ofMineral Reserves (as at 31 December 2021)Based on $20.00/oz Silver, $1.10/lb Lead, and $1.30/lb Zinc	7
Table 1.4Summary of Metallurgical Recoveries and Ownership of PirquitasMineral Resources Estimate Exclusive of Mineral Reserves(as at 31 December 2021)Based on $20.00/oz Silver, $1.10/lb Lead, and $1.30/lb Zinc	7
Table 1.5Summary of Chinchillas Mineral Reserves Estimate(as at 31 December 2021)Based on $18.50/oz Silver, $0.90/lb Lead, and $1.05/lb Zinc	8
Table 1.6Summary of Metallurgical Recoveries of ChinchillasMineral Reserves Estimate (as at 31 December 2021)Based on $18.50/oz Silver, $0.90/lb Lead, and $1.05/lb Zinc	9
Table 1.7Silver / Lead Concentrate Relationships	9
Table 1.8Zinc Concentrate Relationships	10
Table 1.9Mining Production Statistics	11
Table 1.10Operating Costs Estimate	13
Table 1.11Metal Price Assumptions	13
Table 1.12Key Economic Assumptions	14
Table 1.12Puna21TRS Results Summary	15

21015Puna21TRS220927Rev0.docx xi

Table 1.13Financial Results	16
Table 1.14After-Tax NPV Sensitivity to Silver Price and Discount Rates	16
Table 3.1Chinchillas Exploitation Concessions	23
Table 3.2Pirquitas Operation Surface Rights	25
Table 7.1Drill Programmes Completed at the Chinchillas Property	54
Table 7.2Drilling Programmes Completed at the Pirquitas Property	57
Table 8.1Summary of QA/QC Samples – Chinchillas	59
Table 10.1Bond Ball Mill Work Index Test Results	76
Table 10.2Preliminary Socavon Recovery Relationships	78
Table 10.3Silver / Lead Concentrate Relationships	78
Table 10.4Zinc Concentrate Relationships	78
Table 11.1Estimation Domain Statistics	82
Table 11.2Estimation Domain Capping	83
Table 11.3Cell Model Limits – Chinchillas	89
Table 11.4Classification Parameters – Chinchillas	92
Table 11.5Statistical Summary of Raw Assay Data Used in 2013 Modelling – Pirquitas	94
Table 11.6Pirquitas Mineralisation Domain Codes	95
Table 11.7Cell Model Limits – Pirquitas	99
Table 11.8Pirquitas 2013 Mineral Resources Classification Results, Valid ata Cut-off of 65 g/t AgEq	102
Table 11.9Summary of Chinchillas Mineral Resources Estimate Exclusive ofMineral Reserves (as at 31 December 2021)Based on $22.00/oz Silver, $0.95/lb Lead, and $1.15/lb Zinc	104
Table 11.10Summary of Metallurgical Recoveries and Ownership of ChinchillasMineral Resources Estimate Exclusive of Mineral Reserves(as at 31 December 2021)Based on $22.00/oz Silver, $0.95/lb Lead, and $1.15/lb Zinc	104
Table 11.11Summary of Pirquitas Mineral Resources Estimate Exclusive ofMineral Reserves (as at 31 December 2021)Based on $20.00/oz Silver, $1.10/lb Lead, and $1.30/lb Zinc	105

21015Puna21TRS220927Rev0.docx xii

Table 11.12Summary of Metallurgical Recoveries and Ownership of PirquitasMineral Resources Estimate Exclusive of Mineral Reserves(as at 31 December 2021)Based on $20.00/oz Silver, $1.10/lb Lead, and $1.30/lb Zinc	106
Table 11.13Summary of Puna21TRS Mineral Resources Estimate Exclusive ofMineral Reserves (as at 31 December 2021)	107
Table 11.14Summary of Cut-off Values and Metallurgical Recoveries of Puna21TRSMineral Resources Estimate Exclusive of Mineral Reserves(as at 31 December 2021)	108
Table 12.1Summary of Chinchillas Mineral Reserves Estimate(as at 31 December 2021) Based on $18.50/oz Silver, $0.90/lb Lead, and $1.05/lb Zinc	111
Table 12.2Summary of Metallurgical Recoveries of ChinchillasMineral Reserves Estimate (as at 31 December 2021)Based on $18.50/oz Silver, $0.90/lb Lead, and $1.05/lb Zinc	112
Table 13.1Knight Piesold Recommended Inter-ramp Angles	116
Table 13.2Mine Design Criteria	117
Table 13.3Production Schedule for Chinchillas Project	122
Table 14.1Mill Production Summary 2018 to 2021	128
Table 16.1Metal Price Assumptions	134
Table 16.2Concentrate Marketing Terms and Charges	135
Table 18.1Capital Costs Estimate	157
Table 18.2Operating Costs Estimate	158
Table 19.1Metal Price Assumptions	159
Table 19.2Key Economic Assumptions	159
Table 19.3Puna21TRS Results Summary	161
Table 19.4Production Forecast	162
Table 19.5Cash Costs	164
Table 19.6Operating Costs and Revenues	164
Table 19.7Total Project Capital Costs	165
Table 19.8Financial Results	166
Table 19.9After-Tax NPV Sensitivity to Silver Price and Discount Rates	166

21015Puna21TRS220927Rev0.docx xiii

Table 19.10After-Tax NPV5% Sensitivity to Operating and Capital Cost Changes	167
Table 19.11Estimated Cash Flow	167

FIGURES

Figure 1.1Mining Production Profile	11
Figure 1.2Process Feed Profile	12
Figure 1.3Silver Production	12
Figure 1.4After-Tax Annual and Cumulative Cash Flow	17
Figure 3.1Puna Operation Location	21
Figure 3.2Puna Corporate Structure	22
Figure 3.3Property Map Showing Chinchilla, Chinchilla I and Chinchilla II Concessions	24
Figure 3.4Property Map Showing Pirquitas Concessions	26
Figure 6.1Map Showing Tertiary Volcanism from Mega Caldera ComplexesNear the Chinchillas Deposit	35
Figure 6.2Oligocene-Miocene Volcanic Arc. Subvolcanic Intrusions	36
Figure 6.3Bolivian Tin-Silver-Zinc Belt with Major Deposits	38
Figure 6.4Schematic Geological Model on West–East Cross-Section showingChinchillas Deposit	39
Figure 6.5View of the Chinchillas Deposit, Looking East	40
Figure 6.6Chinchillas Geology Map	41
Figure 6.7Interbedded Sequence of Marine Sandstone and Pelite withNear-Vertical Dip at Chinchillas	42
Figure 6.8Brecciated Basement Sediments with Fine Volcanic Matrix Near theContact Between Pyroclastic Sequence and Basement Sediments	43
Figure 6.9Typical Chinchillas Medium Grained Pyroclastic Breccia	45
Figure 6.10Silver Mantos and Mantos Basement Zones with Drillhole Locations and Mineralised Zones Projected to Surface	47
Figure 6.11East–West Cross-Section with Deep Manto Mineralisation	48

21015Puna21TRS220927Rev0.docx xiv

Figure 6.12Pirquitas Geology Map	49
Figure 6.13Anticline Developed in Interbedded Sandstone, Siltstone and Shale of the Ordovician Acoite Formation, Pircas River Valley, Pirquitas Mine Area	50
Figure 6.14Chevron Fold and High Angle Thrust Fault in the Acoite FormationHost Rocks, North Wall San Miguel Open Pit	51
Figure 6.15Map of the Pircas Valley showing Main Ag-Sn-Zn Vein Systems	52
Figure 7.1Location of Drillhole Collars at the Chinchillas Deposit	55
Figure 8.1Ag Values for the Two Types of Duplicates – Chinchillas	61
Figure 8.2Ag Values from CRM 1-CH and 2-CH – Chinchillas	62
Figure 8.3Ag Values from CRM 3-CH – Chinchillas	62
Figure 8.4Pb Values from CRM 1-CH, 2-CH and 3-CH – Chinchillas	63
Figure 8.5Zn Values from CRM 1-CH and 3-CH – Chinchillas	63
Figure 8.6Zn Values from CRM 2-CH – Chinchillas	64
Figure 8.7Comparison of Ag, Pb, and Zn Laboratory Duplicates – Chinchillas	65
Figure 9.1Vein model with Exploration Drillhole AR-315 Outside of the Vein	70
Figure 10.1Drillhole CGA, Variation in Ag Grade and Fe:S Ratio Downhole	74
Figure 10.2Metallurgical Sample Locations within the Two Pit Shells (Mining Phases)	75
Figure 11.1Isometric View showing the Chinchillas Drillhole Database used inResource Modelling	81
Figure 11.2Cross-Section showing Rock Types and Silver Grades in Drilling	83
Figure 11.3Isometric View of Simplified Lithology Model	84
Figure 11.4Bedding Attitudes Represented as Disks Mapped on Topographic DTM	85
Figure 11.5K-means Four-Cluster Multi-Element Geochemistry Model	86
Figure 11.6Volumes with >30% Probability of Exceeding 15 g/t Ag	87
Figure 11.7Estimation Domains based on K-Means Clusters ofMulti-Element Geochemistry	88

21015Puna21TRS220927Rev0.docx xv

Figure 11.8Chinchillas Grade Tonnage Comparison within Pit 3 Volume– Ag, Zn, and Pb	90
Figure 11.9Chinchillas Cross-Section at 7,512,400 mN showing Ag Composites and Estimates	91
Figure 11.10Isometric View showing the Pirquitas Drillhole Database used inResource Modelling	93
Figure 11.11Pirquitas Plan showing Drillholes Coded using Wireframe Vein Models	95
Figure 11.12Plan showing Grade Control Bench Assays	96
Figure 11.13Isometric View of Final Pirquitas Wireframe Vein Models– 2013 Resource Model	97
Figure 11.14Pirquitas Cross-Section showing Two Distinct Oploca Vein Sets	100
Figure 13.1Overview of Knight Piesold Site Investigation	114
Figure 13.2Surface Mapping and orientated data from borehole GCA-209-G	115
Figure 13.3Chinchillas 2021 Ultimate Pit Design	120
Figure 13.4Long-Section of the 2021 Pit Design	121
Figure 13.5Chinchillas General Mine Site Layout	123
Figure 14.1Chinchillas Processing Flow Sheet Overview	124
Figure 14.2Chinchillas Crushing Circuit	125
Figure 14.3Grinding and Lead / Silver Recovery Circuits for Chinchillas	126
Figure 14.4Zinc Recovery Circuit for Chinchillas	127
Figure 15.1Access Road for the Project and Proposed Modifications	129
Figure 15.2Alignment and Gradient of the Tailings Line for In-pit Disposal	131
Figure 16.1Concentrate Production	135
Figure 17.1Response Test Hydraulic Conductivity by Lithology	138
Figure 17.2Project General Arrangement and Water Management Features	142
Figure 17.3Grassland Steppes on the Western Edge of the Project Area	143
Figure 17.4Shrub Land on the Northern Edge of the Project Area	144

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Figure 17.5Vega Habitat	145
Figure 17.6Laguna de los Pozuelos Buffer Zones	147
Figure 19.1Mining Production Profile	162
Figure 19.2Process Feed Profile	163
Figure 19.3Silver Production	163
Figure 19.4After-Tax Annual and Cumulative Cash Flow	166

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1 EXECUTIVE SUMMARY

1.1 Introduction

This Puna 2021 Technical Report Summary (Puna21TRS) is an independent Technical Report Summary that has been prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300) for SSR Mining Inc. (SSR), on the Puna project (Puna, or the Project).

The Project comprises the Chinchillas property and the Pirquitas property, both of which are located the Jujuy Province in far north Argentina.

Puna is directly owned (100%) by SSR through a subsidiary company Puna Operations Inc. (POI), which, through other 100% owned subsidiaries owns Mina Pirquitas S.A. (MPSA). MPSA operates the Project.

Ore from the Chinchillas mine is transported to the Pirquitas plant for processing. The Chinchillas mine is located approximately 45 km from the Pirquitas plant. The open pit mine at Pirquitas has been completed.

The Puna21TRS Qualified Persons (QPs) have reviewed the supplied data and information and accept this information as being accurate and complete and suitable for use in the Puna21TRS. Information and data supplied by SSR that were outside the areas of expertise of the QPs and was relied upon when forming the findings and conclusions of this report are detailed in Section 25. Any individual or entity referenced as having completed work relevant to the Puna21TRS, but not identified therein as a QP, does not constitute a QP for the Puna21TRS.

The Puna21TRS should be construed in light of the methods, procedures, and techniques used to prepare the Puna21TRS. Sections or parts of the Puna21TRS should not be read in isolation of, or removed from, their original context.

1.2 Land Tenure and Ownership

The Chinchillas property is composed of three contiguous claims totalling 2,041 ha, and the Pirquitas property includes surface rights covering an area of approximately 7,500 ha, which is used for purposes such as housing, infrastructure facilities, processing facilities, tailings facility and other facilities to support mining operations for the Project.

POI holds a 100% interest in the property.

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1.3 Property Description and Location

1.3.1 Property Description and Location – Chinchillas

The Chinchillas property is located in the Puna region of north-western Argentina, in the province of Jujuy, department of Rinconada, approximately 290 km north-west of San Salvador de Jujuy, the capital of Jujuy Province. The property is centred at approximately 3,473,150E and 7,512,360N (Gauss Kruger, Argentina, Posgar Zone 3; 22°30′13″ S, 66°15′39″ W) at elevations ranging from 4,000–4,200 metres above sea level (masl).

Access to the Chinchillas property is by paved road to the town of Abra Pampa via National Route No. 9 and then 66 km west across public gravel roads, through the village of Santo Domingo. Santo Domingo is serviced with electricity and water. Abra Pampa has a hospital, and, along with San Salvador, provides other supplies necessary for exploration.

Access between the Pirquitas Operation and the Chinchillas property is via National Route No. 40 that leads to Provincial Route No. 70.

1.3.2 Property Description and Location – Pirquitas

The Pirquitas property is also located in the Rinconada Department in the Province of Jujuy, approximately 45 km south-west of the Chinchillas property and approximately 335 km north-west of San Salvador. Activities at the Property are centred at 22°42′ south latitude and 66°30′ west longitude at elevations of between 4,000–4,450 masl.

1.4 Geological Setting and Mineralisation

The Chinchillas and Pirquitas properties are within the Bolivian tin-silver-zinc belt that extends from the San Rafael tin copper deposit in southern Peru into the Puna region of Jujuy. Deposits with similar environments and styles of mineralisation include San Cristóbal, Potosí, and Pulacayo.

These deposits are generally associated with intrusion of dacite dome complexes. Mineralisation is hosted in shear zones and breccias within the dacite domes and/or shear zones and breccias within the host rocks. More rarely, as in the case of the Chinchillas property and San Cristóbal, the deposits also contain flat-lying manto bodies within sedimentary and pyroclastic rocks that are cut by ‘feeder’ shear zones. All the deposits have large vertical extents.

The Chinchillas and Pirquitas properties are within the Puna geological belt. Stratigraphy in the belt includes metamorphosed Proterozoic sedimentary rocks in the basement, Paleozoic marine back arc sedimentary rocks, and younger volcanic and continental sedimentary rocks. In the Jujuy Province, the Puna terrane is an important host for mineral deposits, including mesothermal quartz veins with native gold and base metals; polymetallic quartz-sulfide veins with base and precious metals; gold, tin, and copper placer deposits; sedimentary exhalative (SEDEX) deposits with lead zinc–silver; and Bolivian-type tin-silver-sulfide veins related to intrusive stocks.

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1.4.1 Geological Setting and Mineralisation – Chinchillas

The Chinchillas deposit is within a dacitic volcanic centre. The deposit was controlled by a dilational jog on a regional scale east–west trending fault where magma has intruded through marine meta-sedimentary basement rocks. The explosive volcanic eruption resulted in an elliptical shaped topographic depression approximately 2 km long by 1.6 km wide, subsequently infilled with pyroclastic rocks including breccias and tuffs. At the contact between pyroclastic volcanic rocks and basement metasedimentary rock, a wide zone of hydraulic fracturing and brecciation formed. Dacitic lavas, flow domes and subvolcanic intrusions occur on the southern margin of the basin at the contact between metasedimentary and pyroclastic rocks.

Significant silver-lead-zinc mineralisation occurs in four main areas at Chinchillas: the Silver Mantos and Mantos Basement zones in the west part of the Project, and the Socavon del Diablo and Socavon Basement zones in the east part.

1.4.2 Geological Setting and Mineralisation – Pirquitas

The Pirquitas deposit is hosted by the Ordovician Acoite Formation, a strongly folded package of low-grade metamorphosed marine sandstone, siltstone, and minor shale beds. These rocks crop out within fault-bounded and likely uplifted structural blocks that occur south-west and east of the mine area. Late Ordovician to Early Devonian compressional tectonism resulted infolding of the Paleozoic sedimentary rocks and development of well-defined axial planar cleavage. High-angle thrust faults were also generated during this event. In the area of the mine, axial planes of folds strike north to north–north-east and are sub-vertical to moderately inclined.

Sulphide-rich veins cut the axial planes of the folds and the related axial planar cleavage at high angles. Four main vein sets are recognised on the Pirquitas property.

Bolivian-type Ag-Sn deposits generally consist of sulphide and quartz-sulphide vein systems typically containing cassiterite and a diverse suite of base and trace metals, including silver, in a complex assemblage of sulphide and sulfosalt minerals. The vein systems are generally spatially and likely genetically associated with epizonal (subvolcanic) quartz-bearing peraluminous intrusions 1–2 km in diameter but may be entirely hosted by the country rocks into which the intrusive stocks were emplaced.

1.5 Exploration

1.5.1 Exploration Activities – Chinchillas

At Chinchillas, surface exploration programmes included detailed mapping with emphasis on structure, rock chip sampling, trenching, soil sampling and talus sampling. These programmes identified major structural zones, the strong east–west control on basin formation, and new mineralised target areas. Geophysical surveys were also conducted, including induced polarisation / resistivity (IP), controlled source audio-frequency magneto-telluric (CSAMT) surveys, and magnetics.

Prior to SSR, Golden Arrow completed five drilling programmes (~50,000 m of drilling) that contributed to the initial resource database.

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1.5.2 Exploration Activities – Pirquitas

Prior to SSR, the Sunshine Mining and Refining Company (Sunshine) completed comprehensive mineral exploration on the Pirquitas property, including geological mapping, geophysical surveying (44 line-kilometres of ground magnetics and 19.2 line-kilometres of induced polarisation surveying), underground rock samplin,g and multiple programmes of reverse circulation (RC) and diamond drilling (DD). In May 2004, SSR fully acquired Sunshine’s ownership in the property and continued to advance the project through detailed drilling, underground resource definition, and mining.

1.6 Sample Preparation, Analyses, and Security

All drilling was completed by professional drilling companies using standard industry methods.

Sample and assay procedures applied in the drilling programme are consistent with generally accepted industry best practices. The statistical analysis of quality control data shows good accuracy and precision with no significant contamination.

1.7 Data Verification

No material sample bias was identified during the review of the drill data and assays. SSR has identified an issue with some pre-2009 drillhole collar locations in the Pirquitas drilling and has made a reasonable attempt at rectifying the issue. This issue is not expected to have a material effect on the quantity and quality of the Mineral Resources inventory and should be able to be managed operationally.

Data collection procedures are in accordance with generally accepted industry best practices and the resultant data is suitable for use in mineral resource estimation.

1.8 Mineral Resources Estimates

1.8.1 Mineral Resources Estimate – Chinchillas Property

The Mineral Resources have been estimated for the Chinchillas property in conformity with generally accepted practices and reported in accordance with S-K 1300. The Chinchillas resource model presented in this Puna21TRS was completed on 28 August 2020.

The Mineral Resources estimate has been generated for the Mantos deposit from drillhole sample assay results and the interpretation of a geological model which relates to the spatial distribution of silver, lead, and zinc. Interpolation characteristics were defined based on the geology, drillhole spacing, and geostatistical analysis of the data.

Estimations were made from three-dimensional (3D) cell models based on geostatistical applications using commercial software. The model uses a cell size of 8 m L x 8 m W x 5 m H.

The Mineral Resources were classified according to proximity to sample data locations.

Table 1.1 and Table 1.2 summarise the estimate of Mineral Resources for the Chinchillas project, effective as at 31 December 2021. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues. The quantity and grade of reported Inferred Mineral Resources are uncertain in nature and there has been insufficient exploration to classify these Inferred Mineral Resources as Indicated or Measured Mineral Resources. It cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resources as a result of continued exploration.

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In determining the cut-off grade, the reasonable prospects for eventual economic extraction requirement generally implies that the quantity and grade estimates meet certain economic thresholds taking into account an open pit extraction scenario with road transport and processing at the Pirquitas Operation. This includes consideration of the technical and economic parameters listed above, but also includes additional operating costs, estimated at $12/t, related to the handling and transportation of ore from the Chinchillas property to the Pirquitas Operation.

1.8.1.1 Mineral Resources Estimate – Socavon Deposit

A review of the pit optimisation work for the Socavon deposit was undertaken using the NSR and other assumptions used for the Mantos deposit. The review concluded that there was no suitable pit shell produced to meet the standard of reasonable prospects for extraction. Therefore, the Socavon Mineral Resources previously reported by SSR have not been included in the 2021 Puna Mineral Resources.

Table 1.1 Summary of Chinchillas Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $22.00/oz Silver, $0.95/lb Lead, and $1.15/lb Zinc

Mineral Resources Classification		Tonnage<br>(kt)		Grades						Contained Metal
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(koz)		Lead<br>(klb)		Zinc<br>(klb)
Measured		1,110		99.2		0.86		0.31		3,540		21,015	7,552
Indicated		4,904		101.1		0.88		0.19		15,943		95,632	20,454
Measured + Indicated		6,013		100.8		0.88		0.21		19,483		116,647	28,006
Inferred		165		101.9		0.48		0.16		540		1,746	582

1. Mineral Resources are reported based on 31 December 2021 topography surface.

2. The Mineral Resources are contained within a pit shell generated using an NSR cut-off of $33.20/t.

3. The Mineral Resources estimates are based on metal price assumptions of $22.00/oz silver, $0.95/lb lead, and $1.15/lb zinc.

4. Metallurgical recoveries vary with grade and average recoveries are, 98% silver, 95% lead and 63% for zinc.

5. The point of reference for Mineral Resources is the point of feed into the processing facility.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

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Table 1.2 Summary of Metallurgical Recoveries and Cut-off Values of Chinchillas Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $22.00/oz Silver, $0.95/lb Lead, and $1.15/lb Zinc

Mineral Resources Classification		Tonnage<br>(kt)		Grades						Metallurgical Recovery				Cut-off<br><br>NSR<br><br>($/t)
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(%)		Lead<br>(%)		Zinc<br>(%)
Measured		1,110		99.2		0.86		0.31		98		95	63	33.2
Indicated		4,904		101.1		0.88		0.19		98		95	63	33.2
Measured + Indicated		6,013		100.8		0.88		0.21		98		95	63	33.2
Inferred		165		101.9		0.48		0.16		98		95	63	33.2

1. Mineral Resources are reported based on 31 December 2021 topography surface.

2. The Mineral Resources are contained within a pit shell generated using an NSR cut-off of $33.20/t.

3. The Mineral Resources estimates are based on metal price assumptions of $22.00/oz silver, $0.95/lb lead, and $1.15/lb zinc.

4. Metallurgical recoveries vary with grade and average recoveries are, 98% silver, 95% lead and 63% for zinc.

5. The point of reference for Mineral Resources is the point of feed into the processing facility.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

1.8.2 Mineral Resources Estimate – Pirquitas Deposit

The Mineral Resources have been estimated for the Pirquitas property in conformity with generally accepted guidelines and reported in accordance with S-K 1300. The Pirquitas resource model presented in this Puna21TRS was completed on 30 September 2013 and the reasonable prospects of the underground mining scenario was completed on 24 January 2018.

The Mineral Resources estimate has been generated for the Mining Area veins from drillhole sample assay results and the interpretation of a geological model which relates to the spatial distribution of silver and zinc. Interpolation characteristics were defined based on the geology, drillhole spacing, and geostatistical analysis of the data.

Estimations were made from 3D cell models based on geostatistical applications using commercial software. The model uses a cell size of 4 m x 4 m x 8 m cells to be compatible with the grade control model.

The Mineral Resources were classified according to proximity to sample data locations.

Table 1.3 and Table 1.4 summarise the estimate of Mineral Resources for the Pirquitas project effective as of 31 December 2021. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues. The quantity and grade of reported Inferred Mineral Resources are uncertain in nature and there has been insufficient exploration to classify these Inferred Mineral Resources as Indicated or Measured Mineral Resources. It cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resources as a result of continued exploration.

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Table 1.3 Summary of Pirquitas Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $20.00/oz Silver, $1.10/lb Lead, and $1.30/lb Zinc

Mineral Resources Classification		Tonnage<br>(kt)		Grades						Contained Metal
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(koz)		Lead<br>(klb)		Zinc<br>(klb)
Measured		79		444.5		0.197		1.17		1,129		343	2,044
Indicated		2,555		287.7		0.016		4.56		23,627		895	256,672
Measured + Indicated		2,634		292.4		0.021		4.46		24,756		1,240	258,715
Inferred		1,080		206.9		0.004		7.45		7,185		95	177,394

1. The Mineral Resources estimates are contained within underground mining shapes based on $90/t to $100/t NSR cut-off.

2. The Mineral Resources estimates are based on metal price assumptions of $20.00/oz silver, $1.30/lb zinc, and $1.10/lb lead.

3. Metallurgical recoveries vary with grade and on average are, 87% silver and 85% for zinc and 50% for lead.

4. The point of reference for Mineral Resources is the point of feed into the processing facility

5. Mineral Resources are reported exclusive of Mineral Reserves.

6. SSR has 100% ownership of the Project.

7. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

8. Totals may vary due to rounding.

Table 1.4 Summary of Metallurgical Recoveries and Ownership of Pirquitas Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $20.00/oz Silver, $1.10/lb Lead, and $1.30/lb Zinc

Mineral Resources Classification		Tonnage<br>(kt)		Grades						Metallurgical Recovery				Cut-off<br><br>NSR<br><br>($/t)
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(%)		Lead<br>(%)		Zinc<br>(%)
Measured		79		444.5		0.197		1.17		87		50	85	100
Indicated		2,555		287.7		0.016		4.56		87		50	85	100
Measured + Indicated		2,634		292.4		0.021		4.46		87		50	85	100
Inferred		1,080		206.9		0.004		7.45		87		50	85	100

1. The Mineral Resources estimate is contained within underground mining shapes based on $90/t to $100/t NSR cut-off.

2. The Mineral Resources estimates are based on metal price assumptions of $20.00/oz silver, $1.30/lb zinc, and $1.10/lb lead.

3. Metallurgical recoveries vary with grade and on average are, 87% silver and 85% for zinc and 50% for lead.

4. The point of reference for Mineral Resources is the point of feed into the processing facility

5. Mineral Resources are reported exclusive of Mineral Reserves.

6. SSR has 100% ownership of the Project.

7. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

8. Totals may vary due to rounding.

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1.9 Mineral Reserves Estimates

1.9.1 Mineral Reserves Estimate – Chinchillas

Open pit mining is carried out by MPSA as an owner-mining operation with ore hauled from the Chinchillas pit to the Pirquitas plant. The Mineral Reserves were developed based on mine planning work completed in 2021 that included pit optimisation and re-design of the pit phases. Table 1.5 and Table 1.6 summarise the Mineral Reserves for Chinchillas. The Chinchillas Mineral Reserves estimate has been generated for the Mantos deposit based on the following inputs: metal prices, resource model, geotechnical information, operating costs, mineral processing recoveries, concentrate transport, and off site costs and charges. Costs for all areas of the operation are estimated from actual costs. These were used to calculate a net smelter return (NSR) and $44.11/t NSR was used for the Mineral Reserves cut-off.

Metal prices for the Mineral Reserves cut-off were estimated after analysis of consensus industry forecasts and compared to metal prices used in other published studies. The prices selected were then reduced from the average long-term prices to take a conservative view of the long-term price. The long-term prices for the cut-off were assumed to apply from the start of 2026. The metal prices are representative of the range of price estimates publicly reported for Mineral Reserves cut-offs.

Table 1.5 Summary of Chinchillas Mineral Reserves Estimate (as at 31 December 2021) Based on $18.50/oz Silver, $0.90/lb Lead, and $1.05/lb Zinc

Mineral Reserves Classification		Tonnage<br>(kt)		Grade						Contained Metal
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(koz)		Lead<br>(klb)		Zinc<br>(klb)
Proven		2,379		168.9		1.33		0.34		12,918		69,735	17,827
Probable		5,041		155.3		1.29		0.25		25,174		143,344	27,780
Probable Stockpiles		187		141.0		1.33		0.50		846		5,470	2,056
Proven + Probable		7,606		159.2		1.30		0.28		38,938		218,681	47,692

1. Mineral Reserves are reported based on 31 December 2021 topography surface.

2. The Mineral Reserves estimates are based on metal price assumptions of $18.50/oz silver, $0.90/lb lead, and $1.05/lb zinc.

3. The Mineral Reserves estimates are reported at a cut-off grade of $44.11/t NSR.

4. Economic analysis for the Mineral Reserves has been prepared using long-term metal prices of $21.00/oz silver, $0.90/lb lead, and $1.20/lb zinc

5. Processing recoveries vary based on the feed grade. The average recovery is estimated to be 98% for silver, 95% for lead and approximately 63% for zinc.

6. Metals shown in this table are the contained metals in ore mined and processed.

7. The point of reference for Mineral Resources is the point of feed into the processing facility.

8. SSR has 100% ownership of the Project.

9. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

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Table 1.6 Summary of Metallurgical Recoveries of Chinchillas Mineral Reserves Estimate (as at 31 December 2021) Based on $18.50/oz Silver, $0.90/lb Lead, and $1.05/lb Zinc

Mineral Reserves Classification		Tonnage<br>(kt)		Grades					Metallurgical Recovery					Cut-off<br><br>NSR<br><br>($/t)
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(%)			Lead<br>(%)	Zinc<br>(%)
Proven		2,379		168.9		1.33		0.34	98			95	63	44.11
Probable		5,041		155.3		1.29		0.25	98			95	63	44.11
Probable Stockpiles		187		141.0		1.33		0.50	98			95	63	44.11
Proven + Probable		7,606		159.2		1.30		0.28	98			95	63	44.11

1. Mineral Reserves are reported based on 31 December 2021 topography surface.

2. The Mineral Reserves estimates are based on metal price assumptions of $18.50/oz silver, $0.90/lb lead, and $1.05/lb zinc.

3. The Mineral Reserves estimates are reported at a cut-off grade of $44.11/t NSR.

4. Economic analysis for the Mineral Reserves has been prepared using long-term metal prices of $21.00/oz silver, $0.90/lb lead, and $1.20/lb zinc

5. Processing recoveries vary based on the feed grade. The average recovery is estimated to be 98% for silver, 95% for lead and approximately 63% for zinc.

6. Metals shown in this table are the contained metals in ore mined and processed.

7. The point of reference for Mineral Resources is the point of feed into the processing facility.

8. SSR has 100% ownership of the Project.

9. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

1.10 Metallurgy and Processing

The metallurgical testing of Chinchillas ore types commenced in 2013 and continued until 2016. The first testwork was focused on silver recovery by both leaching and flotation methods with flotation proving to be superior at this early stage. The second programme continued process development of flotation into separate lead / silver and zinc concentrates. The third testwork campaign was designed to advance the flotation process and test specifically these ore types in the Pirquitas mill flow sheet.

The Pirquitas process plant operating performance since commencement on Chinchillas ores is used to provide the concentrate grade recovery and mass pull relationships, Table 1.7 and Table 1.8.

Table 1.7 Silver / Lead Concentrate Relationships

Variable	Variable Formula
Ag Recovery	(–0.0631 x Pb recovery2) + (11.655 x Pb recovery) -447.4
Pb Recovery	(–2.6303x Pb Feed2) + (12.329 x Pb Feed) + 80.654
Zn Recovery	(-5.2817 x Zn Feed2)+(Zn Feed x –6.31) + 20.546
Mass Pull	(–0.0024 x Pb Feed2) + (0.0164 x Pb Feed)+-0.0007

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Table 1.8 Zinc Concentrate Relationships

Variable	Variable Formula
Ag recovery	(–3.4843 x Zn feed2) + (7.2499 x Zn feed)+0.8295
Pb recovery	(0.024 x (Pb feed / Zn feed)2) + (-0.5988 x (Pb feed / Zn feed)+ 3.1292
Zn recovery	(–195921 x (mass pull Zn)2 + (5620.3 x mass pull Zn)+28.709
Mass Pull	(0.007 x Zn feed2) + (0.0041 x Zn feed+0.0011

1.11 Environment, Communities, and Permitting

There are seven communities located in the project's area of influence. and are included in management plans for training and capacity building.

The Puna project does not intrude upon any protected areas. Water quality in the surface waters draining the Project area is typical of a mineralised zone, including some observed elevated metals parameters, but with generally neutral pH. The waste rock is expected to be largely non-acid generating, with a small portion that may be weakly acid generating under certain oxidising conditions. The waste rock with potential for acid production will be placed so any drainage will report to the pit and avoid introduction to the environment.

Although there is no specific mine closure legislation nor bonding requirements in Argentina, a conceptual closure plan has been developed for the Project. Closure costs are estimated at $30.6M. MPSA is also responsible for the closure costs associated with the Pirquitas pit.

An Environmental and Social Impact Assessment (ESIA) was completed for the Chinchillas project and submitted to the regulatory authorities of the province of Jujuy for review, with the license obtained in December 2017. The biannual update of the ESIA was submitted in due time and form, being pending approval by the regulatory authorities of the province of Jujuy. In addition, an addendum to the ESIA for the Pirquitas mine was obtained from MPSA to use the Pirquitas pit for tailings deposition at the Pirquitas Operation, and this authorization must be renewed.

1.12 Production

Future proposed mine production has been scheduled to optimise the mine output and meet the plant capacity. The mining production forecasts are shown in Table 1.9. Mine, process, and metal production are shown in Figure 1.1 to Figure 1.3.

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Table 1.9 Mining Production Statistics

Item	Unit	Total LOM
Ore Processed
Processed	kt	7,352
Ag Feed Grade	g/t	160
Pb Feed grade	%	1.32
Zn Feed grade	%	0.29
Silver Recovery	%	95.5
Concentrate Produced
Lead Concentrate – in Stockpile	kt	4
Zinc Concentrate – in Stockpile	kt	1
Lead Concentrate – Produced	kt	135
Zinc Concentrate – Produced	kt	27
Lead Concentrate - Total	kt	139
Zinc Concentrate - Total	kt	28
Metal Produced
Silver	koz	37,210
Lead	Mlb	204
Zinc	Mlb	29

Metal produced includes current concentrate stockpiles containing 242 koz silver and 5 Mlb lead.

Figure 1.1 Mining Production Profile

OreWin, 2021

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Figure 1.2 Process Feed Profile

OreWin, 2021

Figure 1.3 Silver Production

OreWin, 2021

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1.13 Capital and Operating Costs

The cost estimate was prepared by the SSR technical department at site. The QPs reviewed the assumptions, parameters, and methods used to prepare the cost estimate and are of the opinion that they are sufficient for the purposes of validating the economics of the Mineral Reserves. Total capital expenditure is estimated to be $99M.

The life-of-mine (LOM) operating costs are approximately $52.67/t of ore milled, as summarised in Table 1.10.

Table 1.10 Operating Costs Estimate

Cost Component	Amount <br>($M)	LOM Average<br>($/t milled)
Mining	110	15.01
Processing	183	24.95
G&A	93	12.71
Total Operating Costs	387	52.67

1.14 Economic Analysis

The estimates of cash flows have been prepared on a real basis as at 1 January 2022 and a mid-year discounting is used to calculate NPV.

The projected financial results include:

•After-tax NPV at a 5% real discount rate is $228M.

•Mine life of five years.

The estimated total cash costs for the LOM is $11.63/oz payable silver. The all-in sustaining costs (AISC) for the LOM, which includes infrastructure capital, capital development and reclamation, averages $13.57/oz payable silver. Unit costs include concentrate in stockpile. Silver provides the primary revenue for the analysis, with contributions from lead and zinc. Credits from lead and zinc are included in the cash cost.

Metal price assumptions used for the economic analysis are shown in Table 1.11.

Table 1.11 Metal Price Assumptions

Commodity	Unit	2022	2023	2024	2025	Long-Term
Silver	$/oz	24.00	23.00	22.00	21.00	21.00
Lead	$/lb	1.00	0.95	0.93	0.92	0.90
Zinc	$/lb	1.30	1.20	1.20	1.20	1.20

Other key economic assumptions for the discounted cash flow analyses are shown in Table 1.12.

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Table 1.12 Key Economic Assumptions

Model Assumption	Unit	Lead Concentrate	Zinc Concentrate
Treatment Charge and Refinery Charge(TCRC)	$/t Conc.	1,191	724
Payability – Silver	%	95	75
Payability – Lead	%	95
Payability – Zinc	%		85
Deduction – Lead	%	3
Deduction – Zinc	%		8
Minimum Payout Factor	%	63	39
Royalty	%	3	3
Export Duty (revenue minus TCRC's)	%	4.5	4.5
Puna Credit (revenue minus TCRC's)	%	2.5	2.5

The key results of the Puna21TRS are summarised in Table 1.13. The projected financial results at a range of discount rates for undiscounted and discounted cash flows and before and after tax are shown in Table 1.14.

The results of NPV5% sensitivity analysis to a range of changes in silver price (primary commodity) and discount rates is shown in Table 1.15. A chart of the cumulative cash flow is shown in Figure 1.4.

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Table 1.13 Puna21TRS Results Summary

Description	Unit	Total LOM
Ore Processed
Ore Tonnes Treated	kt	7,352
Ag Feed grade	g/t	160
Pb Feed grade	%	1.32
Zn Feed grade	%	0.29
Silver Recovery	%	95.5
Concentrates
Lead Concentrate – in Stockpile	kt	4
Zinc Concentrate – in Stockpile	kt	1
Lead Concentrate – Produced	kt	135
Zinc Concentrate – Produced	kt	27
Lead Concentrate - Total	kt	139
Zinc Concentrate - Total	kt	28
Metal Produced
Silver	koz	37,210
Lead	Mlb	204
Zinc	Mlb	29
Key Financial Results
Mine Site Cash Cost	$/oz payable silver	11.61
Royalties and Refining Costs1	$/oz payable silver	6.10
Credits	$/oz payable silver	–6.08
Total Cash Costs (CC) (after credits)1	$/oz payable silver	11.63
All-in Sustaining Costs (AISC)	$/oz payable silver	13.57
Site Operating Costs	$/t milled	52.67
Average Silver Price	$/oz	22.38
NPV1	$M	228
Discount Rate	%	5
Mine Life	years	5

1 Metal produced includes current concentrate stockpiles containing 242 koz silver and 5 Mlb lead.

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Table 1.14 Financial Results

Discount Rate		NPV (M)
	Before-Tax
Undiscounted		279	253
2%		268	242
5%		253	228
10%		231	206
12%		223	199
15%		212	188
18%		202	179
20%		195	173

All values are in US Dollars.

Note: NPV includes concentrate in stockpile

Table 1.15 After-Tax NPV Sensitivity to Silver Price and Discount Rates

After-Tax NPV		Silver Price (/oz)
	10.00
Discount Rate		M	M	M	M	M	M	M	M	$M
Undiscounted		–17	105	192	204	253	278	327	401	474
2%		–17	101	183	195	242	266	313	384	454
5%		–16	95	172	183	228	250	294	360	427
10%		–14	86	156	166	206	226	266	327	387
12%		–14	83	150	160	199	218	257	315	373

All values are in US Dollars.

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Figure 1.4 After-Tax Annual and Cumulative Cash Flow

OreWin, 2021

1.15 Interpretation and Conclusions

1.15.1 Mineral Resources

Mineral Resources for the Puna21TRS are reported in accordance with S-K 1300.

Areas of uncertainty that may materially impact the Mineral Resources estimates include:

•Assumptions used to generate the data for consideration of reasonable prospects of eventual economic extraction for the Puna deposit.

•Environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues.

•Commodity prices and exchange rates.

•Cut-off grades.

1.15.1.1 Chinchillas

The resource model developed for the Chinchillas deposit uses accepted modelling and grade estimation methods. The model is a reasonable reflection of deposit geology. The approach used to generate the cell model is in accordance with accepted industry standards. The QP has checked the data and methods used to develop the resource model and has validated the resource models. The methods used for the estimate of Mineral Resources and Mineral Reserves is in accordance with S-K 1300.

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1.15.1.2 Pirquitas

The resource model developed for the Pirquitas deposit uses accepted modelling and grade estimation methods. The model is a reasonable reflection of deposit geology. The approach used to generate the cell model is in accordance with accepted industry standards. The QP has checked the data and methods used to develop the resource model and has validated the resource models. The methods used for the estimate of Mineral Resources and Mineral Reserves is in accordance with S-K 1300.

1.15.2 Mineral Reserves

Mineral Reserves for the Puna21TRS are reported in accordance with S-K 1300.

Areas of uncertainty that may impact the Mineral Reserves estimate include:

•Any changes to the resource model as a result of further definition drilling at the site.

•Changes to mining conditions that have an impact to operating costs, production rates or mining recovery factors.

•Commodity prices and exchange rates.

1.16 Recommendations

The QPs are not aware of any significant risks and uncertainties that could be expected to affect the reliability or confidence in the information discussed herein.

1.16.1 Further Assessment

The key areas for further studies / work are:

•Potential remains to expand the current Mineral Resource, and to define new Mineral Resources on the property.

•Optimisation of metal prices and cost input parameters.

•More detailed planning and design for rock storage and the general site layout.

•Additional metallurgical laboratory testwork as detailed in Section 10.4.

•Update site standard operating procedures to include a transparent Mineral Resources estimation and Mineral Reserves estimation process, clearly documenting the key input parameters applied, and an audit trail of approvals for each phase of the work performed.

•Continue with ongoing review of capital and operating cost estimates and performance and productivity tracking.

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2 INTRODUCTION

The Puna21TRS has been in prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300). This is the first Technical Report Summary for the Puna project.

The Puna project (Puna, or the Project) is directly owned (100%) by SSR Mining Inc. (SSR) through a subsidiary company Puna Operations Inc. (POI) which through other 100% owned subsidiaries owns Mina Pirquitas S.A. (MPSA). MPSA operates the project. SSR has reported that the total cost of the gross mineral properties, plant, and equipment as of 31 December 2021 was $372.4M.

SSR is a gold mining company with four producing assets located in the USA, Turkey, Canada, and Argentina, and with development and exploration assets in the USA, Turkey, Mexico, Peru, and Canada. SSR is listed on the NASDAQ Capital Markets (NASDAQ:SSRM), the Toronto Stock Exchange (TSX:SSRM), and the Australian Stock Exchange (ASX:SSR).

The Project comprises Chinchillas and the Pirquitas property, both of which are located the Jujuy Province in far north Argentina. Ore from the Chinchillas mine is transported to the Pirquitas plant for processing. The Chinchillas mine is located approximately 45 km from the Pirquitas plant. The open pit mine at Pirquitas has been completed.

2.1 Terms of Reference

The Puna21TRS is an independent Technical Report Summary (TRS) on the Project, prepared for SSR by the Puna21TRS Qualified Persons (QPs). The Puna21TRS is based on information and data supplied to the QPs by SSR and other parties where necessary. Any individual or entity referenced as having completed work relevant to the Puna21TRS, but not identified therein as a QP, does not constitute a QP. Puna21TRS QPs have reviewed the supplied data and information and accept this information as being accurate and complete and suitable for use in the Puna21TRS. The primary source of data for the Puna21TRS is the Puna 2021 Project Update Report.

Section 25 describes any information and data supplied by SSR that was outside the areas of expertise of the QPs and was relied upon when forming the findings and conclusions of this report.

The QPs have used their experience and industry expertise to produce the estimates and approximations in the Puna21TRS. It should be noted that all estimates and approximations contained in the Puna21TRS will be prone to fluctuations with time and changing industry circumstances.

The purpose of this Puna21TRS is to report the Mineral Resources and Mineral Reserves for the project. This report is a Feasibility Study (FS) that represents forward-looking information. The forward-looking information includes metal price assumptions, cash flow forecasts, projected capital and operating costs, metal recoveries, mine life and production rates, and other assumptions used in the FS. Readers are cautioned that actual results may vary from those presented. The factors and assumptions used to develop the forward-looking information, and the risks that could cause the actual results to differ materially are presented in the body of this report under each relevant section.

The conclusions and estimates stated in the Puna21TRS are to the accuracy stated in the Puna21TRS only and rely on assumptions stated in the Puna21TRS. The results of further work may indicate that the conclusions, estimates, and assumptions in the Puna21TRS need to be revised or reviewed.

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The Puna21TRS is intended to be used by SSR, subject to the terms and conditions of its contract with OreWin. Recognising that SSR has legal and regulatory obligations, OreWin has consented to the filing of the Puna21TRS with US SEC. Except for the purposes legislated, any other use of this report by any third party is at that party's sole risk.

A list of the references used to prepare the Puna21TRS is provided in Section 24.

2.2 Qualified Persons

The following people served as the QPs as defined in subpart 1300 of US Regulation S-K Mining Property Disclosure Rules (S-K 1300):

•Sharron Sylvester, BSc (Geol), RPGeo AIG (10125), employed by OreWin Pty Ltd as Technical Director – Geology, was responsible for the preparation of the Mineral Resources, Sections 1 to 3; Sections 6 to 9; Section 11; and Sections 22 to 25.

2.3 Qualified Persons Property Inspections

Gregory Gibson visited the project 17 June 2021 to 1 July 2021 and 28 January 2022 to 9 February 2022. The site visits included briefings on geology, mine operations, processing, environmental, permitting, and site inspections of current mining and plant and infrastructure. In addition, Gregory has weekly calls with site leadership regarding the day-to-day operations and quarterly reviews of the operation performance.

Bernard Peters has not visited the site due to travel restrictions.

Sharron Sylvester has not visited the site due to travel restrictions.

2.4 Units and Currency

This Technical Report Summary uses metric measurements except where otherwise noted. The currency used is US dollars ($) unless otherwise stated.

2.5 Effective Dates

The report has several effective dates, as follows:

•Effective date of the Technical Report Summary: 31 December 2021

•Effective date of Mineral Resources: 31 December 2021

•Effective date of Mineral Reserves: 31 December 2021

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3 PROPERTY DESCRIPTION

3.1 Location

The Chinchillas property is located in the Puna region of north-western Argentina, in the province of Jujuy, department of Rinconada, approximately 290 km from the provincial capital of San Salvador de Jujuy (Figure 3.1). The property is centred at approximately at 3,473,150 mE and 7,512,360 mN (Gauss Kruger, Argentina, Posgar Zone 3; 22°30′13″ S, 66°15′39″ W) at elevations ranging from 4,000–4,200 metres above sea level (masl).

The Pirquitas Operation is also located in the Rinconada Department in the Province of Jujuy. The property is centred at 22°42′ south latitude and 66°30′ minutes west longitude. The city of San Salvador de Jujuy, (Jujuy) the provincial capital, is located approximately 335 km south-east of the property (Figure 1.2). The property is characterised by sparsely vegetated, mountainous terrain at elevations of between 4,000–4,450 masl.

Figure 3.1 Puna Operation Location

SSR, 2021

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3.2 Ownership

SSR has 100% ownership of the Puna project. Puna is directly owned (100%) by SSR Mining Inc. (SSR) through a subsidiary company Puna Operations Inc. (POI) which through other 100% owned subsidiaries owns Mina Pirquitas S.A. (MPSA). MPSA operates the project. The corporate structure that links the Puna project and SSR is shown in Figure 3.2.

Figure 3.2 Puna Corporate Structure

SSR, 2021

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3.3 Mineral Tenure

3.3.1 Chinchillas Mineral Tenure

Exploitation concessions in Argentina are called ‘Minas’. Minas are defined by the following categories:

•First Category Minas include substances such as gold, silver, platinum, iron, lead, copper, zinc, aluminium, lithium, potassium, etc., and

•Second Category Minas comprise substances such as precious stones in riverbeds, any metal not included in the first category, and others.

The Mina is comprised of one or more ‘pertenencias’, which are units of mining properties. Pertenencias must be rectangular in shape.

In disseminated deposits, such as Chinchillas, the pertenencias can encompass up to 100 ha. The mining property fee or ‘canon’ for a Mina is charged every year. It is currently ARS$320 per pertenencia per year (article 215 Mining Code).

Individuals are entitled to explore for, exploit, and dispose of Minas as owners by means of a legal licence or legal concession granted by the competent authority under the provisions of the Argentine Mining Code. The legal concessions granted for the exploitation of Minas are valid for an undetermined period of time and are considered ‘real property’ giving the concessionaire the right to recover metals from the subsurface vertically underneath the concession, provided that the title holder complies with the obligations set out in the Argentine Mining Code.

The Chinchillas property consists of three contiguous First Category Minas that cover an area of approximately 2,042.56 ha, as set out in Table 3.1 (see also Figure 3.3).

Table 3.1 Chinchillas Exploitation Concessions

Concession	File No.	Area <br>(ha)
Chinchilla	469-M-56	329
Chinchilla I	079-D-96	830.98
Chinchilla II	1943-V-2013	882.58

The Chinchilla Mina comprises four pertenencias, while the Chinchilla I and Chinchilla II Minas each comprise nine pertenencias.

All Minas are valid and in good standing.

By July 2015, Valle Del Cura S.A. (VDC) completed option payments to earn a 100% interest in the Chinchilla and Chinchilla I properties, to a total of $1,866,000 paid.

Subsequently, Mina Pirquitas S.A., upon commencement to build a mine on these two properties, paid $1,200,000 to the vendors.

The Chinchilla II Mina was acquired directly by VDC and is not subject to option payments.

Concentrates produced at the Project are subject to a maximum 3% ‘mouth of mine value’ royalty that is payable to the Province of Jujuy. This royalty payment is based on the net recoverable value of the contained metals less certain operating costs.

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MPSA and SSR have advised that all necessary permits and titles are in place for the current operations. Additional permitting updates may be required but MPSA advise that these are expected to be approved.

Figure 3.3 Property Map Showing Chinchilla, Chinchilla I and Chinchilla II Concessions

SSR, 2021

3.3.1.1 Chinchillas Surface Rights

MPSA entered into agreements with occupants and owners of the land on which Mina Chinchilla, Mina Chinchilla I, and Mina Chinchilla II are located to acquire the rights to carry out the Project. All of the Minas comprising the Chinchillas property, which provide exploration and exploitation rights, are valid and in good standing.

3.3.1.2 Chinchillas Permitting

According to the biannual Environmental Impact Study renewal, MPSA also submitted the second Update of Chinchillas mine in October 2021. This report is currently being reviewed by Mining Authorities.

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3.3.1.3 Chinchillas Environmental Liabilities

Prior to initiating work on the Chinchilla Mina, an inspection was performed by the mining and environmental authorities regarding potential pre-existing environmental liabilities. There are remnants of historical mining activities in the Project area, such as small buildings, small areas of workings excavated in the 1960’s, historical drilling platforms, trenches and holes. All of these liabilities were declared as pre-existing in Golden Arrow’s ESIA for the Chinchilla Mina, there were no findings and/or requests by the environmental authorities, and the Chinchilla ESIA report was approved.

3.3.1.4 Chinchillas Tenure Factors and Risks

Except as set out herein, to the extent known, there are no additional factors or risks that may affect the access, title, right or ability to perform work on the Chinchillas property.

3.3.2 Pirquitas Mineral Tenure

3.3.2.1 Pirquitas Operation Surface Rights

The Pirquitas Operation includes the surface rights to a group of nine contiguous land parcels covering an area of approximately 7,500 ha, as set out in Table 3.2 and shown in Figure 3.4. This area corresponds to the surface property owned by MPSA, the area of the mining concession is larger.

3.3.2.2 Pirquitas Exploitation Concessions

Mina Pirquitas comprises 54 mining properties (concessions) that cover an area of approximately 9,742 ha shown in Figure 3.4. These parcels were used for purposes such as housing, infrastructure, processing, and tailings facilities. MPSA is the freehold title holder of the area covered by such surface rights.

Table 3.2 Pirquitas Operation Surface Rights

Parcel No.	Registration No.	Area <br>(ha)
531	L-1111	1,000.1
532	L-1112	1,000.0
533	L-1113	750.0
534	L-1114	749.6
535	L-1115	1,000.0
536	L-1116	1,000.0
537	L-1117	1,005.0
538	L-1118	496.0
539	L-1119	500.1

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Figure 3.4 Property Map Showing Pirquitas Concessions

MPSA, 2020

3.3.2.3 Pirquitas Operation Permitting

The capacity of the current tailings facility at Pirquitas has been exhausted and to maintain mining and processing production disposal of tailings has been into the Pirquitas pit. Mining at the Pirquitas pit was completed in January 2017, a number of upgrades have been undertaken to allow tailings to be transported from the Chinchillas Project to a portion of the Pirquitas pit. These upgrades included constructing a pipeline for in-pit disposal, and construction of the discharge system from the tailings transport pipeline, an in-pit water reclaim system, and a pipeline from the Pirquitas pit to the Pirquitas plant to return water for reuse. These upgrades have allowed for additional tailings capacity for the processing of Chinchillas ore.

The use of the Pirquitas pit for tailings deposition at the Pirquitas Operation is a modification to the mining activities not contemplated in MPSA’s ESIA for the Pirquitas mine until 2016. On August 2017 MPSA issued to Mining Authorities an Addendum of the 2016 ESIA Update that included the upgrades to conduct the tailings to the pit of Mina Pirquitas. The permit was obtained on 24 September 2018 by Resolution No. 056/2018. Since then, MPSA has submitted to Mining Authorities the ESIA Update for Mina Pirquitas in September 2020, which is under review.

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3.3.2.4 Pirquitas Tenure Factors and Risks

Except as set out herein, to the extent known, there are no additional factors or risks that may affect the access, title, right or ability to perform work on the Pirquitas property.

3.4 Other Significant Factors and Risks

SSR have advised that there are no other known significant risks that may affect access, title or the right or ability to perform mining related work on the Property.

Legal matters such as statutory and regulatory interpretations affecting the mine plan and environmental matters are outside the expertise of the QPs (see Section 25).

The Puna21TRS QPs consider it reasonable to rely on SSR because SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

Additionally, the project has been in operation for a number of years, and following a review of the current supplied information, the opinion of the Puna21TRS QPs is that the current plans appear adequate to address any issues related to environmental compliance, permitting, and local individuals or groups.

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4 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

Ore from the Chinchillas mine is transported to the Pirquitas plant for processing. The Chinchillas mine is located approximately 45 km from the Pirquitas plant.

4.1 Accessibility

The Chinchillas property is accessed most directly from the provincial capital of San Salvador de Jujuy via National Route No. 9, northwards along the Humahuaca River to the town of Abra Pampa. The route continues along Provincial Route No. 7 westward for 66 km, through the village of Santo Domingo. The roads are maintained by the Province and are accessible year round. Several temporary rivers cross the route so four wheel drive vehicles are recommended in the rainy season.

The other route to the Chinchillas property and to the Pirquitas Operation follows National Route No. 9 northwards from San Salvador de Jujuy to Purmamarca, then turns north-west on paved road No. 52 to the town of Susques. From there, National Route No. 40 heads to Provincial Route No. 70 that leads to Chinchillas at the Fundiciones mountain pass. This route is more appropriate for heavy transport vehicles and is used by traffic to the Pirquitas mine and mill, located approximately 45 km to the south-west of Chinchillas along the route.

Concentrate shipments from Pirquitas are currently trucked to Susques, Jujuy from Pirquitas via Route 77, and from there to Buenos Aires via Route 9. At arrival to the terminal, the material is directly dispatched from the port facilities to the concentrate buyers.

4.2 Physiography, Climate, and Vegetation

The Chinchillas deposit terrain has an elliptical, caldera-like shape with steep rolling hills surrounding the caldera depression. It is located near the Fundiciones mountain pass, with the Rinconada and Carahuasi ranges extending from north–south. Elevations range from about 4,000–4,200 masl. The highest elevation in the area is Cerro Granada at 5,696 masl, 28 km to the south-west. The Uquillayoc river runs through the Project area and is fed by many small tributaries.

At Pirquitas, elevations on the property range from 4,000–4,450 masl. The processing plant, tailings impoundment and main workers camp are located in the eastern third of the Pirquitas property in an area of relatively open ground that lies at an elevation of 4,100 masl, and the Pirquitas pit, which ceased mining operations in January 2017, is situated about 7 km west of the mill at a slightly higher elevation.

The regional climate is similar at both Chinchillas and Pirquitas and is arid to semi-arid, tropical-subtropical influenced by high desert (Blasco, 2011). Rain is scarce and mainly occurs during the rainy season (November to March), with a mean annual precipitation of 300 mm. The annual mean temperature is 18°C, however during the winter it ranges down to –7.7 ºC to 7.5 ºC. Dry and windy conditions often prevail in the area. Natural vegetation is patchy to sparse and consists of xerophilous and steppe bushes like iro (Festuca ortophylia), and coirón (Stipachrysophylla). Acantoliphia haustata is the predominant species with the Yareta (Azorella compacta), less frequent. The tola (Parastrepia ssp.) and small trees like the queñoa (Polylepis tomentella) can be found in depressions (Blasco, 2011).

Animal species found in the area include mammals such as llamas, puna foxes, and vizcachas, as well as several mice species, chinchillas and ferrets. Other fauna in the area include lizards, and birds such as small rheas, owls, ducks, condors, and falcons (Blasco, 2011).

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4.3 Local Resources and Infrastructure

Chinchillas and Pirquitas are located in the rural zone of Rinconada Department, with an approximate population of 2,500 people. It covers an area of 6,407 km2, includes over twenty small communities, and has basic public services including a police department and health centre. The nearest community to Chinchillas is the village of Santo Domingo, and nearest to Pirquitas is the village of Nuevo Pirquitas. Historically, the local population was mainly employed in ranching, however the operation at Pirquitas has created a significant local trained mining workforce. Basic amenities are supplied from Susques and Abra Pampa, while supplies for mining are obtained through the provincial capital of San Salvador de Jujuy, which has an airport with daily commercial air service to Buenos Aires.

The nearest hospital is located in Abra Pampa, 66 km east of Chinchillas.

4.3.1 Chinchillas Infrastructure

The Chinchillas site has offices, workshops, a lunchroom, change room, explosives magazines, security and first aid buildings, solid waste storage facility, open pit, and waste dumps at the site. Existing exploration infrastructure includes two office containers, a core logging facility, a core cutting machine, two storage tents, two cisterns for diesel fuel (1,500 and 10,000 litres) and six warehouses of 144 m2 each, for the storage of the core boxes.

To generate electricity, the Pirquitas Operation uses natural gas to power three Wärtsila generator sets, each with a capacity of five megawatts (MW) of power. In addition, the same electrical plant has three diesel-powered Cummins generators, each yielding 1.1 MW. There is 6.7 km of gas pipeline on the Pirquitas property. The pipeline is 152 mm diameter and constructed of API5L Grade B steel with 4.8 mm wall thickness in normal applications and 7.1 mm wall thickness at river or drainage crossings.

Power for the Chinchillas mine site supplied along existing power lines from the natural gas powered generators at Pirquitas. EJESA is the local power authority that owns the lines. The power line from Pirquitas that goes directly past the rural EJESA line at the town of Nuevo Pirquitas (approximately 5 km from Pirquitas). The rural power line then goes from Nuevo Pirquitas to all villages along Route No. 40 and Route No. 70 and directly to Santo Domingo. This line is able to carry the 1 MW load for Chinchillas, with a small spur line (approximately 4 km in length) to take power into the mine.

No ore processing is done at Chinchillas therefore power requirements are minimal. In the event of power loss at Pirquitas. Back-up power from the EJESA grid that amounts to 100 kVa can be drawn. This back-up power is designated for critical telecommunications systems and the first aid building.

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4.3.2 Pirquitas Infrastructure

Pirquitas has been a permitted commercial mine operated by SSR since December 2009, with existing infrastructure that includes:

•A processing plant;

•A permitted tailings facility;

•A fully serviced workers camp sufficient for approximately 670 personnel;

•A communications system including cellular and intranet access;

•Fully serviced office buildings; and

•Wastewater treatment facilities, organic waste landfill and a recycling centr.e

The Pirquitas processing plant consists of primary, secondary, and tertiary crushing operations which deliver ore to a stockpile. The crushing circuit throughput is 6,000 tpd. Ore is transferred from the crushed ore stockpile to a ball mill and after that a differential flotation circuit to obtain lead / silver and zinc concentrates.

The Pirquitas plant uses a tailings thickener to improve water recovery. Post thickened tailings are deposited in the tailings storage facility and secondary water recovery is achieved using barge mounted reclaim pumps.

MPSA has the surface rights covering the Pirquitas Operation. Electricity is produced from natural gas and diesel generators at the Pirquitas site.

Water supply is from a San Marcos, which is located within the property a short distance downstream from where the Pirquitas River drains into the Collahuaima River. Domestic water is pumped from a diversion upstream of the open pit for use at the camp. Potable water is supplied by MPSA from bottled water.

Pirquitas has a trained workforce for the processing plant and open pit mining operations, including local workers, operators, supervision, management, and senior staff.

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5 HISTORY

5.1 Chinchillas History

Chinchillas was first prospected and mined in small scale in the eighteenth century by Jesuit missionaries. Relics of furnaces used to melt lead and silver can still be found at the Chinchillas property (Kulemeyer, 2011). In 1956, Mr. Antonio Mercado requested a concession based on the discovery of galena veins in the basement rock. In 1968, the mine was sold to Ing. Pichetti, who later formed the Sociedad Pirquihuasi Company together with the Pirquitas Company, and some adits and tunnels were opened for small scale production. In 1982, the mine licence expired, and the mine was acquired by Shell CAPSA S.A. From December 1982 to 1989, a consulting geologist for Shell, Jorge Daroca, carried out exploration work and, after Shell dropped the property, Mr. Daroca requested it for himself, convinced of the good potential of the area (Daroca, undated). Roads, remnants of infrastructure, and minor underground workings remain from this activity, but no records of this work are available.

In 1994, Aranlee Resources conducted surface sampling and drilled seven reverse circulation drillholes for a total of approximately 780 m. Assay results from this work are available, but there are no samples for re-analysis or quality control information, therefore the data have not been incorporated into the Mineral Resources estimate. In 2004 Silex, a subsidiary of Apex Silver, conducted preliminary reconnaissance work including trenching, pitting and surface sampling, with a total of 165 samples taken. Between October 2007 and July 2008, 40 manual pits and nine trenches were sampled. Surface mapping was also completed at different scales across the Chinchillas property, and a total of 1,036 surface samples were collected. At the beginning of 2008, Quantec Geoscience Argentina S.A. (Quantec) performed a 16 km IP resistivity survey, comprising nine sections. The pole-dipole interval was 50 m with 300 m depth readings. The objective of the programme was to detect and delineate sulfides related to an intermediate to high-sulfidation epithermal system, however the mineralised zones at Chinchillas do not appear to be related to chargeability. Nevertheless, there is a strong resistivity contrast between volcanic units and basement schists and the resistivity data have been an effective tool for imaging the volcanic diatreme shape (Quantec, 2008). Silex subsequently drilled 2,220 m in seven diamond drillholes with drillhole samples taken at 1–2 m intervals. Silex had planned to drill 22 holes but cut the programme short during the 2008–2009 global financial crisis. In early 2009 Apex entered Chapter 11 bankruptcy protection, and with a payment due on the property, opted to drop Chinchillas in favour of its more advanced El Quevar project. The core from the Silex drill programme remains at Chinchillas (Silex, 2008 and Caranza and Carlson, 2012).

In 2011, Golden Arrow acquired the property, completed five phases of drilling over the subsequent five years and outlined mineral resources which are summarised in six technical reports and preliminary economic assessments (Davis and Howie 2013, Davis et al., 2014, Davis et al., 2015, Davis et al., 2016, Kuchling et al., 2014, Kuchling et al., 2015). In October 2015 Golden Arrow announced that it had entered into the Agreement with SSR to form a joint venture comprising of the Chinchillas property, the Pirquitas pit and the Pirquitas Operation. The agreement included an 18-month pre-development period to advance Chinchillas, including the infill drilling, engineering and environmental studies, and permitting.

On 18 September 2019, the Company completed the acquisition of the remaining 25% interest in Puna from Golden Arrow Resources Corporation for aggregate consideration totalling approximately $32.4M. The transaction allowed the Company to consolidate ownership in Puna and streamline its reporting.

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5.2 Pirquitas History

Between the 1930s and 1995, the area of the Pirquitas mine had multiple small mining operations to recover silver and tin from placer and vein deposits.

The Argentine branch of Sunshine Mining and Refining Company acquired the Pirquitas mining concessions in November 1995. In the years following its acquisition of Pirquitas, Sunshine Argentina carried out comprehensive mineral exploration on the property, including underground rock sampling and multiple programmes of revere circulation and diamond drilling. These culminated in a feasibility study in February of 2000.

In May 2002, Silver Standard acquired 43.4% of Sunshine Argentina, Inc. (Sunshine Argentina) from Stonehill Capital Management of New York and in October 2004. Silver Standard acquired the remaining 56.6% of Sunshine Argentina from Elliott International L.P., The Liverpool Limited Partnership and Highwood Partners, L.P. Silver Standard operated the Pirquitas mine property as Sunshine Argentina until it changed the company name to Mina Pirquitas, Inc. in May 2008, and further changed the name to MPLLC in December 2014. In August 2018, Mina Pirquitas LLC. changed its name to Mina Pirquitas S.A.(MPSA).

On 24 November 2015, MPSA was incorporated as 1056353 B.C. Ltd., and changed its name to Puna Operations Inc. on 2 May 2017.

Silver Standard approved the start of the Pirquitas mine in October 2006 commenced construction in 2007. The Pirquitas processing plant has been in continuous operation since such date.

The Pirquitas plant has not been expanded since start up; however, minor changes in the flotation flow sheets have occurred to optimise performance. Since 2010, no tin concentrate production has occurred.

Historical records for metal production from the Pirquitas property between 1933 and 1989 indicate that approximately 777,600 kg of silver, or about 25 Moz, along with 18,200 t of tin were recovered by previous operators. An additional 9,100 t of tin was reportedly recovered from the placer deposits found downstream from the lode deposits.

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6 GEOLOGICAL SETTING, MINERALISATION, AND DEPOSIT

6.1 Regional Geology

North-western Argentina consists of three main geological belts or terranes that together trend north–north-east. These are, from east to west, the Sub-Andean Range (Sierras Subandinas), the Eastern Cordillera (Cordillera Oriental), and the Argentine Altiplano or Puna belt.

These belts are distinguished by their basement lithology complexes, tectonic histories, magmatism, metallogeny and geomorphological features. The Pirquitas and Chinchillas deposits are located in the Puna belt.

6.1.1 The Sub-Andean Belt

The Sub-Andean belt comprises multiple north to north-west trending, low mountain ranges separated by broad flatlands. Elevations range from about 300 masl to a maximum of 2,500 masl. An Early Cambrian to Middle Ordovician carbonate platform, which defines a passive continental margin, dominates this belt. Middle to Upper Ordovician clastic marine rocks cover the carbonate platform in the eastern and central sectors. Paleozoic sedimentary successions display regional-scale open folds. Large intrusions and volcanic complexes related to Andean tectonism are not present in this belt. Mineral deposits of economic significance are rare, although natural gas fields are exploited in the eastern lowlands.

6.1.2 The Eastern Cordillera

The Eastern Cordillera is a 70–130 km wide fold and thrust belt with elevations ranging from 1,300–6,200 masl. Proterozoic basement consisting of medium grade metamorphosed sedimentary rocks are unconformably overlain by Paleozoic sedimentary rocks deposited in a back arc basin. The back arc sequence is composed of Early Cambrian to Middle Ordovician clastic marine sedimentary rocks, which in turn are unconformably overlain by Silurian to Devonian sedimentary rocks (Ramos, 2000). The Paleozoic successions are locally covered by Cretaceous sedimentary rocks belonging to the Salta Group.

Late Ordovician to Devonian collision of the composite Arequipa-Antofalla metamorphic basement terrane with the Pampian terrane, which forms the crustal basement in of the majority of north-western Argentina, resulted in folding and faulting of the Paleozoic rocks at Pirquitas (Ramos, 2000). The faults and axial planes related to the large-scale folds formed during this event strike north to north-east. Uplift of structural blocks has exposed elongate, Ordovician-age batholithic granitoid intrusions.

The metallogeny of the Eastern Cordillera is relatively simple. The most important mineral deposit in the belt is the Ordovician age Aguilar sedimentary exhalative (SEDEX) type Pb-Zn(-Ag) deposit, located about 50 km south of Abra Pampa.

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6.1.3 Puna Belt

To the west of the Eastern Cordillera, at elevations of 3,900–6,700 masl, is the Puna belt. The Puna belt consists of nearly the same sedimentary sequences that occur in the Eastern Cordillera. Late Ordovician to Early Devonian compressive tectonism also affected the Paleozoic rocks in the Puna belt, but to a lesser degree than in the Eastern Cordillera. A Paleogene compressive event related to Andean-style tectonics resulted in minor folding and thrust-faulting. By the Late Miocene the tectonic regime transitioned to extension, resulting in basin and range geomorphology. Thinning of the upper crust resulted in the upwelling of magma and the development of andesitic to dacitic stratovolcanoes as well as multiple very large calderas (Figure 6.1). Large volumes of regionally extensive ignimbrite sheets erupted from the calderas, with approximately 1,800–1,200 km3 of material ejected from the Valdema caldera alone (Soler et al., 2007). Sub-aerial volcanism continued into the Pleistocene. This volcanic activity, and associated mineral deposits, was concentrated along corridors defined by lineaments such as Coranzuli Lipez, El Toro Olacapato and Arizaro (Figure 6.2) (Ramos, 1999, Coira et al., 2004, Gorustovich et al., 2011).

Younger rocks include basaltic lavas, continental sedimentary rocks, and the formation of high-altitude salt flats. In terms of mineral deposit endowment, the Puna belt is by far the most important of the three terranes in Jujuy Province. Below are the main deposit types documented in the Puna belt:

•Devonian mesothermal quartz veins and saddle reefs containing native gold, minor base metals and accessory gangue minerals of ankerite and chlorite, with the Rinconada district being the most important for this type of mineralisation.

•Polymetallic quartz-sulphide veins related to eroded Neogene volcanic centers, with the veins containing variable amounts of Pb, Zn, Sb, As, Ag, and Au.

•Bolivian-type Sn-Ag sulphide-rich veins related to Middle to Late Miocene subvolcanic intrusive stocks.

•Pleistocene to recent placer deposits of Au (Rinconada), Sn (Pirquitas) and Au-Cu (Eureka).

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Figure 6.1 Map Showing Tertiary Volcanism from Mega Caldera Complexes Near the Chinchillas Deposit

Note the location of the Chinchillas deposit relative to major faults Modified from Caffe 2002

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Figure 6.2 Oligocene-Miocene Volcanic Arc. Subvolcanic Intrusions

MPSA, 2020

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6.2 Deposit Types

The Chinchillas and Pirquitas deposits are within the Bolivian tin-silver-zinc belt which occupies the back-arc portion of the central Andes and extends from the San Rafael tin-copper deposit in southern Peru to northern Argentina (Figure 6.3). The Bolivian tin-silver deposits are typically associated with felsic volcanic domes of broadly rhyodacitic composition (Cunningham et al., 1991). Bolivian-type Ag-Sn deposits generally consist of sulphide and quartz-sulphide vein systems typically containing cassiterite and a diverse suite of base and trace metals, including Ag in a complex assemblage of sulphide and sulfosalt minerals. The vein systems are generally spatially and likely genetically associated with epizonal (subvolcanic) quartz-bearing peraluminous intrusions one to 2 km in diameter, although the mineralisation may be entirely hosted by the country rocks into which the intrusive stocks were emplaced. The Chinchillas deposit is modelled as a Tertiary-aged diatreme volcanic centre that has intruded Paleozoic sedimentary basement rocks. The mineralisation occurs mostly as disseminations, veinlets, and matrix fill (Figure 6.4).

Most of these deposits depicted in Figure 6.3 are characterised by the intrusion of dacite dome complexes with mineralisation hosted in shear zones and breccia within the dacite domes and/or within shear zones and breccia within the host rocks. At Pulacayo, Potosí and San Cristóbal, where associated domes are present, there is significant mineralisation within the domes. More rarely, as in the case of Chinchillas and San Cristóbal, the deposits include disseminated mineralisation in flat lying manto bodies within sedimentary and pyroclastic rocks. Chinchillas demonstrates phreatomagmatic diatreme morphology associated with a dome structure.

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Figure 6.3 Bolivian Tin-Silver-Zinc Belt with Major Deposits

Golden Arrow, 2013

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Figure 6.4 Schematic Geological Model on West–East Cross-Section showing Chinchillas Deposit

Golden Arrow, 2015

6.3 District Geology

6.3.1 Chinchillas Geology

The Chinchillas silver lead zinc deposit is in the north–north–east trending Puna belt in the western half of Jujuy province and southern part of the Rinconada Range (Figure 6.1). The range has a regional north–north–east trend and is delimited by thrust faults to the west and east. Miocene–age volcanic dome complexes and associated hydrothermal alteration are present in the area, including Cerro Redondo, Pan de Azucar, Rachaite, and the Chinchillas dome complex. High angle faulting and folding also characterise the area. Chinchillas is located within a structural window at the intersection of north-west fracturing associated with the Lipez-Coranzuli regional lineament, the east–west controlling structure, and lesser north-east trending structures.

The Chinchillas deposit is hosted by the Ordovician Acoite Formation and Miocene dacite. The Acoite Formation, described by Board et. al., (2011), is a strongly folded package of low-grade metamorphosed marine sandstone, siltstone and minor shale beds. Deformation of these sedimentary rock occurred during the Ocloyic Phase (Coira et al., 2004) of the late Ordovician. The Acoite Formation is unconformably overlain by Cretaceous marine clastic sedimentary rocks. The Cretaceous sedimentary rocks are overlain by Oligocene to Middle Miocene dacite tuff, continental sedimentary rocks, and volcaniclastic lithologies. The dacitic volcanic centre has an age of 13±1 Ma (Caffe and Coira, 2008) and is a product of a phreatomagmatic diatreme. The resulting topographic depression is elliptical in shape, approximately 2 km long by 1.6 km wide, and infilled with pyroclastic rocks (breccias and tuffs). At the contact between pyroclastic volcanic rocks and basement metasedimentary rocks is a zone of hydraulic fracturing and brecciation up to 150 m wide which is the main host of basement mineralisation. Dacitic lavas, flow domes and subvolcanic intrusions occur on the southern margin of the basin at the contact between metasedimentary and pyroclastic rocks (Figure 6.5).

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Pyroclastic breccias and tuffs erupted from the volcanic centre and filled in the resulting depression, contouring the vent walls. This most likely occurred via airfall deposition and flows of ignimbrites as there is no observed evidence of water-lain deposits or sediments. The breccias and tuffs are mainly matrix-supported and rarely are clast-supported. The clasts are sub-rounded to angular and vary from fine grained to large metre-scale blocks. The clasts are predominantly fragments of re-worked pyroclastic tuffs, lava, dacite, and basement pelite and sandstone. Most of the volcanic clasts and matrix are altered by intense hydrothermal activity, whereas the sedimentary basement clasts are generally better preserved (Figure 6.9).

Three main dacite domes outcrop along the south-east edge of the Chinchillas basin between the pyroclastic breccias and basement contact. The domes have a medium to fine-grained porphyrytic texture with phenocrysts of quartz, (35% to 45%) plagioclase, biotite and minor sanidine (Caffe and Coira, 2008). The dacite domes are generally massive with limited flow banding and some flow brecciation along the margins. Drilling confirms that the dacite outcrops are part of larger bodies below the Socavon del Diablo area. At surface they lie horizontally above tuff breccias.

Figure 6.5 View of the Chinchillas Deposit, Looking East

Note: outcrop of the sedimentary basement rocks, the volcaniclastic sequence infilling the depression, and the dacite domes flanking the southern border of the deposit.

MPSA, 2020

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Figure 6.6 Chinchillas Geology Map

MPSA, 2020

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Figure 6.7 Interbedded Sequence of Marine Sandstone and Pelite with Near-Vertical Dip at Chinchillas

MPSA, 2020

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Figure 6.8 Brecciated Basement Sediments with Fine Volcanic Matrix Near the Contact Between Pyroclastic Sequence and Basement Sediments

MPSA, 2020

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Figure 6.9 Typical Chinchillas Medium Grained Pyroclastic Breccia

With Dacitic Volcanic Clasts Dominant and Secondary Dark Grey Clasts of Basement Sandstone and Pelite

MPSA, 2020

6.3.1.1 Chinchillas Alteration

Typical hydrothermal alteration is described below for basement sedimentary sequences, pyroclastic volcanic rocks and dacite flows.

Alteration in the Marine Sedimentary Basement

In the basement sedimentary sequence mineralisation is restricted to breccias, fracture filling, and veinlets with variable frequency and intensity. Alteration of the host pelite or sandstone is typically very weak, typified by carbonate, clay, and chlorite alteration proximal to sheared structures. Abundant siderite with lesser iron and manganese oxides are observed on fractures. Disseminated diagenetic pyrite is abundant in sedimentary rocks.

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Figure 6.9 Contact Between Dacite Flow Overlaying the Tuff Breccias

MPSA, 2020

Alteration in Pyroclastic Tuffs and Breccias

Pyroclastic tuffs and breccias have undergone several different types of alteration, including clay alteration, sericitisation, silicification, and carbonate alteration primarily as siderite. Clay alteration is most extensive with feldspar, silica and pumiceous fragments altered to various assemblages including quartz–adularia–sericite, illite–quartz-sulfide, and siderite-sphalerite-pyrite. Biotite is commonly altered to sericite-kaolinite-quartz (Caffe, 2013). Extensive fine-grained silicification within the suite of rocks is also documented. Clay alteration, sericitisation and silicification are observed to overprint each other, indicating the alteration event was prolonged and the result of a range of temperature and pressure. Carbonate alteration is locally pervasive and appears late in the paragenesis based on thin section analysis(Marshall and Mustard, 2012). Plagioclase feldspar is commonly replaced by siderite and illite (Caffe, 2013).

Alteration in the Dacitic Domes

Porphyritic dacite rocks were hydrothermally altered to sericite and siderite with minor silicification. Alteration is more developed in the matrix and in the plagioclases crystals (Caffe, 2013).

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6.3.1.2 Chinchillas Mineralisation

Mineralisation at Chinchillas is dominated by silver with lesser amounts of lead and zinc. Mineralisation occurs as disseminated sulfides, matrix infilling within the volcanic tuffs, and as matrix and fracture filling in breccias within the basement metasediments. Dacite volcanic rocks are rarely mineralised in shear zones, veinlets or vein-like structures. Within the basement lithologies shear zones and faults are more commonly mineralised. Depth of oxidation is a few metres within the volcanic rocks and is insignificant within the basement rocks. Silver, lead, and zinc-bearing minerals include silver sulfosalts, freibergite, boulangerite, tetrahedrite, schalenblende, sphalerite, and galena. Associated mineral associations include chalcopyrite, quartz, pyrite, siderite, limonites, manganese oxides, cerussite, smithsonite, anglesite, and malachite (Marshall and Mustard, 2012 and Coira et al., 1993).

The geological model for the Chinchillas deposit includes significant silver-lead-zinc mineralisation in the Silver Mantos and Mantos Basement zones in the western part of the Project (Figure 6.10 and Figure 6.11). Similar mineralisation is present at the adjacent Socavon deposit. A recent review of the Socavon deposit has resulted in its removal from Mineral Resources inventory in 2022.

The main structural elements controlling the location of mineralisation are the contact between basement sediments and overlying volcanic rocks and the dominant east–west and subordinate north-west, north, and north–north-east trending structures that control the Chinchillas volcanic centre. The phreatomagmatic explosion that produced the diatreme generated a symmetrical cylindrical shaped caldera, with mineralised brecciated basement rocks along the contacts and disseminated mineralisation in sub-horizontal tuff layers.

Silver Mantos Mineralisation

Mineralisation is disseminated in several shallow (+/– 5°) dipping layers hosted within clay altered pyroclastic tuffs and breccias. The mineralisation occurs between surface and 100 m depth in sub-horizontal mantos that range between two and 60 m thick, averaging greater than 20 m in thickness. These layers are open for expansion to the east.

Mantos Basement Mineralisation

Located below the Silver Mantos, the Mantos Basement comprises an area 600 m wide and up to 210 m thick, with an average thickness of 80 m, dipping at approximately 40° to the east (Figure 6.100). The zone has been traced down dip approximately 350 m. The Mantos Basement is hosted entirely within basement pelites and sandstones and is comprised predominantly of breccias, crackle breccias with minor small veinlets, and fracture fill.

Socavon del Diablo Mineralisation

The Socavon del Diablo zone (Socavon) is located in the eastern area of the deposit (Figure 6.10). Mineralisation is dominated by manto-style disseminated sulfides within favourable shallow dipping volcanic tuff horizons.

Mineral occurrences, textures, alteration and ore types within the volcaniclastic lithologies are similar to those described for the Silver Mantos target but the mineralisation is thought to be related to a different fluid event based on compositional differences. There may have been a different vent source within the volcanic centre as the Socavon del Diablo mineralisation is generally lower in silver and higher in zinc content.

A recent review of the economics of the Socavon deposit has resulted in its removal from Mineral Resources inventory in 2022.

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Socavon Basement Mineralisation

The Socavon Basement zone is mainly hosted within the Ordovician interbedded pelite and sandstone basement. The east limit of the Socavon del Diablo zone is a dacitic dome intruded in the tuff units and flowed over the tuff at surface (Figure 6.9). Immediately to the east of the dacite dome, biotitic sub-horizontal tuff layers of up to 80 m thick cover the Socavon Basement zone. Here, the mineralisation is hosted in breccias filled with argentiferous galena and a stockwork of sphalerite-siderite-galena within a halo of low-grade zinc of up to 320 m thickness.

The most significant mineralisation in this target is located at more than 150 m depth from surface. The mineralised fluids may have precipitated sulfide minerals as a result of interaction with the water table or decrease in pressure.

A recent review of the economics of the Socavon deposit has resulted in its removal from Mineral Resources inventory in 2022.

Figure 6.10 Silver Mantos and Mantos Basement Zones with Drillhole Locations and Mineralised Zones Projected to Surface

MPSA, 2020

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Figure 6.11 East–West Cross-Section with Deep Manto Mineralisation

MPSA, 2020

6.3.1.3 Chinchillas Resource Expansion Opportunities

Mineralisation at Chinchillas in the Silver Mantos, Mantos Basement, Socavon Basement, and Socavon del Diablo are still open to expansion, particularly the deeper zones of Silver Mantos and Socavon Basement. Other targets include: the northern slope of the basin; the area between the Silver Mantos and Socavon zones; and the dacite domes.

A recent review of the economics of the Socavon deposit has resulted in its removal from Mineral Resources inventory in 2022.

6.3.2 Pirquitas Geology

The majority of the Pirquitas property covers intensely folded Ordovician Acoite Formation marine sedimentary rocks (Figure 6.12). Well exposed along the length of the Pircas River valley, this formation is composed of interbeds fine to medium grained lithic wacke tens of centimetres to a few meters thick, greywacke siltstone, and less abundant black shale that range in thickness from a few centimetres up to several metres. Underlying the north-eastern sector of the property is a sequence of continental sedimentary rocks, mainly hematite-stained arkosic sandstone intercalated with thin polymictic conglomerate beds and cream-coloured reworked dacitic tuff units. This sequence is inferred to belong to the shallow east–north-east dipping Tiomayo Formation of Early to Middle Miocene age. Several kilometres east of the property, a medium-grained granodiorite intrusion forms the small mountain of Cerro Galan, which represents the only substantial intrusive rock body proximal to the mine area.

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Figure 6.12 Pirquitas Geology Map

MPSA, 2020

Folds with shallow-plunging hinge lines and amplitudes ranging from tens of metres to several hundred metres crop out in this area (Figure 6.13). Mining on the north face of the San Miguel open pit has exposed a ‘textbook’ example of a chevron-style anticline (Figure 6.14). High-angle, mostly reverse faults cut the folds, displacing fold limbs by metres to tens of metres.

Axial planar cleavage is well developed in the Paleozoic rocks, especially in the siltstone and shale beds. The well-formed cleavage does not appear to have acted as a receptive structural fabric for quartz-hosted Ag-bearing Fe-Zn-Sn-Pb sulphide veins, although a minor amount of weakly-auriferous quartz veins were deposited along cleavage planes.

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Figure 6.13 Anticline Developed in Interbedded Sandstone, Siltstone and Shale of the Ordovician Acoite Formation, Pircas River Valley, Pirquitas Mine Area

MPSA, 2020

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Figure 6.14 Chevron Fold and High Angle Thrust Fault in the Acoite Formation Host Rocks, North Wall San Miguel Open Pit

MPSA, 2020

The Pirquitas mine open pit exploits previously un-mined portions of the Potosí, San Miguel and Chocoya vein systems. Sheeted sulphide veins and associated disseminated mineralization of the San Miguel system occur in a swarm that is 160 m wide in the north-south direction and a maximum of 400 m along strike in the east–west direction. The Potosí Vein is located on the northern margin of the current pit; the Chocaya Vein system is located on the southern margin and the uppermost part of the Oploca system, known as the Oploca breccia, was exploited by the southern edge of the open pit.

A major system of sulphide-rich veins cut the axial surfaces of the folds and the related cleavage fabric at high angles. Three main and one minor vein sets are recognised at the Pirquitas mine:

Vein Set 1

In the dominant orientation veins strike close to 290° and are generally subvertical. Veins with this orientation include the majority of those in the Potosí, San Miguel, Chocaya, and Oploca areas (Figure 6.15). The Potosí Vein is the largest known single vein on the property, with a strike length of approximately 500 m and maximum thickness of 2.5–3.0 m. Other veins of this orientation typically have a strike length between 100 and 500 m, with average widths of 30–50 cm. The larger of these veins include localised matrix-supported breccias with angular clasts of quartz-sericite altered wallrock in a matrix of Fe and Zn +/– Sn-Ag-Cu sulphides.

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Vein Set 2

The secondary vein set is represented by the Veta Blanca and Colquechaca, veins and narrow (50 cm to 2 m) veins in the Oploca area. The veins are steeply to moderately south dipping and strike close to 310°.

Vein Set 3

The Crucero vein is a series of saddle reefs that follow the axial plane of the antiform in the middle of the Pirquitas pit. Sulphide mineralisation within the Crucero vein is irregularly developed along fractures within white crystalline syn-deformational quartz.

Vein Set 4

At approximately 200 m below surface to the south of the pit is a 4 m thick, 100 x 200 m vein that dips 30° to the north-east and strikes close to 320°. In addition to the veins, zinc-rich mineralisation is hosted within pipe-like breccia bodies that are interpreted to be breccia diatremes.

The Pirquitas open pit exploited previously un-mined portions of the Potosí and San Miguel veins in addition to a set of sheeted sulphide veinlets with associated disseminated mineralisation. The sheeted veins occur in a swarm that is 120–140 m wide in the north–south direction and a maximum of 300 m along strike in the east–west direction. The Potosí Vein is in the northern margin of the current pit; the Chocaya Vein system is south of the open pit (Figure 6.15)

Figure 6.15 Map of the Pircas Valley showing Main Ag-Sn-Zn Vein Systems

Red outline represents approximate limits of San Miguel open pit SSR, 2020

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6.3.2.1 Pirquitas Alteration

Hydrothermal alteration is not particularly well-developed in the host rocks of the Pirquitas deposit (Board et. al., 2011). An assemblage of sericite+quartz+disseminated pyrite replaces original wallrock minerals along the margins of the larger veins, thus forming thin bleached halos to the veins. This sericite-quartz-pyrite alteration is also recognised in wallrock clasts within vein breccia. Disseminated subhedral pyrite is widespread in the deposit, generally constituting less than a few percent of the wallrock by volume; it tends to be more abundant in shale and siltstone beds.

6.3.2.2 Pirquitas Mineralisation

The fracture and breccia-hosted mineralisation at the Pirquitas mine consists of Fe and Zn sulphides with accessory cassiterite (Sn oxide) and a large variety of Ag-Sn-Zn (+/– Pb-Sb-As-Cu-Bi) sulphides and sulfosalts. Crystalline quartz, along with chalcedony in the upper levels of the system, and kaolinite are the main gangue minerals in the veins and mineralised breccias. The main sulphides, specifically pyrite, pyrrhotite, sphalerite and wurtzite, form colloform bands parallel to vein margins, which together with crustiform and drusy vein textures suggest that the mineralisation is epithermal in origin. The vein textures imply that the mineralisation was deposited from relatively low temperature hydrothermal fluids within about 500 m of the paleosurface. However, mineralogical evidence suggests that the initial temperature of the mineralising fluids was possibly greater than 400°C. A detailed study by L. Malvicini (1978) provides relationships between 26 sulphide and sulphosalt mineral phases.

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7 EXPLORATION

7.1 Surficial Exploration

Emphasis was placed on mapping lithologies, alteration and structures to understand the controls of the mineralisation. In the basement rocks bedding, foliation and brecciation were recorded. A handheld x-ray fluorescence (XRF) analyser was used to measure approximate silver, lead, and zinc values at all prospective outcrops.

Geophysical surveys (IP / resistivity, CSAMT, magnetics) conducted in 2013, together with the re-interpretation of the 2008 IP survey, was used to target the Chinchillas South area, detecting deep structure and defining the contact between the tuff unit and basement rocks. The methods used to explore the Chinchillas property are in accordance with industry standards and there are no indications of sample biases.

SSR’s exploration at Pirquitas has predominantly involved RC and DD drilling. Sunshine Argentina completed detailed geological mapping on the property and commissioned approximately 44 line-kilometres of ground magnetics surveying and 19.2 line-kilometres of induced polarisation surveying centred on what is now the San Miguel open pit. Sunshine Argentina’s drilling programmes ended in September 1998, after which the parent company completed an internal pre-feasibility study of the project. Since fully acquiring the project in 2005, the Company has carried out additional geophysical programmes, including in 2012 a 14.4 line-kilometre Quantec Titan-24 DC-IP survey, a ground gravity and differential GPS survey, and in 2018 a Drone Airborne magnetic survey. Between 2008 and 2021 numerous prospecting and geological mapping surveys evaluated the mineral potential of the property.

7.2 Drilling

7.2.1 Chinchillas Summary

The historical drilling programmes at Chinchillas are summarised in Table 7.1. Aranlee Resources completed the first programme in 1994, which comprised seven RC holes. The results from the Aranlee holes were not used in any mineral resource modelling as there is no quality control data.

Table 7.1 Drill Programmes Completed at the Chinchillas Property

Company	BHIDs Sequence	Count	Year	Metres drilled
Aranlee Resources	CH-1–7	7	1994	782
Silex Argentina S.A.	CHD-010–016	7	2007–2008	2,220
Golden Arrow <br>(Phase 1-V)	CGA-017–297	284	2012–2015	45,803
Golden Arrow / SSR <br>(Phase VI – VII)	CGA-212W + CGA-298–340	44	2016	8,945

The average recovery from the 45,803 m of Golden Arrow drilling used in the 2017 Mineral Resources estimate was 94%, including the first 6 m where recovery was commonly less than 50%. Figure 7.1 shows the location of the Golden Arrow drilling separated into six different phases. For details on Chinchillas historical drilling refer to the 2017 technical report, (Kuchling et. al., 2017).

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Figure 7.1 Location of Drillhole Collars at the Chinchillas Deposit

Golden Arrow, 2016

7.2.2 Pirquitas Summary

Initial drilling on the Pirquitas property was conducted by Sunshine Argentina and included a total of 51,863.62 m in 241 drillholes (Table 7.2).

The 2005 drilling programme was designed to test targets in the Oploca, Llallagua, and Colquechaca areas (Table 7.2). The 2007 and 2008 drilling programmes included exploration drilling, resource definition drilling, drilling for metallurgical testing, and condemnation drilling. All drilling was conducted from surface, with the majority completed by RC methods (approximately 84% of the total metreage drilled). Diamond drillholes were generally drilled HQ-size, sequentially reducing to NQ then BQ at depth, as needed.

Diamond core drilling was conducted between July 2010 and September 2011. The majority of this drilling was for resource definition in and around the existing open pit (approximately 89% of the drillholes), with the remainder consisting of exploration drillholes targeting the Cortaderas Breccia Zone (approximately 6% of the drillholes) and other exploration targets (e.g., Veta Blanca).

In 2012, diamond core drilling was conducted between March and November. Most of the drilling was for resource definition in the Cortaderas Breccia Zone (approximately 89% of the drillholes), with the remaining drillholes being exploration drillholes at the pit margins.

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The main objective of the 2018 Potosí drilling programme was to determine if the vein extends eastwards, and to study whether the vein crosses the proposed descending ramp that would access the Cortaderas vein breccia located 500 m to the north. While the main 2–3 m wide Potosi vein does not project into the area, there is a thin and commonly less than 1 m vein intersected in select drillholes.

In 2019–2020, the objective of the deep Granada drill programme was to test the theorised intersection between the south-west dipping Cortaderas vein breccia and the steeply north-dipping Potosi vein beneath the San Miguel pit. Three HQ holes were completed between 6 October 2019 and 31 January 2020 for a total length drilled of 3,430.40 m. Holes from the Granada programme intersected two different mineral compositions that correspond with historically described veins. Mineralisation commonly fills open spaces related to fracturing, brecciation, and faulting, usually as massive or semi-massive veins or veinlets. These veins and veinlets typically range in width from 5–30 cm but may locally approach 3.5 m. The Granada target was not encountered at the anticipated depth, but the programme did identify significant intersections of gold values below the elevation of previous mining within the San Miguel open pit (approximately 4,000 m). The Au grades are commonly associated with elevated concentrations of Ag, As, Bi, Cu, and Sn. The most encouraging broad, low-grade interval from this programme was intersected in GR-396.

As a result of the elevated gold results encountered in the deep drilling in the 2019–2020 Granada drilling programme, select reject and pulp material from historical Cortaderas drilling programmes were re-analysed in 2021 using fire assay gold and multi-element ICP. There were no gold analyses included for any of the original Cortaderas drilling programmes. Two phases of sampling were completed with the first phase including samples from 13 drillholes. The intervals were selected to test a range of high-grade Ag, Zn, and/or Sn intercepts and addition to multiple elevations ranging from the upper portion of the vein (elevation 4,100–4,200 m) to deeper in the breccia system (elevation 3,850–4,000 m). Gold demonstrates a positive correlation between Sb, Cu, Ag, Bi, Mn, and Zn. As a result of the anomalous results, a more-detailed Phase 2 sampling programme was completed that included re-analysis of samples from two drillholes, DDH-214 and DDH-230. An additional 198 rejects were submitted for evaluation and represent continuous intervals (DDH-214: 150 m; DDH-230: 219 m). These results further support an elevation control to the gold mineralisation, with higher grades in the lowermost area of the deposit.

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Table 7.2 Drilling Programmes Completed at the Pirquitas Property

Company		Programme Description		Count		Year		Metres drilled
Sunshine Argentina		San Miguel Deposit (DD)		46		Pre-2004		12,645.72
	Underground (DD)		25		Pre-2004		4,284.50
	San Miguel Deposit (RC)		170		Pre-2004		34,933.40
SSR		Oploca (4), Llallagua (6), Colquechaca (4)		14 *		2005		3,299.65
	San Miguel (24), Cortaderas (6), San Miguel (4), Potosí (1)		35 *		2007		7,723.45
	San Miguel (115), Potosí (52), Oploca (32), Cortaderas (12), Pircas (4), Médanos (10)		225 *		2008		41,112
	San Miguel (38), Oploca (17), Veta Blanca (2), Cortaderas (4)		61 *		2010–2011		12,665.40
	San Miguel (69), Cortaderas (5), Other Targets (5)		79 *		2011		17,549.95
	Cortaderas (126), Médanos (1), West of Pit (9), South of Pit (4), North of Pit (2)		142 *		2012		52,804.30
	Pirquitas Property		17		2013		6,923.00
	Pirquitas Surface (16) and underground (2)		18		2014		3,553.00
	Pirquitas Underground		44		2015		10,961.00
	Potosi – East Extension		15		2018		2,399.30
	Deep Granada		3		2019–2020		3,430.40

* Drillholes used in the 2013 resource modelling

7.3 Drill Core Handling Protocol

The diamond drill core is extracted from the core tube and placed in appropriate boxes marked with drillhole number and the hole depth in metres. The boxes are transported, by pickup truck, from the drill site to the core shack at the end of each shift by trained personnel. The drill contractor used a single shot Reflex survey instrument to measure the down hole deviation. Following completion of the hole, a PVC tube is cemented at the drill collar with hole number, depth, and azimuth inscribed on a metal ticket.

Measurements of core recovery and geotechnical measurements (fracture frequencies and rock quality designation (RQD)) are recorded. The core boxes are then photographed and select intervals are temporarily removed for specific gravity measurements. Geological descriptions are recorded and the samples for analysis are marked at 1 m intervals in mineralised zones and 2 m intervals in areas with no expected mineralisation. The drill core is split using an electric diamond core saw and sampled according to the marked intervals.

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8 SAMPLE PREPARATION, ANALYSES, AND SECURITY

8.1 Sample Preparation, Analyses, and Security – Chinchillas

The following summarises the sample preparation, analysis and security details used by Golden Arrow and SSR in their drill campaigns at Chinchillas and remains unchanged from the most-recent Technical Report (Davis et al, 2016). For details of methodologies used by Silex in the earlier drill campaign, the reader is referred to the penultimate Technical Report (Davis and Howie, 2013).

8.1.1 Sampling Method – Chinchillas

Following the splitting of core, half the core is returned to the box while the other half is bagged. Corresponding tags are inserted, one in the plastic sample bag and the second in the core box. Quality control samples are inserted in sample bags and allocated in order for the laboratory to have a control sample in every batch.

8.1.2 Sample Custody and Security – Chinchillas

Samples bags are placed in larger sacks (between six and ten samples per sack) and are sealed. Sealing numbers are recorded in the Chain of Custody database. The sacks were shipped by private truck to the Alex Stewart (Assayers) Argentina S.A. laboratory in Mendoza, (Alex Stewart) where the sample preparation and analysis are performed.

Samples are received by the laboratory and the reception is reported to Golden Arrow. No damaged or missing samples were ever reported during transportation.

8.1.3 Sample Preparation – Chinchillas

Samples are prepared by method ‘P-5’, which includes drying the samples at 90°C, crushing the entire sample up to 80% passing 10 mesh, splitting 1,000 g with a Jones riffle splitter and pulverising to 95% passing 140 mesh. The pulverised material or pulp is then sampled, and 200 g of pulp is sent to the laboratory.

8.1.4 Sample Analysis – Chinchillas

Alex Stewart was the primary laboratory and ALS in Peru (ALS) was used as the secondary laboratory for check samples (see Section 8.1.6.4 for details). All samples are tested for a suite of 39 elements including silver, lead, and zinc by a four-acid digestion method and analysis by Inductively Coupled Plasma inductively coupled plasmaatomic emission spectroscopy (ICP) (method ICP-MA-39). Silver greater than 200 ppm is assayed by fire assay using a 50 g sample with gravimetric finish (method Ag4A-50). Lead and zinc greater than 10,000 ppm are re-assayed by an oxidising acid digestion for ore grade material and reading by ICP (method ICP-ORE).

In order to speed the reception of assay results, ALS acted as the primary laboratory for one batch of 876 samples in the Phase V programme. Quality control procedures were applied in the same manner as with the rest of the samples.

Alex Stewart is an international laboratory certified under ISO 9001:2008, ISO 17025:2008 and ISO 14001: 2004. Alex Stewart is independent from Golden Arrow and SSR.

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8.1.5 Density – Chinchillas

To determine density, samples of drill core measuring about ten centimetres in length, at approximately 15 m intervals are collected. Samples are dried for two hours at 90°C in an electric oven. After cooling, the samples are sealed with plastic (cellophane) film. The weight of the plastic is ignored in the calculations since the volume is insignificant (less than 1 g of plastic film compared with the 900 g average weight of each sample). The samples are weighed in air and then weighed again while submerged in water. The formula used to calculate density values is as follows:

Density = (Weight in air) / (Weight in water)

A total of 2,586 samples of drill core were tested for density from Phases II, III, IV and V drilling. The results averaged 2.59 t/m3 for the basement rocks, 2.40 t/m3 for the dacites and 2.08 t/m3 for the tuffs, with an overall average of 2.31 t/m3.

8.1.6 Quality Assurance and Quality Control – Chinchillas

Golden Arrow established a quality assurance and quality control (QA/QC) system for its drilling programmes. The system specified the procedures for handling and sampling of drill core including, logging procedures, the frequency of inclusion of QA/QC samples and the procedure for the chain of custody between the drill and the assay lab. QA/QC samples, including blanks and certified reference materials (CRM) are inserted in each batch in the field to check the precision and accuracy of the laboratory. This section reports the results from the Phase V programme. Results from prior phases of drilling are detailed in the previous Technical Reports (Davis & Howie, 2013; Davis et al., 2014, Davis et al., 2015). The QA/QC results from previous drilling programmes indicate the samples from those programmes are of sufficient quality to support Mineral Resources estimation.

A total of 1,792 quality control samples were inserted as shown in Table 8.1.

Table 8.1 Summary of QA/QC Samples – Chinchillas

Type of Sample	Number of Samples	Percentage of Total<br>(%)
Core samples	10,468	85.4
Coarse Blanks	369	3.0
Fine Blanks	377	3.1
Coarse Duplicates Lab. 1	185	1.5
Fine Duplicates Lab. 1	191	1.6
Fine Duplicates Lab. 2	293	2.4
Reference Material	377	3.1
Total	12,260	100

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8.1.6.1 Blanks – Chinchillas

Coarse and fine blanks were used to detect contamination problems and cross labelling in the process. The blank used was not a certified material from a vendor. The coarse blank, named BL-CH-1G, was made from a tuff breccia with no silver mineralisation and low-grade base metals values. It was sampled by Golden Arrow personnel and assayed by Alex Stewart Assayers.

The blank material used for QA/QC purposes was not certified by a round robin process at several accredited laboratories; however, assay QA/QC results indicate the material appears to be sufficiently homogeneous to detect sample contamination. The acceptance values were three times the reference value. In the case of the Ag the acceptance value was three times the detection limit (1.5 ppm Ag).

From the 369 coarse blank samples, all Ag, Pb, and Zn values are under the acceptance limit except for one sample with 221 ppm Pb, just above the limit of 198 ppm Pb.

The fine blanks were made from the fine rejects of coarse blanks of the previous drilling phase. They were named BL-CH-2F, BL-CH-2aF and BL-CH-3F. The original assays were averaged, and internal reports were produced. The acceptance values were three times the reference value. During the Phase V drilling programme, a total of 377 fine blanks were inserted in the batches as part of the QA/QC programme. Ag values were always below the acceptance limit of 1.5 ppm Ag. Lead and zinc values were also below the acceptance limit except for two outliers in Pb and Zn. These outliers might reflect some contamination in the laboratory but the absolute values, even above the acceptance limit, are not considered significant.

8.1.6.2 Coarse and Fine Duplicates – Chinchillas

During the Phase V drill programme coarse and fine duplicates were incorporated in the quality control process. A total of 185 of the coarse rejects (at 10 mesh) were re-labelled with a new number, re-assayed at Alex Stewart and considered as coarse duplicates. The same procedure was applied to 191 fine rejects (pulps), and these were considered as fine duplicates. Assay of the fine duplicates is not intended to validate the assay process since each part of the duplicate pair was assayed in the same laboratory. Pairs of values below 3 ppm Ag were removed due to the poor precision of results. Figure 8.1 shows a summary of the coarse and fine duplicates for Ag comparing the mean percentage difference (MPD) to the accumulated MPD. The MPD is calculated as the percentage of Ix1-x2I / (x1+x2) / 2.

Curves for Pb and Zn show similar tendency as for Ag.

Field duplicates were not taken during the Phase V drill programme. As shown in previous phases, the comparison between quarter-core versus half- core had low representativeness and usefulness.

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Figure 8.1 Ag Values for the Two Types of Duplicates – Chinchillas

SSR, 2020

8.1.6.3 Certified Reference Materials – Chinchillas

A set of certified reference materials (CRM) standards was used to check the accuracy and precision of the laboratory. The same three CRMs used during Phases III and IV were used during the Phase V programme, referred to as 1-CH, 2-CH and 3-CH. These standards were originally prepared by ACME-Mendoza, at the request of Golden Arrow, from rejects of previous drill core from the Chinchillas property. CRMs 1-CH and 2-CH have low (41 ppm) and intermediate (146 ppm) Ag grades and were packaged in 30 g envelopes because they do not require fire assay. Standard 3-CH has higher silver content (862 ppm) and, therefore, was packaged in 120 g envelopes to accommodate the larger sample requirements of the fire assay testing.

A total of 148 CRM of 1-CH, 157 of 2-CH, and 72 of 3-CH were inserted along the Phase V drilling. The assay results from the 1-CH all fall within three standard deviations (SD)of the accepted value (Figure 8.2). In the case of the 2-CH, only one value is above three SD of the accepted value. The results of 3-CH, shown in Figure 8.3, indicate that all assay results are within two SD of the accepted value.

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Figure 8.2 Ag Values from CRM 1-CH and 2-CH – Chinchillas

MPSA, 2020

Figure 8.3 Ag Values from CRM 3-CH – Chinchillas

MPSA, 2020

The results for lead, shown in Figure 8.4, indicates some outliers in standard 2-CH. Samples immediately before and after this potentially suspect standard result were re-assayed and no significant difference was detected from the original assays.

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Figure 8.4 Pb Values from CRM 1-CH, 2-CH and 3-CH – Chinchillas

MPSA, 2020

In the case of zinc, reference materials 2-CH and 3-CH were assayed by method ICP-MA and all values are within +/– two SD, except for one sample that is less than three SD from the accepted value (Figure 8.5 and Figure 8.6).

Figure 8.5 Zn Values from CRM 1-CH and 3-CH – Chinchillas

MPSA, 2020

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Figure 8.6 Zn Values from CRM 2-CH – Chinchillas

MPSA, 2020

8.1.6.4 Secondary Laboratory for Checks – Chinchillas

ALS was used as secondary laboratory. A total of 293 pulps were sent to ALS to be tested by method ME-ICP61 based on a four-acid digestion and reading by ICP. Samples greater than 1% Pb or 1% Zn were re-tested using ore grade method Pb-OG62 and Zn-OG62. Samples greater than 100 ppm Ag were re-assayed by fire assay with gravimetric finish (method Ag-GRA22). ALS is part of an international laboratory system and has ISO 9001:2008 and 17025:2005 certifications. ALS is independent from Golden Arrow and SSR.

As with the field / coarse duplicates, the laboratory duplicate pairs with values close to the lower limit of detection were removed due to the poor precision of results, leaving only the greater than 3 ppm Ag values.

Figure 8.7 shows the MPD of the Ag, Pb, and Zn values in check samples between the primary and secondary laboratory.

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Figure 8.7 Comparison of Ag, Pb, and Zn Laboratory Duplicates – Chinchillas

MPSA, 2020

8.1.7 Conclusions and Recommendations – Chinchillas

In the opinion of the QPs the sample preparation, security, and analytical procedures meet or exceed industry standards for data quality and integrity. There are no factors related to sampling or sample preparation that would materially impact the accuracy or reliability of the samples or the assay results. The outcomes of the QA/QC procedures indicate that the assay results are within acceptable levels of accuracy and precision and the resulting database is sufficient to support the estimation of Mineral Resources.

8.2 Sample Preparation, Analyses, and Security – Pirquitas

The following summarises the sample preparation, analysis and security details used by Sunshine Argentina and SSR in their drill campaigns at Pirquitas and remains unchanged from the most-recent NI 43-101 Technical Report (Board et al, 2011).

8.2.1 Sampling Method – Pirquitas

8.2.1.1 Sunshine Argentina

RC drillhole cuttings were collected and split into 30–40 kg samples at the drill rig. A three-tier Jones-style splitter was used to split these samples. A 3–5 kg sample was sent to the relevant analytical laboratory for sample preparation and analysis.

Drillhole core (HQ and NQ) was marked for sampling and cut in half using a diamond saw. One half of the core was geologically logged and stored on site. The other half of the core was sent to the laboratory for sample preparation and analysis.

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A total of 2,788 underground channel samples were collected over a total distance of 1,600 m from mineralised veins and sheeted vein systems in the main ore zone of the San Miguel zone, as well as from the Oploca, Potosí, Blanca, San Pedro, and Llallagua vein systems. Samples of approximately 2 kg per linear metre were chiselled from channels.

8.2.1.2 SSR

Dry RC samples were split to approximately one-eighth (~12.5%) of the original sample size using a three-tier Jones-style splitter, then split to one-sixteenth using a one-tier Jones splitter. This sample was sent to the laboratory. Wet RC samples were initially halved using a wet splitter, with one half split again to one-eighth using a three-tier Jones splitter, then halved again to produce a one-sixteenth sample to be sent to the laboratory.

Drillhole core was marked for sampling and cut in half using a diamond saw. One half of the core was geologically logged and stored on site. The other half of the core was sent to the laboratory for sample preparation and analysis. All drillhole samples generated from the 2010–2011 drilling programmes were diamond drillhole core samples.

8.2.2 Sample Custody and Security – Pirquitas

The analytical laboratories took possession of the samples at the Pirquitas site, and the samples were in their custody throughout the sample preparation and analysis steps, including sample transportation from site to the respective analytical laboratory.

SSRs sampling protocol included the labelling of sample bags and closing with a security seal. The samples were then sent to Jujuy by company truck.

8.2.3 Sample Preparation – Pirquitas

8.2.3.1 Sunshine Argentina

Sunshine Argentina’s drilling programme was effectively conducted in two phases, with the transition being marked by a change in analytical laboratories from American Assay Laboratories (AAL) to the SGS Chile laboratory partway through its drilling programme. RC drillholes AR 001-AR 092 and diamond core drillholes DDH 001-DDH 042 were analysed by AAL, RC drillholes AR 093-AR 164 and diamond core drillholes DDH 043-DDH 069 were analysed by SGS Chile.

Sample preparation procedures were similar at both analytical laboratories:

•Samples were initially dried for two to three hours at 105°C.

•Dried samples were crushed to less than 18 mm in diameter using a jaw crusher, through to less than 2 mm to less than 0.18 mm in diameter using a roll crusher.

•A Jones-style riffle splitter was used to collect sample splits of approximately 250 g (AAL) and 400 g (SGS Chile).

•Sample splits were pulverised in ring / disk pulverisers to less than 0.10 mm in diameter, homogenised, and packaged for analysis.

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All coarse rejects from the AAL prepared sample splits were stored on-site at Pirquitas; a minimum of 0.25 kg per sample was returned for on-site storage at Pirquitas by SGS Chile. A split of each sample pulp was also returned for on-site storage at Pirquitas.

8.2.3.2 SSR

RC and DD samples were shipped to the ALS Chemex analytical laboratory in Mendoza, Argentina. The following sample preparation was conducted by ALS Chemex:

•Samples were logged into the ALS Chemex Webtrieve sample tracking system (ALS Chemex procedure LOG-21), weighed (WEI-21), and then dried (DRY-21).

•Dried samples were crushed to between 70% and 80% passing a nominal –2 mm (CRU-31 or CRU-35), and split using a riffle splitter (SPL-21) to produce a representative 250 g split for pulverisation. The sample split was pulverised to better than 85% passing 75 µm (PUL-31 or PUL-32, depending on sample size).

8.2.4 Sample Analysis – Pirquitas

8.2.4.1 Sunshine Argentina

Sample pulps were digested in aqua regia and analysed for Ag using atomic absorption spectrometry (AAS). Samples with values higher than 500 ppm Ag were analysed a second time using fire assay methods. For Sn analyses, the sample pulps were fused with sodium peroxide and caustic pellets to ensure the Sn was completely dissolved before being analysed by AAS.

A total of six assay laboratories were used during Sunshine’s two drilling phases:

•Phase I – After sample preparation, AAL sent the samples to the Laboratorio Quimíco Guayacan Ltda. Analytical laboratory in La Serena, Chile for Ag analysis, and to the AAL analytical laboratory in Santiago, Chile for Sn analysis. Samples were also submitted to the Centro de Investigación Minera y Metalúrgica (CIMM) in Santiago, Chile for check assaying of Ag, and to the Instituto de Investigaciónes Minero-Metalúrgicas in Oruro, Bolivia for check assaying of Sn.

•Phase II – Prepared samples were sent to the SGS Chile analytical laboratory in Quilicura, Santiago, Chile for assaying, and to the Acme Labs in Santiago, Chile analytical laboratory for check assaying purposes. The analytical laboratories received 60 g pulps for Ag analyses and 20 g pulps for Sn analyses.

8.2.4.2 SSR

The analytical methodology changed during SSRs 2005–2008 drilling programme. Samples were initially analysed using the ICP mass spectrometry method, then aqua regia digestion followed by 36 element atomic emission ICP spectroscopy (ME-ICP41). Ag grades were found to be understated by both the ICP mass spectrometry method and, to a lesser degree, the ICP agua regia method. As a result of this SSR elected to switch to a third method: Four-acid ‘near total’ digestion followed by 34 element atomic emission ICP spectroscopy (ME-ICP61a, including Sn). Over limit Pb (>10%), Zn (>10%), and Ag (>200 ppm) grades were re-analysed using a four-acid digestion followed by AAS finish (Pb, Zn, or Ag-AA62 procedures). Ag grades still over limit (>1,500 ppm) were analysed by fire assay with a gravimetric finish (Ag-GRA21). Additional Sn analyses were conducted using AAS (Sn-AA82). All ICP mass spectrometry samples were re-assayed using this method by ALS Chemex.

Four-acid ‘near total’ digestion followed by 34 element atomic emission ICP spectroscopy (ME-ICP61a, including Sn) was the primary analytical technique used during the 2010–2011 drilling programme.

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8.2.5 Density – Pirquitas

The method of density determination is not specifically discussed in historical documentation.

8.2.6 Quality Assurance and Quality Control – Pirquitas

8.2.6.1 Sunshine Argentina

Approximately 12% of the samples submitted for assays were QA/QC samples consisting of field standard, blank, and duplicate control samples. The QA/QC results showed that:

•Overall, Ag analyses of CRM and blank control samples were within acceptable limits.

•Field duplicate control samples of Ag were considered acceptable.

•Sn analyses were initially biased low, resulting in the re-assaying of 3,252 samples. No significant biases were noted in the Sn assay data for the Phase II drilling. Limited cross-contamination in Sn assay data was rectified through a programme of sample batch re-assaying. Sn data displayed a relatively high degree of inherent variability.

8.2.6.2 CRM

CRM standard, blank, and field duplicate control samples were inserted into the sample stream on a one-in-twenty basis. Approximately 5% of the total number of submitted samples was submitted to the third-party analytical laboratory for check assaying. QA/QC samples included six different reference CRMs covering a representative range of Ag, Sn, and Zn grades, blanks generated from barren sandstone, and field duplicates (prepared as discussed above).

The QA/QC results showed that:

•The control values of the CRMs were not initially correctly calibrated, resulting in extensive failures of the field standard control samples relative to no failures in the analytical laboratory standard control samples. Recalibration of these values indicates that the key assay data from the 2005 through 2008 drilling programmes are unbiased and accurate.

•Field blank control samples indicated that sample cross-contamination was generally not an issue during the analytical work conducted on SSRs 2005–2008 drilling data.

•Field duplicate control samples, whilst indicating a degree of variability in the assay data, were reported at acceptable levels of precision for Ag, Sn, and Zn, given the nugget effect (inherent variability) and the variability associated with quarter-core versus half-core samples.

8.2.7 Conclusions and Recommendations – Pirquitas

In the opinion of the QPs the sample preparation, security, and analytical procedures meets industry standards for data quality and integrity. There are no factors related to sampling or sample preparation that would materially impact the accuracy or reliability of the samples or the assay results. The outcomes of the QA/QC procedures indicate that the assay results are within acceptable levels of accuracy and precision and the resulting database is sufficient to support the estimation of Mineral Resources.

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9 DATA VERIFICATION

9.1 Database Validation

9.1.1 Collar Coordinate Validation

9.1.1.1 Collar Coordinate Validation – Chinchillas

Validation of collar elevation data at Chinchillas was done by comparing elevations from DGPS field surveys against the satellite photo digital elevation model (DEM). Precision of the DGPS is between 15–70 cm.

9.1.1.2 Collar Coordinate Validation – Pirquitas

Drillhole collar locations at Pirquitas were validated by an independent surveyor for the 2011 resource modelling study.

In the 2013 modelling study, it became apparent that there was a discrepancy in some pre-2009 holes in the form of displacement of mineralised vein intervals relative to the vein interpretation and grade control data. An example of this issue is shown for drillhole AR-315, which has a high-grade interval that falls outside of the Veta Blanca vein model (Figure 9.1). The vein model in this location has been informed by many grade control drillholes, causing the position of AR-315 to be called into question.

A thorough investigation was undertaken, and similar issues were identified in 96 drillholes.

Efforts were made to identify the possible source of the issue and, if a transposition error, remedy those errors definitively, however the age of the data and the inability to re-survey the collars due to them having been mined made this unachievable.

To remedy the issue in the modelling, the collar locations of the affected holes were adjusted to bring the vein intercept into expected location, making it concordant with observations in surrounding holes. While this is not an ideal situation, the intervals in question are likely to be captured by grade control at the time of mining, therefore it is not expected to cause a volumetric difference. However, further ongoing assessment would be required to definitively identify and remedy the incorrect historical data. Some new drilling should be considered to re-check vein locations in the areas afflicted by suspect holes.

9.1.2 Downhole Survey Validation

The down-hole survey data were validated by searching for large discrepancies between the dip and azimuth reading against the previous reading. No significant discrepancies were found.

Before the beginning of Phase III drilling at Chinchillas it was noted that the correction of the magnetic declination between true north and magnetic north was correct in angle but had the opposite direction. For this reason, all azimuths of drillholes of Phases I and II were corrected by 13° counterclockwise. No other adjustments were necessary for the other drilling phases.

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Figure 9.1 Vein model with Exploration Drillhole AR-315 Outside of the Vein

MPSA, 2013

9.1.3 Assay Verification

To validate the Chinchillas data, the following checks were confirmed:

•The maximum depth of samples were checked against hole depth;

•The values of less than the detection limit were converted into a positive number half the detection limit;

•The highest Ag values and at least one random value from each drillhole were checked against the original assay certificate;

•The units were converted from ppm into percent (%) for Pb and Zn values; and

•Silex drillhole assay data were validated as reported in a previous Mineral Resources estimate (Davis et al, 2013).

For Pirquitas, approximately 10% of the pre-2010 drilling assay data set was checked and compared to the original assay certificates, to generate additional confidence in this data. Detailed checks of assay data from the 2010–2011 drilling programme was undertaken, with iterative corrections made for any anomalies (generally typographic errors, including mis-labelled samples, and mis-labelled sample intervals).

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9.2 QA/QC Protocol

A review of the Chinchillas QA/QC protocols was conducted prior to drilling and formalised in a detailed QA/QC manual developed by Golden Arrow. Onsite reviews were conducted during all drilling phases by a QP. The procedures for core processing, the insertion of blanks and standards were examined and considered appropriate.

At Pirquitas, QA/QC information for all exploration drilling programmes was analysed. Review of real-time QA/QC data monitoring was undertaken by SSR, especially timing and effectiveness of remedial action taken with respect to failed batches.

9.3 Geological Data Verification and Interpretation

While several geology variables were captured during core logging, only lithology was used to constrain the Chinchillas Mineral Resources estimation. Therefore, geology data verification was limited to determining that the lithology designation was correct in each sample interval. This included the following:

•FROM – TO intervals for gaps, overlaps and duplicated intervals;

•Collar and SampleID mismatches; and

•Correct geology codes.

A geological legend was provided by Golden Arrow and compared to the values logged in the database. Data were examined on screen for discrepancies in logging.

9.4 Assay Database Verification

The assay data from 15 randomly selected drillholes, representing approximately 5% of the Chinchillas database, was manually compared to the original assay certificates. These holes contained a total of 1,890 individual samples, in which eight samples were found to have differences in the values of the second decimal value. Differences of this nature are not considered to be ‘errors’ as they have no measurable impact on the estimation of Mineral Resources. The results of this test indicate the database is sound and free of errors.

For Pirquitas, approximately 10% of the pre-2010 drilling assay data set was checked and compared to the original assay certificates. Detailed checks of assay data from the 2010–2011 drilling programme was undertaken.

9.5 QP Opinion

It is the opinion of the QPs that the data is adequate for the purposes used in the Puna21TRS. No material sample bias was identified during the review of the drill data and assays. Review of the data validation processes indicates that the drill data is adequate for the estimation of Mineral Resources.

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10 MINERAL PROCESSING AND METALLURGICAL TESTING

10.1 Chinchillas

The metallurgical development of Chinchillas ore types commenced in 2013 and continued through 2016. The first testwork was focussed on silver recovery by both leaching and flotation methods, with flotation proving to be superior at the early stage. The second programme continued process development of flotation into separate lead / silver and zinc concentrates. The third testwork campaign was designed to advance the flotation process and test specifically these ore types to the Pirquitas mill flow sheet.

10.1.1 Initial Testwork 2013

A scoping metallurgical test programme was initiated in January 2013. This testwork was undertaken by Bureau Veritas Commodities Canada Ltd. All testing was bench-scale. Results from the early testwork stages are summarised in a previous NI 43-101 Technical Report (Kuchling et al, 2014).

10.1.2 Second Phase Testwork 2014

The second testing programme was conducted on composite samples from the silver Mantos zone (MAN-2), the Socavon Del Diablo Zone (SOC-2) and the Mantos Basement zone (BAS-1). This programme included locked cycle testing and provided the most representative view of the overall metallurgical performance of the samples to date. The following summary is an excerpt from the final report titled “2014 Project Report on Metallurgical Testing on the Chinchillas Project” prepared by Bureau Veritas Commodities Canada Ltd, Inspectorate Metallurgical Division, (Chen and Redfearn, 2014):

“Seven core samples (received on October 15, 2013 weighing 102 kg), were air dried and separated into three composites. Each composite was individually crushed to 6 mesh, mixed and split into the required samples for testing. Silver contents range from 94.2–150.6 g/t and base metals include lead and zinc.

In this testing programme, it was confirmed that Chinchillas samples are usually amenable to the conventional lead and zinc sequential flotation process. For most of the samples, the majority of silver was recovered in the lead circuit. Overall silver, lead, and zinc recoveries are above 95%. Most rougher concentrates responded well to the subsequent cleaner flotation stages. Upgrading of composites BAS-1, MAN-2, and SOC-2 generated lead final concentrates with grades ranging from 65% to 79% lead and zinc final concentrates with grades from 52% to 62% zinc.”

Locked cycle tests on three samples (BAS-1, MAN-2, and SOC-2) showed that high silver and lead recoveries in the lead circuit can be achieved along with good lead final concentrate grades. For composites BAS-1 and SOC-2, good final zinc concentrates grading 51.8% and 60.1% respectively were obtained.

To assist with future metallurgical development, mineralogical analysis was undertaken on the three ore types (BAS, MAN, SOC) and two flotation testwork concentrates (BAS lead second cleaner concentrate and lead scavenger concentrate generated during one of the flotation tests).

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The report concluded:

“The three composites assayed 100–150 g/t silver and 0.6% to 2.2% lead. Freibergite was the dominant silver bearing mineral, constituting over 75% of the total feed silver. The remaining silver was contained in pyrargyrite, stephanite and tetrahedrite. The lead was mostly contained in galena.

The three composites also assayed 70–300 g/t copper and 130–330 g/t arsenic. The copper was predominantly carried by freibergite and chalcopyrite.

The arsenic was mostly carried by arsenopyrite and krutovite.”

The objective of this second phase flotation testwork was to produce sequential lead / silver and zinc concentrates. This was successful with high recoveries achieved of the target metals to marketable quality concentrates. The mineralogical analysis highlighted that the lead was contained in galena, and the silver was contained in the very typical series of silver sulfosalt minerals.

10.1.3 Third Phase Testwork 2016

The 2016 flotation testing programme was developed to determine the compatibility of Chinchillas mineralisation types to the Pirquitas process plant flow sheet and capacity. Testwork included comminution and focused on producing lead / silver and zinc concentrates by sequential flotation. In addition, a comparison between the flotation reagent scheme used in the historical testwork programmes and the current Pirquitas scheme was undertaken.

The testwork was completed at ALS Metallurgy, Kamloops, British Columbia, Canada.

10.1.3.1 Selection of Drill Intervals for Testing

A review of the drill assay database assays was used to imply mineralogy; specifically, iron to sulfur ratio (Fe:S, a proxy for pyrite content). It was suggested at the start of the testwork programme that silver might be partially associated with pyrite. A typical example of both silver content and Fe:S versus drillhole depth is shown for drillhole CGA-35 in Figure 10.1. However, the varying iron to sulfur ratio appeared to be independent of Ag grade – therefore, a poor association with pyrite.

The criteria for selection of individual core intervals for selection for metallurgical testing were:

•Within pit shell (excluding the SOC zone, not in the initial mine plan)

•Ag grades similar to mine plan grades

•Fe:S ratio into High and Low classes

•Lithology into either Manto or Basement

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Figure 10.1 Drillhole CGA, Variation in Ag Grade and Fe:S Ratio Downhole

MPSA, 2020

Nominally four separate drillhole intervals were identified for each of the four mineralisation types. These were named ‘Manto Low’ and ‘Manto High’, and ‘Basement Low’ and ‘Basement High’ with the designations corresponding to Fe:S ranges of >15 or <5 respectively. Figure 10.2 shows the selected drill interval locations within the pit. The pit is planned in two mining phases, the first shown as the red pit shell and the second as the green pit shell.

These identified intervals were recovered from the Chinchillas site drill core library and re-sawn into quarter core by Golden Arrow geological staff. Once securely bagged and labelled, approximately 350 kg of material was shipped directly to the laboratory in Kamloops, Canada.

In addition to the economic metals, additional analysis was completed for lead and zinc oxides, total and sulfide sulfur and silver (by both fire assay and three-acid ICP methods).

Observations included:

•Low amounts of lead and zinc oxide with no effect expected on flotation.

•High proportion of the total sulfur is present as sulfide (i.e., limited sulfates).

•Variation of silver by the two methods is low which implies most silver is sulfide hosted.

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Figure 10.2 Metallurgical Sample Locations within the Two Pit Shells (Mining Phases)

MPSA, 2020

The Master composites ranged in grade between 154–238 g/t Ag, 0.31%–1.77% Pb and 0.16%–1.02% Zn. The Fe:S ratio ranged from 5–30. In terms of composite grades, the selection of samples was based on an initial mine plan. This initial plan had no mining in Socavon zones and therefore, no samples were selected from this Project area for the 2016 flotation testwork programme. Variability samples are selected to cover a range of grades above and below this mine plan.

Metallurgical testwork development followed a general plan of:

•comminution testing;

•reagent optimisation on the four Master composites;

•batch rougher / cleaner flotation on Master and Variability composites;

•locked cycle flotation on the four Master composites; and

•additional flotation tailings were generated for thickening tests and water chemistry.

10.1.3.2 Comminution

Two of the Master composites, Basement Low and Manto Low, and individual composite CGA-89 Manto High, were tested for Bond Work Index (BWi) values. For comparison, the Pirquitas plant design was 15.2 kWh/t (Table 10.1).

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Table 10.1 Bond Ball Mill Work Index Test Results

Composite Sample	BWi, kWh/t
Basement Low	11.5
Manto Low	15.5
Manto High (CGA-89)	16.2

10.1.3.3 Master Composite Rougher Flotation

The previous metallurgical programme in 2013 utilised a flotation reagent scheme quite different from the standard Pirquitas flotation reagent scheme. The initial series of batch sequential rougher flotation tests were performed on the four Master composites testing these two alternate reagent schemes. Neither of these schemes utilised sodium cyanide for pyrite or sphalerite depression.

Primary grind was maintained in the target P80 size range of 120–160 µm, consistent with both previous testwork and Pirquitas operating experience on similar ore types.

The Pirquitas reagent scheme recovered more silver to the lead concentrate. For Basement Low and High samples, the increase in silver recovery to the lead / silver concentrates was 3.6% and 11.8%. For Manto Low and High, the increase in silver recovery to the lead / silver concentrates was 19.6% and 28.7%. Therefore, the Pirquitas reagent scheme was used for all subsequent flotation testing (both batch rougher / cleaner and locked cycle work).

10.1.3.4 Master Composite Rougher / Cleaner Flotation

For each of the four Master composites, a rougher / regrind / cleaner test was completed, yielding separate lead and zinc concentrates.

For all Master composites, a high Pb grade lead concentrate was produced, with the contained Ag grade varying directly with the lead to silver proportion in the heads. Open circuit cleaning recovery was good. For the very low zinc grade Manto High composite, no zinc flotation was attempted. The remaining three Master composites produced marketable zinc concentrates.

10.1.3.5 Variability Composite Rougher / Cleaner Flotation

For each of the Variability composites, a rougher / cleaner flotation test was completed to assess the effect of head grade variation on metal recoveries and cleaner concentrate grades.

Pirquitas’ operating experience has demonstrated difficulty in achieving a marketable grade zinc concentrate when zinc feed grades are below 0.4% Zn. For the Chinchillas variability testwork, no zinc flotation was completed for any composite with a head grade below 0.2% Zn.

As with lead / silver flotation, there is generally consistent flotation performance between the Master and the Variability composites.

This reagent scheme employed at the Pirquitas plant, using ZnSO4, lime, AP3418A and MIBC, avoids the use of cyanide in the lead flotation stage, thus eliminating any cyanide concerns with tailings effluent and the possible need for cyanide destruction.

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10.2 Socavon / Chinchilla

10.2.1 Testwork 2018

In 2018, ALS Metallurgy, Argentina conducted preliminary metallurgical testwork on Socavon / Chinchilla and Pirquitas Manto / Basement composites.

The Socavon / Chinchillas composites were tested by flotation to determine if a lead / silver and a marketable zinc concentrate could be produced. Additionally, to access the blend results of the Socavon / Chinchilla and Pirquitas Manto / Basement composites. Two metallurgical test composites were constructed representative of the Socavon zone of the Chinchillas deposit. The composites were identified as Socavon / Chinchillas Composite A and Socavon / Chinchillas Composite D. An additional composite was constructed for blend testing and was identified as Pirquitas Basement / Manto composite.

The primary objective of the test programme was to determine the effect of blending the Socavon / Chinchillas composites with a Pirquitas Manto / Basement composite. Lead feed grade measured 0.56% and 0.78% in Socavon / Chinchillas Composite A and D, respectively. The Pirquitas Basement / Manto composite measured the highest lead grade at 0.96%. Lead soluble in ammonium acetate digestion, indicative of lead in oxide form measured about 20% of the total lead in the Socavon / Chinchillas Composite A.

Zinc feed grade measured 0.56% and 1.78% for Socavon / Chinchillas Composite A and D, respectively. The Pirquitas Basement / Manto composite measured approximately 0.76% Zn. Zinc soluble in an ammonium acetate digestion measured 0.06% or less indicating a low percent of the zinc as non-sulphide minerals. Silver feed grade measured 16–34 g/t Ag for the Socavon / Chinchillas Composites A and D, respectively, and measured significantly higher in the Pirquitas Basement / Manto composite at 190 g/t Ag. Sulphur content in the samples ranged from 0.7%–4.4% S.

Initial cleaner testing on the Socavon / Chinchillas Composites A and D indicated that approximately 71% and 87% of the lead could be recovered to lead concentrates grading 55% and 30% respectively. Lower lead recoveries measured in the Socavon / Chinchillas Composite A test was likely due to the high percentage of the lead in the feed associated with lead oxide minerals. About 87% and 83% of the silver was also recovered to the respective lead concentrates. The low lead grade of the Composite D concentrate was due the high levels of zinc dilution in the lead concentrate. Optical microscope assessment of the lead concentrate indicated that the majority of the zinc minerals were well liberated, and rejection should be possible with the correct flotation chemistry.

The Pirquitas Basement / Manto composite was tested using similar test conditions and recovered 94% of the lead and 92% of the silver to the lead cleaner concentrate which graded about 60% lead and 1.1g/t Ag. Zinc recovery to the lead concentrate was about 20%, similar to the Socavon / Chinchillas Composite A.

A series of blend tests were conducted using the Socavon / Chinchillas and Pirquitas composites. Results from these tests indicated that lead and silver recoveries to the lead concentrates typically decreased when higher ratios of the Socavon / Chinchillas composites were added to the various blends. The lead grade in the Socavon / Chinchillas Composite D blend tests decreased as more of the composite was used in the blends, as a result of increased zinc in the lead concentrate. This trend was also observed for zinc in the zinc circuit.

Measured lead and zinc recoveries in the blend tests were typically lower than the calculated recoveries of the various blend ratios. However, the measured concentrate grades for the blend tests were higher than the calculated concentrate grades and it is likely they exist on a similar grade recovery curve. Zinc recoveries to the lead concentrates were about 5% lower than calculated values. This suggests that there was room for improvement within the lead circuit in the baseline Socavon / Chinchillas tests.

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Tests were conducted using MBS as a replacement for zinc sulphate and the results indicated that the addition of MBS at the dosages employed had no significant effect on depressing zinc in the lead circuit. Additionally, closed circuit testing would be beneficial for providing a better estimate of metallurgical performance in the zinc circuit.

Preliminary relationships, Table 10.2, of recovery and mass pull to produce a silver / lead concentrate and zinc concentrate for Socavon material have been derived.

Table 10.2 Preliminary Socavon Recovery Relationships

				Unit		Silver / Lead Concentrate		Zinc Concentrate
Recovery		Silver		%		90		2.5
	Lead		%		55 x Pb% feed + 45		1
	Zinc		%		(Pb/Zn feed x –46) + 56		Zn feed x 13.8 + 21.5
Mass Pull to concentrate		Concentrate		t conc./<br>t feed		Pb% feed x 0.055 – 0.02		Zn% feed x 0.004 + 0.001

Further optimisation flotation testwork is recommended including liberation testing, mineralogy and flotation reagent optimisation.

10.3 Metallurgical Performance Estimates

The Pirquitas process plant operating performance since commencement on Chinchillas ores is used to provide the concentrate grade recovery and mass pull relationships, Table 10.3 and Table 10.4.

Table 10.3 Silver / Lead Concentrate Relationships

Variable	Variable Formula
Ag Recovery	(–0.0631 x Pb recovery2) + (11.655 x Pb recovery) -447.4
Pb Recovery	(–2.6303x Pb Feed2) + (12.329 x Pb Feed) + 80.654
Zn Recovery	(-5.2817 x Zn Feed2)+(Zn Feed x –6.31) + 20.546
Mass Pull	(–0.0024 x Pb Feed2) + (0.0164 x Pb Feed)+-0.0007

Table 10.4 Zinc Concentrate Relationships

Variable	Variable Formula
Ag recovery	(–3.4843 x Zn feed2) + (7.2499 x Zn feed)+0.8295
Pb recovery	(0.024 x (Pb feed / Zn feed)2) + (-0.5988 x (Pb feed / Zn feed)+ 3.1292
Zn recovery	(–195921 x (mass pull Zn)2 + (5620.3 x mass pull Zn)+28.709
Zn recovery
Mass Pull	(0.007 x Zn feed2) + (0.0041 x Zn feed+0.0011

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10.4 Recommendations for Additional Testwork

It was identified that low zinc grade material would not generate a saleable zinc concentrate and the zinc circuit would be bypassed when Zn grade is less than 0.4%.

The Chinchillas mineralogy showed lead occurred predominantly as galena, silver as a series of sulfosalts and zinc as sphalerite.

Additional metallurgical laboratory testwork should include the following:

•Detailed geometallurgical study to understand the distribution of possible future smelter penalty elements (e.g., antimony for lead concentrate and silica for zinc concentrate).

•Additional Bond Work and Abrasion Index testing on samples throughout the deposit.

10.5 QP Opinion

It is the opinion of the QPs that the data is adequate for the purposes used in the Puna21TRS and the analytical procedures used in the analysis are of conventional industry practice.

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11 MINERAL RESOURCES ESTIMATES

11.1 Mineral Resources Estimate – Chinchillas

The Chinchillas Mineral Resources estimate was developed in August 2020 by independent consultant company Red Pennant Geoscience. The Puna21TRS QPs have reviewed and accepted this information for use in the Puna21TRS. The Chinchillas Mineral Resources have been estimated in accordance with generally accepted industry guidelines and are reported in accordance with S-K 1300. Mineral Resources are not Mineral Reserves and they do not have demonstrated economic viability.

The effective date of the resource cell model is 28 August 2020. The effective date of the Mineral Resources is 31 December 2021 after accounting for depletion from mining from the August 2020 model.

The previous Mineral Resources estimate for the Chinchillas property had an effective date of 12 April 2016 and is described in the NI 43-101 Technical Report dated 27 May 2016, (Davis, et al., 2016).

11.1.1 Available Data – Chinchillas

The database available at the time of the resource modelling comprised a total of 335 diamond drill (DD) holes with 56,641 m of logged data and 55,905 m of assay data.

The spatial distribution of the drilling is shown in Figure 11.1.

Figure 11.1 Isometric View showing the Chinchillas Drillhole Database used in Resource Modelling

MPSA, 2020

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11.1.2 Compositing – Chinchillas

Raw sample lengths were generally up to 2 m in waste rock and 1 m or less in mineralised rock. A minimum sample length of 0.1 m was permitted on samples from highly mineralised structures such as veins, stockworks, and breccias.

A composite length of 5 m was considered most suitable for the Chinchillas drillhole data. Data were composited to the selected composite length within the interpreted wireframe solid. Residual lengths were retained.

11.1.3 Exploratory Data Analysis – Chinchillas

Exploratory data analysis was conducted to understand the distribution of the metals within the different lithological units and as well in different structural domains.

Five estimation domains were used to inform the estimation process, the lithological data along with the rock geochemistry informed these estimation domains. The raw and composited statistics for silver, lead and zinc are provided in Table 11.1.

Table 11.1 Estimation Domain Statistics

	Count	Min.	Max.	Mean	SD	CV
Domain 1
Ag (g/t)	3,270	0.25	2,908.62	81.70	181.23	2.22
Pb (%)	3,270	–	29.64	0.79	1.63	2.07
Zn (%)	3,270	–	8.99	0.29	0.59	2.04
Domain 2
Ag (g/t)	3,635	0.25	8,970.34	72.71	211.79	2.91
Pb (%)	3,635	–	28.00	0.69	1.37	1.98
Zn (%)	3,635	0.01	6.10	0.39	0.60	1.52
Domain 3
Ag (g/t)	5,444	0.25	5,395.45	87.77	238.53	2.72
Pb (%)	5,444	–	15.38	0.62	1.02	1.66
Zn (%)	5,444	–	11.90	0.74	1.19	1.60
Domain 4
Ag (g/t)	3,061	0.25	2,466.14	47.68	112.85	2.37
Pb (%)	3,061	–	12.44	0.34	0.72	2.14
Zn (%)	3,061	–	13.08	0.21	0.57	2.72
Domain LO
Ag (g/t)	20,821	0.25	2,466.14	47.68	112.85	2.37
Pb (%)	20,821	–	10.90	0.07	0.15	2.10
Zn (%)	20,821	–	15.31	0.15	0.38	2.52

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The majority of the samples were analysed by ICP for a suite of 39 elements. The silver, lead, zinc, and sulfur data was extracted from the main database for use in the development of the resource model.

The database contains a total of 2,586 samples that have been tested for density. These samples were obtained from core selected at approximately 15 m intervals down most drillholes giving a relatively consistent distribution of density data throughout the deposit areas.

Individual assay sample intervals ranged from 0.1–10 m and averaged 1.34 m in length. Some 72% of the samples were exactly 1 m in length and 25% of the samples are 2 m long. Values analysed below the detection limit (<DL) were assigned values equal to one half of the detection limit (½DL).

Diamond drill core recovery averages 96%. Recoveries do not vary significantly between rock types (average recoveries: tuff 95%, dacite 98%, basement breccia 97% and basement 97%). There was no indication of a relationship between core recovery and grade.

11.1.4 Evaluation of Outlier Grades – Chinchillas

The 5m composited data were examined for outliers using cumulative probability plots and for loss metal of due to capping of higher-grade values. Multiple scenarios were run to understand the effect of capping and compared with the tonnes and grade within the mined out areas. The capping shown in Table 11.2 was used in the estimation domains.

Table 11.2 Estimation Domain Capping

	Domain 1	Domain 2	Domain 3	Domain 4	Domain LO
Ag (g/t)	900	1,000.00	1,300.00	843.50	12.60
Pb (%)	8	7.90	6.30	5.00	1.00
Zn (%)	3	2.50	5.00	2.50	1.50

11.1.5 Geological Model – Chinchillas

As described in Section 6.2, the Chinchillas deposit is interpreted to be formed as a result of a Tertiary aged diatreme intrusion into a host of Paleozoic basement schists. Heat from the intrusion resulted in mineralisation in the form of disseminations, veinlets, and matrix filling within the volcanic breccias and tuffs as well as within the original schists.

11.1.5.1 Lithological Model – Chinchillas

The general spatial distribution of the main lithological units at Chinchillas is shown in cross-section in Figure 11.2. The higher grade silver-lead-zinc mineralisation occurs predominantly in the tuffaceous phase of the intrusive rocks and also within the brecciated zone in the underlying basement schists. However, relatively high-grade mineralisation can be found in all rock types.

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Figure 11.2 Cross-Section showing Rock Types and Silver Grades in Drilling

MPSA, 2020

The mineralisation in the Mantos area of the deposit exhibits two general styles or trends; a more flat-lying mantos-style distribution which is more common in the tuffs and a second basement trend of mineralisation which tends to be sub-parallel to the basement / tuff contact.

The comprehensive logging of lithological types (20) and alteration style (6) and intensity (5) results in the potential for 600 combinations. The lithological and alteration codes were rationalised into a small number of units for practical purposes.

A simplified 3D implicit model was created of the key lithological units (Figure 11.3):

•‘so’ (surficial materials)

•‘dac’ (dacite intrusive)

•combined ‘ctb’, ‘ftb’ and ‘mtb’ tuff and breccia units

•Ss2 basement

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Figure 11.3 Isometric View of Simplified Lithology Model

MPSA, 2020

11.1.5.2 Structural Model – Chinchillas

High resolution satellite imagery and SRTM and client-supplied topography data were combined into a 3D model. As the outcrop patterns are not obscured by vegetation, it was possible to carry out a ‘virtual’ field mapping exercise to measure the strike and dip of the locality (Figure 11.4). These measurements were used to develop a structural model in the form of structural surfaces to aid interpretation of the morphology and depth extension of the diatreme.

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Figure 11.4 Bedding Attitudes Represented as Disks Mapped on Topographic DTM

MPSA, 2020

11.1.5.3 Multi-Element Geochemistry Model – Chinchillas

An implicit geochemistry-based model was constructed using K-means clusters defined by assay statistics.

A clustering algorithm was used to partition the assay dataset into distinct, exclusive clusters so that the data points within each group show as little variability as possible. The reduction in variability within each of the clusters should help to minimise the coefficient of variation and improve geostatistical estimation within each domain by improving the likelihood of geostatistical stationarity. The advantage of the clustering process is that the variability can be automatically and quantifiably assessed by a computer algorithm much more efficiently than by human visual processing.

The elements used for the clustering exercise were: Ag, Al, As, Ba, Ca, Co, Cr, Fe, Ga, K, La, Li, Mg, Mn, Mo, Na, Nb, Ni, P, Pb, S, Sr, Ti, V, Y, Zn, and Zr.

Generally, the way K-means algorithms proceeds via an iterative refinement process, as follows:

•Each data point is randomly assigned to a cluster (number of clusters is pre-determined),

•The centroid of each cluster (the mean within the cluster) is calculated in n-dimensional space, and

•Each data point is assigned to its nearest centroid (iteratively to minimise the within-cluster variation).

Two cluster cases were assessed: four-cluster and seven-cluster solutions. The seven-cluster solution was geometrically complex and was not pursued further.

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The pattern of the four-cluster solution was simpler than the seven-cluster solution, conformed to the general geometry of known geological features, and provided only 14% more variability. Consequently, it was decided to proceed with the simpler four-cluster solution. The cluster allocation was coded onto the drillhole data (Figure 11.5). A 3D model was made of the four-cluster geochemical model using known geometry, including the trends of the diatreme and associated lithological units. Clusters 1 and 2 were largely confined to the exterior of the diatreme, while clusters 3 and 4 show a flatter geometry interior to the steep west-plunging diatreme.

Figure 11.5 K-means Four-Cluster Multi-Element Geochemistry Model

MPSA, 2020

11.1.5.4 Indicator Probability Constraint – Chinchillas

To provide a limit to estimation and to partly fulfil the requirement of ‘reasonable prospects of eventual economic extraction’, an indicator interpolant was created. Based on experimentation, a 15 g/t Ag value and a 30% probability were used to provide the limiting probability shell. Experimental indicator variograms were modelled and used in the indicator estimation.

The resulting indicator shells conformed to the geometry of the mineralised zones and were used to restrict the four-cluster domains to better mineralised regions (Figure 11.6).

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Figure 11.6 Volumes with >30% Probability of Exceeding 15 g/t Ag

MPSA, 2020

The cluster model combined with the probability shell provided the final estimation domain framework (Figure 11.7).

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Figure 11.7 Estimation Domains based on K-Means Clusters of Multi-Element Geochemistry

MPSA, 2020

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11.1.6 Grade Estimation – Chinchillas

Estimation was carried out within a cell model with 16 m x 16 m x 10 m cells (Table 11.3). Sub-cells of 4 m x 4 m x 5 m were permitted to improve boundary resolution. The model is not rotated.

The cell model was subsequently split down into 8 m x 8 m x 5 m cell sizes for mine planning purposes.

Table 11.3 Cell Model Limits – Chinchillas

Direction	Minimum	Maximum	Cell Size<br>(m)	Number of Cells
East	3,472,100	3,474,404	8	288
North	7,510,644	7,512,996	8	294
Elevation	3,750	4,300	5	110

Cells in the model were coded on a majority basis with the various domains.

The proportion of cells that occur below the topographic surface are also calculated and stored in the model as individual percentage items. These values are used as weighting factors when determining the in situ Mineral Resources for the deposit.

Estimation was undertaken within the four-cluster domains as well as outside of those domains, with only like-coded samples permitted to inform the estimates. Grades within peripheral unestimated blocks and the rock dumps were set to zero.

Three methods were used to populate Ag, Pb, Zn, and S estimates into the cell model:

•Nearest neighbour (NN),

•Inverse distance to the power of two (ID2), and

•Ordinary kriging (OK).

Normal scores variography was modelled and a locally varying orientation was used for both the ID2 and OK estimates. The varying directions follow the generally centrally dipping pattern seen in the geological modelling of the flat Mantos and steep Socavon marginal zones.

The OK grade estimates are regarded as definitive, while the ID2 and NN estimates were used for validation purposes.

11.1.7 Density – Chinchillas

Density was estimated from the previous PFS model using a nearest neighbour approach.

Density was estimated in the 2016 using the ID2 method. Densities are estimated with a maximum of two composites per drillhole and a maximum of six composites in total. The lithology domains provide hard boundary conditions during estimation and samples below 1.75 t/m3 excluded as these are considered to be anomalous.

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11.1.8 Validation – Chinchillas

Grade-tonnage comparisons of the ID2 vs. OK estimates were undertaken to validate the estimates (see Figure 11.8). In general, the ID2 and OK results are similar.

A cross-section showing the conformity between exploration drillhole data and the model estimates is shown in Figure 11.9.

Figure 11.8 Chinchillas Grade Tonnage Comparison within Pit 3 Volume – Ag, Zn, and P

MPSA, 2020

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Figure 11.9 Chinchillas Cross-Section at 7,512,400 mN showing Ag Composites and Estimates

MPSA, 2020

11.1.9 Classification – Chinchillas

Classification was undertaking in accordance with the same method used in the 2016 resource model; that being minimum and maximum distance from drillhole data. The classification criteria used are shown in Table 11.4. Model cells that sit within the economic pit shell and meet the classification criteria are reported as Mineral Resource.

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Table 11.4 Classification Parameters – Chinchillas

Class Description		Class Number		Average Distance to Assay from Three Drillholes
	Minimum		Maximum
Measured		1		0	25
Indicated		2		25	50
Inferred		3		50	75

11.2 Pirquitas

11.2.1 Available Data – Pirquitas

The 2011 Mineral Resources estimate database contains assay data derived from DD and RC drillholes. The finalised valid drillhole database used as input for the modelling contains 551 collars (326 DDH and 225 RC).

The spatial distribution of the drilling completed to date at Pirquitas is shown in Figure 11.1.

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Figure 11.10 Isometric View showing the Pirquitas Drillhole Database used in Resource Modelling

MPSA, 2011

11.2.2 Exploratory Data Analysis – Pirquitas

The 2013 modelling dataset contains 14,246 downhole survey, and 88,220 sample records, of which 88,788 have associated Ag assay data (for a total metreage of 108,815.6 m of Ag assays).

The majority of the samples were analysed by four-acid ‘near total’ digestion followed by 34-element atomic emission ICP spectroscopy (ME-ICP61a, including Sn).

Individual assay sample intervals in the database ranged from 0.001–5.0 m and averaged 1.31 m in length. Values analysed below the detection limit (DL) were assigned values equal to one half of the DL. The basic statistical summary of the assay sample data used in the 2013 resource modelling at Pirquitas is shown in Table 11.5.

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Table 11.5 Statistical Summary of Raw Assay Data Used in 2013 Modelling – Pirquitas

Element	Count	Min.	Max.	Mean	SD	CV
Ag (g/t)	82,788	0.10	41,330	43.45	341.09	7.85
Zn (%)	82,787	0.001	36.6	0.93	2.34	2.51

Drilling recovery generally ranged between 95% and 100% for diamond core drillholes, and generally between 80% and 100% for RC drillholes, except where drillholes intersected old underground workings.

11.2.3 Domain Interpretations – Pirquitas

SSR considers the Pirquitas deposit as comprising the Mining Area, which includes the San Miguel, Potosi, and Oploca Vein zone, and the Cortaderas area, which consists of the Cortaderas Breccia zone and the Cortaderas Valley zone (see Figure 11.10). It is the Mining Area that is the subject area for the 2013 resource modelling.

From July to October 2013, wireframes of the majority of mappable veins within the San Miguel, Potosi, and Oploca areas were created in MineSight software using a 65 ppm Ag-equivalent (AgEq) cut-off to define vein margins. The AgEq is calculated using the following formula:

AgEq = Ag (ppm) + 14 x Zn%

Visual analysis of colour-coded uncomposited grade control bench assay data was conducted to assess the general orientation of the veins within the domains. Figure 11.12 shows the available grade control data, colour coded for AgEq. Clear trends of veins are obvious at this scale, the 010° trend follows the orientation of an anticline hinge where there is a concentration of mineralisation where it is intersected by the more common north-west trending veins.

For the majority of veins, the process of wireframe construction began with the snapping of points on what were determined to be hangingwall and footwall pierce points of veins on the exploration drillholes. This was done in 3D space on a vein by vein basis. A different methodology was applied to the central part of the Crucero vein where it crosses the San Miguel pit and the Potosi Breccia. For these models, bench polygons were made around >65 g/t AgEq grade intercepts and then linked to form wireframes.

The wireframes were trimmed at surface where they were not controlled by exploration drilling to ensure that they did not extend above topography or overburden material. They were also intersected where one vein crossed or touched another in order to reduce the possibility of coding errors.

The result of this process was that the Mining Area was broadly sub-domained into three zones:

•Central San Miguel zone – characterised in the field by essentially sub-vertical veins and vein stockworks.

•Northern Potosí zone – characterised in the field by veins and vein stockworks steeply dipping toward the north.

•Southern Oploca Vein zone – characterised in the field by veins steeply dipping toward the south.

Details of the domain coding used in the generation of the 2013 model are shown in Figure 11.11.

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Figure 11.11 Pirquitas Plan showing Drillholes Coded using Wireframe Vein Models

MPSA, 2013

Table 11.6 Pirquitas Mineralisation Domain Codes

Location		Zone		Mineralisation Domain Code
Mining Area		San Miguel		10
	Potosí		20
	Oploca		30

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Figure 11.12 Plan showing Grade Control Bench Assays

MPSA, 2013

Wireframe models for Pirquitas are presented as an isometric view inside the September 2013 EOM survey pickup and 2013 designed pit in Figure 4.3.

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Figure 11.13 Isometric View of Final Pirquitas Wireframe Vein Models – 2013 Resource Model

MPSA, 2013

After the wireframes were created the ‘LITHO’ field in the drillhole file was coded for the domains and vein wireframes. Each intercept on different veins was given a unique identifier representing a ‘material type’. This assignment of material type codes by vein allowed for the control of grade estimation within the veins. The drillhole traces back-coded for vein material types are shown in Figure 11.11.

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11.2.4 Compositing – Pirquitas

A composite length of 2 m was considered most suitable for the Pirquitas drillhole data. Data were composited to the selected composite length within the interpreted wireframe solid. Residual lengths were retained.

11.2.5 Evaluation of Outlier Grades – Pirquitas

Top cut analyses were conducted on all finalised domain-coded drillhole composite data to assess the potential impact of extreme values during grade estimation.

The process of assessing the need for top cuts was an iterative one and included the analysis of statistics including lognormal probability distribution plots and the reconciliation of estimated parent cell grades to grade control. Ultimately, the best reconciliation to grade control was obtained by applying a 9,000 g/t Ag upper cut and a 20% Zn upper cut.

11.2.6 Continuity Analysis – Pirquitas

A variographic analysis was undertaken to help define the continuity characteristics of the domained data.

Three dimensional variography analysis (pairwise relative) was undertaken on 2 m composite intervals for Ag and Zn. Downhole variograms (omni-directional) were used to determine the nugget effect.

Due to the low amount of data within each vein wireframe the general orientation and shape of the veins or groups of veins (San Miguel and Oploca) was used to determine the orientation and dimensions of the search parameters for grade interpolation. Thus, the variogram parameters from the variogram model, that is the nugget, first, and second structure were the same for all kriging interpolations for both Ag and Zn, the search ellipse orientation and dimensions.

11.2.7 Grades Estimation – Pirquitas

Estimation was carried out within a cell model with 4 m x 4 m x 8 m cells to be compatible with the grade control model. The Cell model limits are shown in Table 11.7. The model is not rotated.

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Table 11.7 Cell Model Limits – Pirquitas

Direction	Minimum	Maximum	Cell Size<br>(m)	Number of Cells
East	751,900	753,340	4	360
North	7,488,300	7,490,340	4	510
Elevation	3,838	4,510	8	84

Cells in the model were coded with the various domains on a majority basis.

The wireframe surfaces and solids representing topography, oxide surface, overburden, and vein interpretation were used to code each cell with the proportion of its volume relative to these features.

11.2.7.1 Grades Estimation – Pirquitas Mining Area

Ag and Zn grades were estimated into Domains 10, 20, and 30 in the cell model using ordinary kriging (OK). The vein boundaries were treated as hard boundaries such that the composites within the vein interpretations were only permitted to inform estimates in cells that fall within the veins (this estimate is stored in a field called ‘AGPV’ for Ag and ‘ZN%V’ for the Zn estimates), and likewise composites outside the veins were only permitted to inform estimates in cells that have some proportion outside of the veins (this estimate is stored in a field called ‘AGPVD’ for Ag estimates and ‘ZNPVD’ for Zn estimates).

Veins in the San Miguel and Oploca areas were estimated separately. The veins in the Oploca area were separated into two sets for estimation: Oploca Set A and Set B veins have generally different orientations these are shown in Figure 11.14.

Figure 11.14 Pirquitas Cross-Section showing Two Distinct Oploca Vein Sets

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MPSA, 2013

Whole cell grades were calculated after estimation completed into fields called ‘AGP’ for silver and ‘ZN%’ for zinc. The calculation takes into account the proportion of the cell that is inside / outside the vein by calculating the weighted average based on stored volume information.

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11.2.7.2 Density – Pirquitas

An analysis of density data was undertaken. It was determined that density could be estimated satisfactorily using a formula underpinned by Ag grade as follows:

agp = agpv * veinp / 100 + agpvd * (100 – veinp) / 100

temp = 0.39 * (ALOG (agp + 10) / ALOG (10)-1) + 2.64 * (1 – oxide) + 1.8 * oxide

bulk density = temp * (1 – void% / 100)

Where:

AGPV = estimated Ag grade from sample inside a vein

VEINP = proportion of the cell that falls inside a vein wireframe

AGPVD = estimated Ag grade from sample outside the veins

OXIDE = proportion of cell in oxide material

VOID% = proportion of cell not in rock (i.e., in underground workings)

2.64 = the average density of fresh waste rock

1.8 = the average density of oxidised rock

11.2.7.3 Validation – Pirquitas

Validation was conducted at the end of each model preparation step, as well as of the completion of the process. Model validation included visual validation in plan and section, and statistical validation by domain, of cell model codes. Additional validation was conducted on cells coded with UGVOID% values of greater than 0.001% to ensure that the in situ volume estimate was correctly discounted for previously mined material.

In addition to the ongoing and iterative validation steps conducted throughout the modelling process, the resultant model that forms the basis of the Mineral Resources estimate was subjected to the following validation steps:

•Visual comparison of estimated grades for Ag and Zn against the input drillhole data on a series of plan views and oblique cross-sections through the model. This was reviewed at numerous times during model generation to ensure that the modelled grades and grade distribution closely reflected that of the input data.

•Comparison of average grades for each grade variable in each domain between the input drillhole data and model estimates, to assess for potential global estimation bias in the model.

•Comparison of average grades for each grade variable in each domain between the input drillhole data and the cell model along northing, easting, and elevation swathes to assess potential spatial bias in the model.

•Grade-tonnage data were reported using different software to ensure validity of final grade-tonnage reports.

•Reconciliation of the cell model to grade control and production data on a bench-by-bench basis in the mined-out part of the deposit.

Results of the various detailed model validation steps summarised above indicate that the 2013 grade estimates are honour the input geological and drillhole data both globally and locally and have an acceptable level of smoothing for the cell size selected.

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11.2.7.4 Mineral Resources Classification – Pirquitas

Following an assessment to determine suitability, the resource confidence classification method for the 2013 model used the same method described in the 2011 NI 43-101 Technical Report. The following is a description of the 2011 analysis.

Mineral Resources classification was conducted using a combination of drillhole spacing, search volume, distance from underground workings, mineralisation continuity considerations, comparisons in locations of high-grade vein and stockwork structures between the model and open-pit observations, reconciliation between the model and grade control and production data, and discussions with mine-based geological staff. Resource confidence classification was focussed on the Ag variable, being the most economically important constituent at the Pirquitas deposit, however, prior to finalisation, the classification was reviewed with respect to Zn and Sn to ensure that there were no anomalies. Care was taken to ensure that coherent zones of high-confidence (Measured Mineral Resources) and reasonably good confidence (Indicated Mineral Resources) estimates were modelled to avoid a scattered resource classification. To this end the various model classification criteria discussed above were incorporated in the generation of wireframe solids for the classification of the Mineral Resources into Measured and Indicated Mineral Resources.

The Mineral Resources confidence classification scheme can be broadly simplified as follows:

•All estimates informed by the third search pass in the Mining Area (Domains 10, 20, & 30) were classified as Inferred Mineral Resource.

•All estimates informed by sufficient samples from drillholes spaced closer than effectively 40 m (generally 20–35 m), in areas where there is reasonably good confidence in the modelled location of the mineralised veins and for which there is reasonably good confidence in the modelled location of the underground workings, were classified as Indicated Mineral Resources.

•Estimates classified as Measured Mineral Resources were informed by sufficient samples from drillholes spaced closer than effectively 25 m (generally less than 15 m), in areas for which there is high confidence in the modelled location of the mineralised veins and for which there is high confidence in the modelled location of the underground workings.

The wireframes used for classification in 2011 were added to the 2013 cell model. The ranges of distance to the cell centroid from the nearest composite, number of holes and number of composites relative to the 2011 resource cell model are given in Table 8.1.

Table 11.8 Pirquitas 2013 Mineral Resources Classification Results, Valid at a Cut-off of 65 g/t AgEq

Mineral Resources Classification	Average Distance	Average No. of Holes	Average No. of Composites
Measured	17.44	7	21
Indicated	29.70	6	6
Inferred	51.85	2	4

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11.3 Mineral Resources Estimate

11.3.1 Chinchillas Mineral Resources

The Chinchillas Mineral Resources estimate is contained within a pit shell generated using an NSR cut-off of $33.20/t that is based on metal prices of $22.00/oz for silver, $0.90/lb lead, and $1.15/lb for zinc.

Metal prices for the Mineral Resources cut-off were selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The metal prices are representative of the range of price estimates publicly reported for Mineral Resources cut-offs. The Chinchillas Mineral Resources are assumed to be mined by open pit.

In determining the cut-off grade, the reasonable prospects for eventual economic extraction requirement generally implies that the quantity and grade estimates meet certain economic thresholds taking into account an open pit extraction scenario with road transport and processing at the Pirquitas plant. This includes consideration of the technical and economic parameters listed above, but also includes additional operating costs, estimated at $12/t, related to the handling and transportation of ore from the Chinchillas property to the Pirquitas plant. Using this operating scenario, the cut-off grade is estimated to be 60 g/t AgEq. It should be noted that this determination considers site operating costs and ignores the pay factors for any concentrate generated and sold to a smelter.

MPSA have advised that there are no known factors related to environmental, permitting, legal, title, taxation, socio-economic, marketing, or political issues that could materially affect the Mineral Resources estimate. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues. The quantity and grade of reported Inferred Mineral Resources are uncertain in nature and there has been insufficient exploration to classify these Inferred Mineral Resources as Indicated or Measured Mineral Resources. It cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resources as a result of continued exploration.

Mineral Resources are reported exclusive of Mineral Reserves.

The Chinchillas Mineral Resources by classification are summarised in Table 11.9 and shown with recovery and ownership detail in Table 11.10.

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Table 11.9 Summary of Chinchillas Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $22.00/oz Silver, $0.95/lb Lead, and $1.15/lb Zinc

Mineral Resources Classification		Tonnage<br>(kt)		Grades						Contained Metal
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(koz)		Lead<br>(klb)		Zinc<br>(klb)
Measured		1,110		99.2		0.86		0.31		3,540		21,015	7,552
Indicated		4,904		101.1		0.88		0.19		15,943		95,632	20,454
Measured + Indicated		6,013		100.8		0.88		0.21		19,483		116,647	28,006
Inferred		165		101.9		0.48		0.16		540		1,746	582

1. Mineral Resources are reported based on 31 December 2021 topography surface.

2. The Mineral Resources are contained within a pit shell generated using an NSR cut-off of $33.20/t.

3. The Mineral Resources estimate is based on metal price assumptions of $22.00/oz silver, $0.95/lb lead, and $1.15/lb zinc.

4. Metallurgical recoveries vary with grade and average recoveries are, 98% silver, 95% lead, and 63% for zinc.

5. The point of reference for Mineral Resources is the point of feed into the processing facility.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

Table 11.10 Summary of Metallurgical Recoveries and Ownership of Chinchillas Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $22.00/oz Silver, $0.95/lb Lead, and $1.15/lb Zinc

Mineral Resources Classification		Tonnage<br>(kt)		Grades						Metallurgical Recovery				Cut-off<br><br>NSR<br><br>($/t)
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(%)		Lead<br>(%)		Zinc<br>(%)
Measured		1,110		99.2		0.86		0.31		98		95	63	33.2
Indicated		4,904		101.1		0.88		0.19		98		95	63	33.2
Measured + Indicated		6,013		100.8		0.88		0.21		98		95	63	33.2
Inferred		165		101.9		0.48		0.16		98		95	63	33.2

1. Mineral Resources are reported based on 31 December 2021 topography surface.

2. The Mineral Resources are contained within a pit shell generated using an NSR cut-off of $33.20/t.

3. The Mineral Resources estimate is based on metal price assumptions of $22.00/oz silver, $0.95/lb lead, and $1.15/lb zinc.

4. Metallurgical recoveries vary with grade and average recoveries are, 98% silver, 95% lead, and 63% for zinc.

5. The point of reference for Mineral Resources is the point of feed into the processing facility.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

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11.3.2 Pirquitas Mineral Resources

The Pirquitas Mineral Resources estimate is contained within underground mining shapes using an NSR cut-off of $100/t that is based on metal prices of $20.00/oz for silver, $1.10 for lead, and $1.30/lb for zinc. The NSR cut-off grade selected for the Pirquitas Mineral Resources assumes underground mining will be used for extraction, that the processing facility could be used for processing. It is recommended that the Mineral Resources estimate be re-evaluated and assessed with a study to determine the development horizon available prior to the completion of the Chinchillas open pit and the impact of the current operation on the Pirquitas Mineral Resource.

The Pirquitas Mineral Resources by classification is summarised in Table 11.11 and shown with recovery and ownership detail in Table 11.11.

SSR has advised that there are no known factors related to environmental, permitting, legal, title, taxation, socio-economic, marketing, or political issues that could materially affect the Mineral Resources estimates. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. The estimates of Mineral Resources may be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues. The quantity and grade of reported Inferred Mineral Resources are uncertain in nature and there has been insufficient exploration to classify these Inferred Mineral Resources as Indicated or Measured Mineral Resources. It cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resources as a result of continued exploration.

Mineral Resources are reported exclusive of Mineral Reserves.

Table 11.11 Summary of Pirquitas Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $20.00/oz Silver, $1.10/lb Lead, and $1.30/lb Zinc

Mineral Resources Classification		Tonnage<br>(kt)		Grades						Contained Metal
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(koz)		Lead<br>(klb)		Zinc<br>(klb)
Measured		79		444.5		0.197		1.17		1,129		343	2,044
Indicated		2,555		287.7		0.016		4.56		23,627		895	256,672
Measured + Indicated		2,634		292.4		0.021		4.46		24,756		1,240	258,715
Inferred		1,080		206.9		0.004		7.45		7,185		95	177,394

1. The Mineral Resources estimate is contained within underground mining shapes based on $90/t to $100/t NSR cut-off.

2. The Mineral Resources estimate is based on metal price assumptions of $20.00/oz silver, $1.30/lb zinc, and $1.10/lb lead.

3. Metallurgical recoveries vary with grade and on average are, 87% silver, 85% for zinc, and 50% for lead.

4. The point of reference for Mineral Resources is the point of feed into the processing facility

5. Mineral Resources are reported exclusive of Mineral Reserves.

6. SSR has 100% ownership of the Project.

7. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

8. Totals may vary due to rounding.

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Table 11.12 Summary of Metallurgical Recoveries and Ownership of Pirquitas Mineral Resources Estimate Exclusive of Mineral Reserves (as at 31 December 2021) Based on $20.00/oz Silver, $1.10/lb Lead, and $1.30/lb Zinc

Mineral Resources Classification		Tonnage<br>(kt)		Grades						Metallurgical Recovery				Cut-off<br><br>NSR<br><br>($/t)
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(%)		Lead<br>(%)		Zinc<br>(%)
Measured		79		444.5		0.197		1.17		87		50	85	100
Indicated		2,555		287.7		0.016		4.56		87		50	85	100
Measured + Indicated		2,634		292.4		0.021		4.46		87		50	85	100
Inferred		1,080		206.9		0.004		7.45		87		50	85	100

1. The Mineral Resources estimate is contained within underground mining shapes based on $90/t to $100/t NSR cut-off.

2. The Mineral Resources estimate is based on metal price assumptions of $20.00/oz silver, $1.30/lb zinc, and $1.10/lb lead.

3. Metallurgical recoveries vary with grade and on average are, 87% silver, 85% for zinc, and 50% for lead.

4. The point of reference for Mineral Resources is the point of feed into the processing facility

5. Mineral Resources are reported exclusive of Mineral Reserves.

6. SSR has 100% ownership of the Project.

7. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

8. Totals may vary due to rounding.

11.3.3 Mineral Resources Statement

The Mineral Resources estimate for Puna was completed by the SSR technical department on site. The Puna21TRS QPs have reviewed and accepted this information for use in the Puna21TRS. The Puna21TRS QPs reviewed the assumptions, parameters, and methods used to prepare the Mineral Resources Statement and are of the opinion that the Mineral Resources are estimated and prepared in accordance with the U.S. Securities and Exchange Commission (US SEC) Regulation S-K subpart 1300 rules for Property Disclosures for Mining Registrants (S-K 1300).

Mineral Resources are reported exclusive of Mineral Reserves and have been summarised by project and resource classification in Table 11.3.

Table 11.4 shows the cut-off values and metallurgical recoveries associated with the Mineral Resources.

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Table 11.13 Summary of Puna21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021)

Mineral Resources Classification			Chinchillas									Pirquitas							TOTAL
	Tonnage<br>(kt)			Grades							Tonnage<br>(kt)		Grades					Tonnage<br>(kt)		Grades
		Ag<br>(g/t)			Pb<br>(%)		Zn<br>(%)		Ag<br>(g/t)					Pb<br>(%)	Zn<br>(%)	Ag<br>(g/t)	Pb<br>(%)			Zn<br>(%)
Measured			1,110			99.2		0.86		0.31		79		444.5	0.197		1.17		1,189		122.2	0.81	0.37
Indicated			4,904			101.1		0.88		0.19		2,555		287.7	0.016		4.56		7,458		165.0	0.59	1.69
Measured + Indicated			6,013			100.8		0.88		0.21		2,634		292.4	0.021		4.46		8,647		159.1	0.62	1.50
Inferred			165			101.9		0.48		0.16		1,080		206.9	0.004		7.45		1,245		192.9	0.07	6.48

1. Mineral Resources are reported based on 31 December 2021 topography surface.

2. The Chinchillas Mineral Resources are contained within a pit shell generated using an NSR cut-off of $33.20/t. The Pirquitas Mineral Resources estimate is contained within underground mining shapes based on $90/t to $100/t NSR cut-off.

3. The Chinchillas Mineral Resources estimate is based on metal price assumptions of $22.00/oz silver, $0.95/lb lead, and $1.15/lb zinc. The Pirquitas Mineral Resources estimate is based on metal price assumptions of $20.00/oz silver, $1.30/lb zinc, and $1.10/lb lead.

4. The Chinchillas metallurgical recoveries vary with grade and average recoveries are, 98% silver, 95% lead ,and 63% for zinc. The Pirquitas metallurgical recoveries vary with grade and on average are, 87% silver, 85% for zinc, and 50% for lead.

5. The point of reference for Mineral Resources is the point of feed into the processing facility.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

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Table 11.14 Summary of Cut-off Values and Metallurgical Recoveries of Puna21TRS Mineral Resources Estimates Exclusive of Mineral Reserves (as at 31 December 2021)

Mineral Resources <br>Classification		Tonnage<br>(kt)		Grades						Metallurgical Recovery				Cut-off<br><br>NSR<br><br>($/t)
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(%)		Lead<br>(%)		Zinc<br>(%)
Measured – Chinchillas		1,110		99.2		0.86		0.31		98		95	63	33.2
Measured – Pirquitas		79		444.5		0.197		1.17		87		50	85	100.0
Indicated – Chinchillas		4,904		101.1		0.88		0.19		98		95	63	33.2
Indicated – Pirquitas		2,555		287.7		0.016		4.56		87		50	85	100.0
TOTAL – Measured		1,189		122.2		0.81		0.37		95		94	63	see above
TOTAL – Indicated		7,458		165.0		0.59		1.69		91		90	64
TOTAL – Measured + Indicated		8,647		159.1		0.62		1.50		92		91	64
Inferred – Chinchillas		165		101.9		0.48		0.16		98		95	63	33.2
Inferred – Pirquitas		1,080		206.9		0.004		7.45		87		50	85	100.0
TOTAL – Inferred		1,245		192.9		0.07		6.48		88		95	76	see above

1. Mineral Resources are reported based on 31 December 2021 topography surface.

4. The Chinchillas metallurgical recoveries vary with grade and average recoveries are, 98% silver, 95% lead, and 63% for zinc. The Pirquitas metallurgical recoveries vary with grade and on average are, 87% silver, 85% for zinc, and 50% for lead.

5. The point of reference for Mineral Resources is the point of feed into the processing facility.

6. Mineral Resources are reported exclusive of Mineral Reserves.

7. SSR has 100% ownership of the Project.

8. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

9. Totals may vary due to rounding.

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11.4 Comparison with Previous Mineral Resources Estimates

11.4.1 Chinchillas Comparison – 2021 vs. 2020

The Mineral Resources have been compared to the previous Mineral Resources, which were based on the EOY 2020 pit surface.

Key changes in the Mineral Resources (contained metal) have resulted from:

•Mining depletion (–6%).

•Socavon Resource write down (–26%).

•A review of the pit optimisation work for the Socavon deposit was undertaken using the NSR and other assumptions used for the Mantos deposit. The review concluded that there was no suitable pit shell produced to meet the standard of reasonable prospects for extraction. Therefore, the Socavon Mineral Resources previously reported by SSR have not been included in the 2021 Puna Mineral Resource.

•Engineering changes (–14%) due to updating of the Resource pit shell used.

11.4.2 Pirquitas Comparison – 2021 vs. 2011

The new Mineral Resources estimate is compared to the previous Mineral Resources estimate from 2011 and detailed in the NI 43-101 Technical Report dated 27 May 2016 (Davis et al., 2016). The Mineral Resources estimated in 2011 were based on metal prices of $11/oz silver, $0.70/lb lead, $0.70/lb zinc, and $5.00/ln Sn, and an NSR cut-off of $15.00/t NSR.

There was no change since EOY 2020 as there was no updated work on the Mineral Resources.

11.5 Conclusions and Recommendations

For the Chinchillas deposit it is recommended that MPSA examine advanced grade control (using reverse circulation drilling) at a grid spacing of 20 m, to determine if it will improve prediction particularly where the grade trends are horizontal. The shallow eastward dip of high grades should be carefully managed by pit mapping and advanced grade control drilling to provide appropriate levels of confidence to manage risk.

For the Pirquitas deposit it is recommended that the Mineral Resources estimate be re-evaluated and assessed with a study to determine the development horizon available prior to the completion of the Chinchillas open pit and the impact of the current operation on the Pirquitas Mineral Resource.

11.6 QP Opinion

The Puna21TRS QPs have not identified any relevant technical and/or economic factors that require resolution with regards to the Mineral Resources estimates.

11.7 Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects

The Mineral Resources reported in the Puna21TRS are suitable for reporting as Mineral Resources using the Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

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12 MINERAL RESERVES ESTIMATES

12.1 Summary

The Mineral Reserve estimates for the Project were completed by the SSR technical department on site. The Puna21TRS QPs have reviewed and accepted this information for use in the Puna21TRS.

This section describes the methodology and parameters used to estimate the Mineral Reserves for the Project.

Open pit mining is carried out by MPSA as an owner mining operation with ore hauled from Chinchillas pit to the Pirquitas plant. The Mineral Reserves were developed based on mine planning work completed in 2021 that included pit optimisation and re-design of the pit phases. The Puna21TRS QPs have reviewed and accepted this information for use in the Puna21TRS. Table 12.1and Table 12.2 summarise the Mineral Reserves for Chinchillas. The Chinchillas Mineral Reserves estimate has been generated for the Mantos deposit based on the following inputs: metal prices, resource model, geotechnical information, operating costs, mineral processing recoveries, concentrate transport, and off site costs and charges. Costs for all areas of the operation are estimated from actual costs. These were used to calculate an NSR of $44.11/t used for the Mineral Reserves cut-off.

Metal prices for the Mineral Reserves cut-off were selected after consideration of the pricing information described in Section 16, which includes a description of the time frame used for the selection of the price and the reasons for selection of such a time frame. The long-term prices for the cut-off were assumed to apply from the start of 2026. The metal prices are representative of the range of price estimates publicly reported for Mineral Reserves cut-offs.

12.2 Mineral Reserves Statement

Table 12.1 summarises the Mineral Reserves statement for the Puna Operations.

12.3 Factors that Affect the Mineral Reserves Estimates

Factors that affect the Mineral Reserves estimates include but are not limited to dilution; metal prices; off-site costs; metallurgical recoveries, pit slope designs; capital and operating cost estimates; and the effectiveness of managing environmental impacts. The main factors that affect the Mineral Reserves estimations reported in this section are:

•Commodity prices, particularly silver price.

•Processing recoveries.

•The effectiveness of managing environmental impacts for waste rock and downstream water flows.

•Pit slope design criteria.

The Mineral Reserves estimate has taken into account all known legal, political, environmental or other risks that could materially affect the potential development of the Mineral Reserves, as discussed in various sections of this Puna21TRS.

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Table 12.1 Summary of Chinchillas Mineral Reserves Estimate (as at 31 December 2021) Based on $18.50/oz Silver, $0.90/lb Lead, and $1.05/lb Zinc

Mineral Reserves Classification		Tonnage<br>(kt)		Grade						Contained Metal
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(koz)		Lead<br>(klb)		Zinc<br>(klb)
Proven		2,379		168.9		1.33		0.34		12,918		69,735	17,827
Probable		5,041		155.3		1.29		0.25		25,174		143,344	27,780
Probable Stockpiles		187		141.0		1.33		0.50		846		5,470	2,056
Proven + Probable		7,606		159.2		1.30		0.28		38,938		218,681	47,692

1. Mineral Reserves are reported based on 31 December 2021 topography surface.

2. The Mineral Reserves estimate is based on metal price assumptions of $18.50/oz silver, $0.90/lb lead, and $1.05/lb zinc.

3. The Mineral Reserves estimate is reported at a cut-off grade of $44.11/t NSR.

4. Economic analysis for the Mineral Reserves has been prepared using long-term metal prices of $21.00/oz silver, $0.90/lb lead, and $1.20/lb zinc

5. Processing recoveries vary based on the feed grade. The average recovery is estimated to be 98% for silver, 95% for lead, and approximately 63% for zinc.

6. Metals shown in this table are the contained metals in ore mined and processed.

7. The point of reference for Mineral Reserves is the point of feed into the processing facility.

8. SSR has 100% ownership of the Project.

9. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

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Table 12.2 Summary of Metallurgical Recoveries of Chinchillas Mineral Reserves Estimate (as at 31 December 2021) Based on $18.50/oz Silver, $0.90/lb Lead, and $1.05/lb Zinc

Mineral Reserve Classification		Tonnage<br>(kt)		Grades						Metallurgical Recovery				Cut-off<br><br>NSR<br><br>($/t)
	Ag<br>(g/t)		Pb<br>(%)		Zn<br>(%)		Silver<br>(%)		Lead<br>(%)		Zinc<br>(%)
Proven		2,379		168.9		1.33		0.34		98		95	63	44.11
Probable		5,041		155.3		1.29		0.25		98		95	63	44.11
Probable Stockpiles		187		141.0		1.33		0.50		98		95	63	44.11
Proven + Probable		7,606		159.2		1.30		0.28		98		95	63	44.11

1. Mineral Reserves are reported based on 31 December 2021 topography surface.

2. The Mineral Reserves estimate is based on metal price assumptions of $18.50/oz silver, $0.90/lb lead, and $1.05/lb zinc.

3. The Mineral Reserves estimate is reported at a cut-off grade of $44.11/t NSR.

4. Economic analysis for the Mineral Reserves has been prepared using long-term metal prices of $21.00/oz silver, $0.90/lb lead, and $1.20/lb zinc

5. Processing recoveries vary based on the feed grade. The average recovery is estimated to be 98% for silver, 95% for lead, and approximately 63% for zinc.

6. Metals shown in this table are the contained metals in ore mined and processed.

7. The point of reference for Mineral Resources is the point of feed into the processing facility.

8. SSR has 100% ownership of the Project.

9. All ounces reported represent troy ounces, and g/t represents grams per metric tonne.

10. Totals may vary due to rounding.

12.4 Comparison with Previous Mineral Reserves Estimates

The Mineral Reserves have been compared to the previous Mineral Reserves, which were based on the EOY 2020 pit surface. Comparison of the 2021 Mineral Reserves with the 2020 Mineral Reserves shows a net decrease in contained silver of 4.07 Moz (–9%).

Changes in Mineral Reserves (contained silver) can be summarised as Mining depletion (–16%) and Engineering changes (+5%) due to model updates, pit design changes and a small impact from increased metal prices.

12.5 Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects

The Mineral Reserves reported in the Puna21TRS are suitable for reporting as Mineral Reserves using the Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

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13 MINING METHODS

13.1 Geotechnical Review

A review of the geotechnical reports provided by MPSA was carried out. The following reports were provided and form the basis of this review:

•Knight Piesold, 2016. 2015 Geotechnical Site Investigation Report

•Knight Piesold, 2016. Pre-Feasibility Pit Slope Design VA201-439/3-2 Rev0

•Knight Piesold, 2016. Hydrogeologic Conceptual Model and Preliminary Pit Inflow Estimates

•Puna Operations Inc, 2016. NI 43-101 Technical Report

It is important to note, the Knight Piesold - 2015 Site Investigations report suggested a study regarding waste rock dumps (WRD) but have not been sourced or reviewed. SSR report that the review has not been undertaken. Knight Piesold prepared a report in 2020 after an inspection of the pit. The QPs did not review this report as it was not available during the Puna21TRS preparation.

13.1.1 Knight Piesold 2015 Site Investigation

The Knight Piesold 2015 site investigations were addressed in a report issued in February 2016. The report and the data therein provided the key source of information for additional Knight Piesold Geotechnical studies.

Figure 13.1 provides an overview of the extent of the Knight Piesold site investigations of which comprised:

•Five cored boreholes, geotechnically logged and with oriented core utilised in defining orientation of defects where possible.

•Packer testing was conducted in two boreholes and falling head testing in the other three boreholes.

•Observation wells in three vertical exploration boreholes and with slug testing therein.

Of particular note, Knight Piesold utilised mapping data of exposures carried out by GAR during exploration, Figure 13.1.

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Figure 13.1 Overview of Knight Piesold Site Investigation

Knight Piesold, 2015

The extent of studies is considered appropriate. Key aspects of note from the studies include the following:

•Two key lithologies are present, metasediments (basement sandstone), which has been overlain by a pyroclastic tuff, Figure 13.1.

•Metasediments are steeply dipping to the west. Intact strength is potentially underestimated from laboratory testing (average UCS of 30 MPa).

•Pyroclastic tuff has shallow bedding and has low intact strength due to alteration and with UCS of 10 MPa.

•The available orientated data in the pyroclastic tuff (from two boreholes) is limited. There are 33 data points, which indicate widely spaced defects.

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•The available orientated data in the metasediments (from four boreholes) is significant. Nearly 700 data points, which indicate a more jointed rock mass but with a lack of orientation of any faulting. Review of the logging and core photos suggest such faulting is present and the absence of understanding of orientation is considered as a significant gap in the studies.

•As shown in Figure 13.2, comparison of the surface mapping (mostly in metasediments) and orientated data from borehole CGA-209-G (in closest proximity to mapping and entirely in metasediments) indicates very good comparison. Of note is the good match between defect sets ‘D1’ (circled in red) and sets ‘C1’ and ‘A1’ (circled in blue). Of note is slightly flatter dip of set ‘D1’ in the surface mapping. The oriented data indicates the bedding (diamond symbols in set ‘D1’) is steeply dipping south to east. The variance from the broader interpretation noted above is owing to rotation introduced by intrusion of an underlying diatreme. However, the bedding noted in the surface mapping is moderately dipping to the south-west (set ‘A2’) and also moderately dipping to the west (set ‘B1’). This is at odds with the provided geological overview where bedding is steeply dipping in the metasediment and shallow dipping in the pyroclastic tuffs. The bedding has the potential to significantly control stability of the overall slopes, therefore this discrepancy needs to be resolved.

•Other oriented boreholes in the metasediments provide somewhat poorer comparison with mapping as this is a function of the blind zones impacting on the other boreholes.

Figure 13.2 Surface Mapping and orientated data from borehole GCA-209-G

Knight Piesold, 2016 Note: Defect sets annotated by OreWin. The blind zone (grey shading), where defects will be under-represented in oriented core.

Defect sets ‘D1’ = circled in red and ‘C1’ and ‘A1’ = circled in blue; bedding = diamond symbols in set ‘D1’

13.1.2 Knight Piesold PFS Pit Slope Design

The Knight Piesold 2016 report 'Pre-Feasibility Pit Slope Design VA201-439/3-2 Rev0' provides the pit slope recommendations for PFS studies. The following comments are based on a review of this report:

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•The kinematic assessment, which has been largely utilised to provide batter angles is considered appropriate for the metasediments. However, owing to the lack of knowledge on faults, there is a degree of uncertainty on interramp scale stability.

•Overall pit wall stability has been addressed through limit equilibrium stability analyses. The analyses utilised the Hoek & Brown (HB) rock mass strength criterion. The inputs of the analysis are considered largely appropriate based on the 2015 Geotechnical Site Investigation Report , however caution of the following three aspects:

•Knight Piesold have assumed a Disturbance Factor (D value) of 0.85. This value may be appropriate near created slope faces where blast damage may be present but is not considered appropriate for the rock mass within the pit slope.

•Knight Piesold have adopted an Ru value of 0.2 in the analyses, which nominally equates to a slope where the phreatic surface aligns with the mined surface and groundwater conditions are nominally 50% hydrostatic. This level of depressurisation is considered optimistic at the PFS stage.

•The HB criterion is poorly suited for rocks of low intact strength such as the pyroclastic tuff. As such, the interramp slope angle (IRA) in the tuff is considered marginally high and an overall angle in the order of 37° is considered more appropriate.

•The Knight Piesold design parameters are shown in Table 13.1. These parameters maintain similar berm widths in all areas and with variation in batter angle. An alternative would be to utilise 70° batter angles in all areas and utilise 10 m wide berms in the south-west and north-west (IRA of 49° maintained), 14 m wide berms in the south (IRA of 43° maintained) and 17 m wide berms in the east (IRA of 39°). A haul road would reduce the overall angle in the east wall but depending on placement may require revision of berm widths.

•Knight Piesold recommend that for the west wall, the uninterrupted IRA not exceed 150 m in height and with a wider bench or haul ramp of the order of 20–30 m to limit overall angle to 46° to 47°. In practice, with the ramp location unknown, would suggest placement of a wider berm in the south-west and north-west at nominally mid-height of the overall slope.

Table 13.1 Knight Piesold Recommended Interramp Angles

Pit Design Sector	Geotechnical Unit	Bench Face Angle<br>(°)	Bench Height<br>(m)	Bench Width<br><br>(m)	Interramp Angle<br>(°)
East	Pyroclastic	60	20	10	43
South	Basement Sandstone and Pyroclastic	60	20	10	43
South-west	Basement Sandstone	70	20	10	49
North-west	Basement Sandstone	70	20	10	49

Recommended angles assume quadruple benching

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13.1.3 2021 Pit Design Slope Criteria

SSR prepared new pit designs in 2021, the interpretation of the slope design recommendations used by SSR in the 2021 designs is shown in Table 13.2. The pit slope angles for both pit and dumps were based on those recommended by Knight Piesold but subject to some alteration by SSR. The 2021 pit designs have not been subjected to an independent geotechnical review and it is important that this review is carried out and the revised designs confirmed to meet the slope design criteria. This should be carried out as soon as possible in 2022.

Ore and waste are mined in 5 m benches. Final wall 20 m benches are formed by joining four working benches together. Haulage roads are 30 m wide, which is sufficient for two-way traffic of 100 t trucks, plus enough space to build a ditch and a safety berm. Interramp angles for the west and east walls are 49° and 43°, respectively. For every 150 m of slope height, either a 20 m geotechnical berm or a haulage road was added to the slope.

Table 13.2 Mine Design Criteria

Criteria	Unit	Value	Remarks
Bench height (final wall)	m	20	Ore and waste will be mined in 5 m benches
Bench face angle	degree	60 and 75	60° in east and south; the rest 75°
Catch bench width	m	11.4–13.3	On final walls
Geotechnical berm	m	20	For every 150 m height
Interramp angle	degree	40, 47, and 50	40° in east and south; 47° in the south-west; the rest 50°
Haulage road width	m	30	Two-way roads, includes berm and ditch
Maximum road grade	%	10
Rock dump face angle of repose	degree	35	25 m lifts and overall slope of 26°

13.1.4 Knight Piesold Hydrogeologic Conceptual Model and Preliminary Pit Inflow Estimates

The 2016 Hydrogeologic Conceptual Model and Preliminary Pit Inflow Estimates report has utilised the available knowledge to summarise hydrogeological parameters, provide a conceptual groundwater model and provide inflow estimates. This study is considered appropriate at a PFS level.

13.1.5 NI 43-101 Technical Report

The NI 43-101 Technical Report of December 2016 provides limited additional geotechnical value to that indicated in the previous Knight Piesold reports. The pit designs utilised overall angles 3° flatter than those indicated in Table 13.1. The flatter slope angle was to allow for haulage ramps (20–30 m width) and a geotechnical berm every 150 m of slope height as proposed by Knight Piesold (albeit this was only recommended for the west wall).

The haul road placement shown in the NI 43-101 Technical Report , and as provided by OreWin, transects the east wall and such that the overall angle of 37° is largely achieved.

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The report places the ex-pit waste dump to the north-east of the pit, which is represented by a red dashed line in Figure 13.1, and where the topography suggests a somewhat uniform slope dipping to the south. The following information is inferred:

•The dump footprint has a dip of nominally 13°.

•The dump would be constructed in a bottom-up sequence.

The waste rock dump (WRD) investigations in the 2015 Geotechnical Site Investigation report utilised test pitting, indicated typically a shallow depth to rock in the dump footprint. Although no previous WRD geotechnical studies have been sourced for this review, it is anticipated with clearing of overburden from the footprint of the first dump lift there should be negligible risk of waste dump instability.

13.1.6 Summary of Geotechnical Studies Review

the following is an overview of comments for the geotechnical studies review for the Chinchillas project:

•Review of the logging and core photos in the metasediments suggest faulting is present. The absence of understanding of orientation is considered as a significant gap in the geotechnical studies. Owing to the lack of knowledge on faults, there is a degree of uncertainty on interramp scale stability. It is recommended that MPSA considers three boreholes in the western quadrant with use of televiewer (ATV) logging. ATV, which uses scanning of the borehole wall, is far more reliable in providing the orientation of major structures which are typically present in recovered core as rubble zones, broken core or highly jointed zones, which invariably cannot be orientated in oriented core as used in the PFS investigations.

•Bedding in the surface mapping shown in Figure 13.2, is moderately dipping to the south-west (set ‘A2’) and moderately dipping to the west (set ‘B1’). This is at odds with the provided geological overview where bedding is steeply dipping in the metasediment and shallow dipping in the pyroclastic tuffs. As bedding has the potential to significantly control stability of the overall slopes this discrepancy needs to be resolved.

•Overall pit wall stability was addressed in the PFS study through limit equilibrium stability analyses utilising the Hoek & Brown rock mass strength criterion. It is considered the inputs as largely appropriate. However, caution of the following three aspects is recommended.

•The level of depressurisation in the analyses is optimistic at PFS stage.

•The Hoek & Brown criterion is poorly suited for rocks of low intact strength such as the pyroclastic tuff. As such, the IRA in the tuff is considered marginally high and an overall angle in the order of 37° is considered more appropriate.

•The Knight Piesold design parameters, shown in Table 13.1, maintain similar berm widths in all areas and with variation in batter angle. An alternative would be to utilise 70° batter angles in all areas and utilise 10 m wide berms in the south-west and north-west (IRA of 49° maintained), 14 m wide berms in the south (IRA of 43° maintained) and 17 m wide berms in the east (IRA of 39°). A haul road would reduce the overall angle in the east wall but depending on placement may require revision of berm widths.

•The 2021 pit designs have not been subjected to an independent geotechnical review and it is important that this review is carried out and the revised designs confirmed to meet the slope design criteria. This should be carried out as soon as possible in 2022.

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13.2 Mining

Open pit mining is carried out by MPSA as an owner-mining operation with ore hauled from Chinchillas pit to the Pirquitas plant.

The Chinchillas deposit is located in the high lands of the Andes. The topography of the property consists of several mountains and hills on the sides of property with a small valley in the middle. The orebody is located mainly in the bottom of the valley with extensions stretching to the west on the hillside. The elevation varies from about 4,090 metres above sea level (masl) in the east side of the valley to 4,300 masl on the peaks in the west. There is a small creek in the middle of valley running from west to the east.

The Chinchillas deposit is mined as a conventional open pit operation. Most of the in-pit haulage for both ore and waste is carried out using 100 t haulage trucks. Ore is mined in 5 m benches and stockpiled in a staging area close to the pit. From the staging area, ore is transported to the crusher at the Pirquitas Operation which is 42 km away from Chinchillas. Throughout the mining operation, low grade ore is stockpiled near the pit rim to be processed at the end of mine life. The mining operation is conducted by the owner. Ore haulage was changed to owner operated in 2021.

Waste rock is mined and hauled to two major on-site rock storage facilities based on geochemical characteristics.

For the mine planning work the NSR is calculated for each block. No dilution is included in the block model. Ore is placed in the ore staging area as it is mined from the pit. The ore is then loaded onto haul trucks and transferred to Pirquitas on a daily basis. Material that falls between the Resource cut-off and the Reserve cut-off is stockpiled separately as mineralised waste.

13.3 Mine Design

The Mineral Reserves were developed based on mine planning work completed in 2021 that included pit optimisation and re-design of the pit phases. The 2021 ultimate pit design and the waste dumps at Chinchillas are shown in Figure 13.3.

Figure 13.4 shows a long-section of the main pit from west to east.

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Figure 13.3 Chinchillas 2021 Ultimate Pit Design

MPSA, 2020

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Figure 13.4 Long-Section of the 2021 Pit Design

MPSA, 2020. (A=west, A’=east)

13.4 Rock Storage Facilities

Some of the waste types have the potential to leach metals and are separated from the neutral waste material. Based on geochemical characteristics, waste is classified into three groups designated ‘A’, ‘B’, and ‘C’. Type ‘A’ waste is stored close to the pit as it has the potential to leach metals. This is so that the drainage can be collected in the pit and if necessary be treated. Types ‘B’ and ‘C’ are stored together in the same location.

According to this classification, two waste rock storage facilities have been designed for Chinchillas to accommodate different rock types. These can be seen in the general site layout Figure 13.3.

Rock storage ‘A’ is close to the pit, on a hill side to the north-east of the Chinchillas pit. The toe of this dump is 100 m offset from the pit rim. Rock storage ‘B’ and ‘C’ are located to the south-east of the active mining area on somewhat flatter terrain. Waste Rock Facilities are built with 25 m lifts and 15 m berms. The angle of repose for each lift is 35° and the overall slope angle of dumps is 26°. Access to the dump is by 30 m wide haulage roads. The total height of the dumps are approximately 100 m.

13.5 Mining Equipment and Personnel

The mine operates 355 days a year with two 12-hour shifts per day. The amount of mining equipment required for the operation varies by the tonnages of material moved in each period.

The mining operation utilises 12 m3 wheel loaders to load 100 t off-highway trucks. Two main drills and two smaller drills will be sufficient for the life of mine. Dozers (D9) and graders provide ancillary support. The graders are also used for maintaining the 42 km ore haulage road.

The mining personnel are grouped into three sections: operation, maintenance, and management / technical.

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13.6 Production Scheduling

Mining production is scheduled throughout the current mine life which is five years. Ore is classified into grade bins: low grade and high-grade ore. Milling cut-off grade is calculated to be $44.11/t NSR.

Table 13.3 Production Schedule for Chinchillas Project

Description		Unit		Total	Project Year
	2022		2023		2024	2025	2026
Total Movement		kt		29,671	10,652	9,168	5,316	3,725	809
Waste Mined		kt		22,469	9,033	7,437	3,655	2,008	336
Ore Mined		kt		7,202	1,619	1,732	1,661	1,717	473
Strip Ratio		kt		3.1	5.6	4.3	2.2	1.2	0.7
Processed		kt		7,352	1,643	1,643	1,647	1,647	773
Silver Feed grade		g/t		160	168	176	158	158	123
Lead Feed grade		%		1.32	1.24	1.38	1.43	1.43	0.94
Zinc Feed grade		%		0.29	0.47	0.34	0.19	0.19	0.22
Ag Recovery		%		95.5	96.2	96.2	96.2	93.6	94.7
Ag Produced		koz		37,210	8,529	8,924	8,026	7,809	3,921
Payable Ag Produced		koz		35,135	8,051	8,427	7,581	7,375	3,701

Metal produced includes current concentrate stockpiles containing 242 koz silver and 5 Mlb lead.

13.7 General Mine Site Layout

Figure 13.5 shows the general site layout of Chinchillas mine.

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Figure 13.5 Chinchillas General Mine Site Layout

Knight Piesold, 2019

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14 PROCESSING AND RECOVERY METHODS

Chinchillas material is processed at a rate of up to 1.7 Mtpa through the existing Pirquitas Operation process plant. The Pirquitas plant was commissioned in 2009 and has since been in continuous operation.

The plant has not been expanded since start-up; however, minor changes in the flotation flow sheets have occurred to optimise performance.

14.1 Process Overview for Chinchillas

The Pirquitas plant was upgraded in 2017 to process the Chinchillas ore types, producing a silver / lead concentrate and a zinc concentrate.

A schematic diagram of the Chinchillas process flow sheet is shown in Figure 14.1.

Figure 14.1 Chinchillas Processing Flow Sheet Overview

MPSA, 2021

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14.1.1 Stockpiling and Crushing

The trucked material is delivered to suitable stockpiles at the primary jaw crusher. The jaw crusher is fed directly via 35–42 t truck dumping or with a front-end loader. Mill operations decides daily feed blending from the ore stocked looking for steady head grade according to the production plan.

Secondary / tertiary crushing and screening operations will reduce this material to an 80% passing size of 9 mm. This material is discharged onto a crushed feed stockpile with four feeders located beneath the stockpile.

The crushing circuit was designed to process up to 6,000 tpd.

Figure 14.2 shows the crushing circuit flow sheet.

Figure 14.2 Chinchillas Crushing Circuit

MPSA, 2021

14.1.2 Grinding

The ball mill circuit grinds crushed ore to the optimum size at a rate of 4,500 tpd. The ball mill is 4.8 m in diameter by 6.25 m long with 2,400 kW of installed power. Mill discharge is pumped to a cyclone nest where the underflow is returned to milling operations and the overflow reports to flotation.

The addition of granular lime to the ball mill feed belt is done for flotation pH control. The pyrite / sphalerite depressant and frother are added into the mill. The lead / silver flotation collector and a reinforcement of frother are added to the cyclone overflow.

14.1.3 Lead / Silver Flotation

The lead / silver flotation section consists of rougher, and in the concentrate cleaning stage with a scavenger stage.

Figure 14.3 shows the lead / silver recovery circuit for Chinchillas ore.

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Figure 14.3 Grinding and Lead / Silver Recovery Circuits for Chinchillas

MPSA, 2021

14.1.4 Zinc Flotation

The zinc flotation circuit consists of rougher, and one stage of conventional cell concentrate cleaning followed by one stage of column cell cleaning.

Figure 14.4 shows the zinc recovery flow sheet for processing Chinchillas material.

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Figure 14.4 Zinc Recovery Circuit for Chinchillas

MPSA, 2021

14.1.5 Concentrate Handling

The Pirquitas silver / lead concentrate dewatering circuit consists of a thickener, holding tank and pressure filter. However, as higher lead feed grades are mined after the first few years of operation, the existing tin concentrate thickener will be recommissioned along with an additional new filter press to handle the higher lead / silver concentrate production.

The Pirquitas zinc concentrate dewatering circuit consists of a thickener, holding tank and filter.

After filtering, the concentrates are bagged into one tonne bulk bags. Sampling will be by manually inserted spear samplers.

14.1.6 Tailings Handling

The Pirquitas plant tailings thickener was designed to treat a low density, tin circuit tailings (approximately 20% solids) at 4,090 tpd. The Pirquitas plant has operated successfully on zinc tailings at higher tonnages. The thickened solids (55% to 58% solids) are pumped 6 km to a portion of the mined-out Pirquitas pit for storage. Water recovery is a combination of tailings thickener overflow and in-pit pond, both recycled to the plant reclaim water system.

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14.2 Process Plant Performance

The Pirquitas process plant has continued to improve performance after upgrading and converting to Chinchillas ores in 2018. These improvements have included better understanding of flotation characteristics of the ores, improved operating and maintenance practices and a change of cyclone in the grinding circuit.

These changes have seen improvement from the expected 4,000 tpd capacity to 4,500 tpd achieved in 2021.

Table 14.1 summarises mill feed and grade, with recovery and production for metals in concentrate.

Table 14.1 Mill Production Summary 2018 to 2021

	Unit	2018	2019	2020	2021
Ore Milled	kt	1,420	1,393	1,118	1,643
Ag Feed	g/t	114	184	164	158
Zn feed	%	0.84	0.54	0.51	0.57
Pb Feed	%	0.85	0.89	0.77	1.12
Silver Recovery	%	72.1	93.2	94.6	95.8
Zinc Recovery	%	39.3	49.2	55.5	65.6
Lead Recovery	%	82.6	85.8	90.2	93.0
Silver Produced	koz	3,747	7,674	5,581	8,010
Zinc Produced	klb	8,775	8,392	6,988	13,641
Lead Produced	klb	3,107	23,958	17,193	37,695

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15 INFRASTRUCTURE

The main approach to infrastructure for the Project is to maximise the use of existing infrastructure and facilities at the Pirquitas Operation and minimise the building of new items at the Chinchillas site.

The Pirquitas Operation includes significant infrastructure used to sustain mining and processing operations over the last seven years, much of which remains suitable for continued operation. These facilities include roads, a gas pipeline, power generation facilities, water diversion systems, tailings dams, mine waste stockpiles, camp facilities, office buildings, maintenance shops and communications systems.

15.1 Ore Haulage

The ore transport road from Pirquitas to Chinchillas is the National Route No. 40 (Route 40) that leads to Provincial Route No. 70 (Route 70). The route required upgrading in order to have the increased traffic, including 35-42 t ore haulage trucks, safely and efficiently travel the route. A road survey was completed, and a road design was developed and constructed to widen roads and improve route conditions, including bypasses of the local villages of Orosmayo and Liviara to minimise social impacts. This design, along with improved river and creek crossings and the requirement for road surface topping. Figure 15.1 shows the access road route.

Figure 15.1 Access Road for the Project and Proposed Modifications

Knight Piesold, 2019

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15.2 Gas Pipeline and Power Supply

For its source of electricity, the Pirquitas Operation uses natural gas to power three Wärtsila generator sets, each with a capacity of five megawatts (MW) of power. In addition, the same electrical plant has three diesel-powered Cummins generators, each yielding 1.1 MW. There is 6.7 km of gas pipeline on the Pirquitas property. The pipeline is 152 mm diameter, constructed of API5L Grade B steel with 4.8 mm wall thickness in normal applications and 7.1 mm wall thickness at river or drainage crossings.

Power for the Chinchillas mine site supplied along existing power lines from the natural gas powered generators at Pirquitas. EJESA is the local power authority that owns the lines. The power line from Pirquitas that goes directly past the rural EJESA line at the town of Nuevo Pirquitas (approximately 5 km from Pirquitas). The rural power line then goes from Nuevo Pirquitas to all villages along Route 40 and Route 70 and directly to Santo Domingo. This line is able to carry the 1 MW load for Chinchillas, with a small spur line (approximately 4 km in length) to take power into the mine.

15.3 Water Supply

Water supply for the Pirquitas Operation comes from the northwards flowing Collahuaima River which lies immediately east of the property. Water is pumped 7 km to the mill from a site known as San Marcos located within the mine property, a short distance downstream from where the Pirquitas River drains into the Collahuaima River. By means of Permit No. 201/002, originally granted to Sunshine Argentina by the Dirección Provincial de Recursos Hidricos de Jujuy and recorded by the Ministerio de Obras y Servicios Publicos on 23 July 1998, the mine is allowed to draw up to 32 L/s of water from the river.

Water supply for the Chinchillas mine will be supplied via local pumping wells. There is allowance for a water distribution system, equipment washing, road dust control, sewage and fire water facilities. Potable water for Chinchillas will be supplied by bottles and larger water totes.

15.4 Tailings

MPSA is currently using the Pirquitas pit as a tailings reservoir. These tailings come from the processing of the Mina Chinchillas ore.

The use of the pit as a tailings reservoir was approved by the Authority through Resolution No. 056/2018, after submitting the Addendum to the Authority in August 2017.

Placing the tailings inside the pit involves transporting them from the process plant to the pit by means of a pumping system and a 6.3 km pipeline to the tailings box located on the edge of the pit, to be discharged to the pit. Likewise, the tailings reservoir has a water recovery system to pump the water (from the tailings, the flows that enter the pit by filtration, direct rain and surface run-off) to the process plant for reuse. This pipe follows the same route as the pipe that transports the tails. The disposal of the tailings in the pit began in April 2019.

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Thickened tailings (55% solids) are transported to a portion of the Pirquitas pit through a pipeline for in-pit disposal, tailings in-pit discharge system from the tailings transport pipeline, in-pit water reclaim system, and pipeline from the Pirquitas pit to the Pirquitas plant for reuse. Water recovery will be a combination of tailings thickener overflow and in-pit pond, both recycled to the plant reclaim water system. These proposed upgrades will allow for additional tailings capacity in connection with the processing of Chinchillas ore. The distance from the Pirquitas plant to the Pirquitas pit is 6 km and the grades vary from 1.7% to 3.0% uphill. The alignment and gradient is shown in Figure 15.2.

Figure 15.2 Alignment and Gradient of the Tailings Line for In-pit Disposal

MPSA, 2021

15.5 Communications Systems

The Pirquitas site is equipped with both cellular and landlines. This equipment uses cell phone towers to communicate to Abra Pampa and is connected via a land line to the Pirquitas mine offices and buildings. On-site communication at Chinchillas is via radio communication and local phone.

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15.6 Camp, Office, and Chinchillas Infrastructure

The Pirquitas camp site is equipped with housing sufficient for a maximum of 673 personnel. This housing is a mix of rehabilitated housing from prior mining operations and modular housing that was installed during construction. It is anticipated that Chinchillas and Pirquitas operating management and senior staff will be housed at the Pirquitas camp while local workers and operators will be transported to their local villages.

Camp food is catered by a contractor and is provided on a seven day per week schedule. Food as required by Chinchillas workers will be delivered daily to Chinchillas.

Office buildings at Pirquitas are a mix of rehabilitated offices from prior mining operations and modular office space installed during mine construction.

The following facilities are located at Chinchillas:

•Mine and administration offices

•Truck shop

•Lunchroom (food preparation and storage is at the Pirquitas camp – daily delivery)

•Change room / Bathrooms / Training room

•Water wells, distribution and sewage system

•Lighting and heating facilities

•IT network

•Explosives magazines, and transfer of emulsion silos from MPLLC

•Fire and lightening protection

•Oil and fuel storage

•Security and first aid buildings

•Solid waste storage facility

Solid waste materials will be collected at the mine site and will be delivered to Pirquitas for recycling. A small landfill facility will be developed at Chinchillas site for small amount of solid waste produced at site. The explosives facilities are located at Pirquitas in accordance with Argentine mining regulations.

The infrastructure and facilities listed above can be seen in general site layout in Figure 13.5.

15.7 Mine Short-Term / Long-Term Ore Stockpiles

In the east side of the pit, adjacent to the pit rim, a pad has been developed using Type ‘C’ waste materials for multi-purpose tasks. The size of the pad is approximately 400 m x 300 m. This includes a staging area for loading ore onto the haulage trucks to be transported to the mill. A short-term ore stockpile of ore will be formed in this area, with the amount of stockpiling varying by period. A small amount (690 kt) of low-grade ore will also be stockpiled on this pad as long-term stockpile. This will be milled at the end of mine life before closing the mine. Refer to Figure 13.5 for general site layout where the location of short-term and long-term stockpiles are shown.

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15.8 Rock Storage Facilities

The mine currently has two waste stockpiles as described in Section 13.4. Rock storage facilities are classified by their geochemistry attributes as discussed in Section 17. Potential acid generating rock (Type ‘A’) will be disposed close to the pit rim so that its drainage will be collected in the pit and treated accordingly at closure. Mineralised waste will be separated and stockpiled with Type ‘A’ material, but adjacent to the ore stockpiles, for potential processing opportunities at a later date. High metal leaching materials (Type ‘B’) will be stored with Type ‘C’ (non-hazardous materials) with a controlled drainage system. Rock storage facilities can be seen in general site layout at Figure 13.5. More information about managing Type ‘A’ and ‘B’ materials can be found in Section 17.

15.9 Other Pirquitas Infrastructure

The Pirquitas site has a permitted wastewater treatment facility for treatment of liquid waste from camp operations. This system is designed to allow for discharge of treated wastewater to national standards.

The site has a landfill for organic waste and a recycling centre for plastics, wood, and metal products. Most wood products are donated to the local communities and are used as fuel or for construction supplies. Scrap steel and specialty steels are recycled via local vendors.

Domestic water comes from a water diversion located in the Medano Canyon area which is approximately 300 m upstream from the Pirquitas mine open pit. Water is pumped from that location to a site water treatment facility for filtering and chlorination and is then used within the camp site. At the date of this Puna21TRS, potable water is currently supplied by bottles and totes for drinking and cooking purposes.

Concentrate shipments from Pirquitas are currently trucked to Susques, Jujuy from Pirquitas via Route No. 77, and from there to Buenos Aires via Route No. 9. At arrival to the terminal, the material is directly dispatched from the port facilities to the concentrate buyers.

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16 MARKET STUDIES

The Project is a poly-metallic project containing three principal metals – silver, lead, and zinc. Production is from two separate concentrates: a high silver content lead concentrate and a zinc concentrate.

The lead concentrate contains most of the recovered silver metal and is the more valuable of the two concentrates. Trace amounts of minor penalty elements are present in both of the concentrates.

16.1 Marketing and Metal Prices

Silver is traded on a global basis on a number of metals and commodity market exchanges. The price is determined by a number of factors that follow short and long-term trends and is most commonly established on the London Metal Exchange.

Metal prices for the economic analysis were estimated after analysis of consensus industry forecasts and compared to metal prices used in other published studies. The metal prices selected have taken into account the current project life. The metal prices are representative of industry forecasts. Lead and zinc are relatively low compared to the consensus prices. The prices used for the economic analysis are shown in Table 16.1.

Table 16.1 Metal Price Assumptions

Commodity	Unit	2022	2023	2024	2025	Long-Term
Silver	$/oz	24.00	23.00	22.00	21.00	21.00
Lead	$/lb	1.00	0.95	0.93	0.92	0.90
Zinc	$/lb	1.30	1.20	1.20	1.20	1.20

No external consultants or market studies were directly relied on to assist with the sales terms and commodity price projections used in the Puna21TRS. The QPs for this Section 16 agree with the assumptions and projections presented.

16.2 Concentrate Terms

The Chinchillas concentrates are commodities and are sold and traded in global markets. Sales are either made directly to smelter operations or through commodity traders.

The logistics, required customs procedures, and exporting requirements are therefore well understood by the MPSA.

Average concentrate terms and transport are based on experience at the Puna Operations and are shown in Table 16.2.

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Table 16.2 Concentrate Marketing Terms and Charges

Item	Unit	Lead Concentrate	Zinc Concentrate
Treatment Charge and Refinery Charge (TCRC)	$/t Conc.	1,191	724
Payability – Silver	%	95	75
Payability – Lead	%	95	–
Payability – Zinc	%	–	85
Deduction – Lead	%	3	–
Deduction – Zinc	%	–	8
Minimum Payout Factor	%	63	39
Royalty	%	3	3
Export Duty (revenue minus TCRC's)	%	5	5
Puna Credit (revenue minus TCRC's)	%	2.5	3

The concentrate quantities produced by period are displayed graphically in Figure 15.2 and are derived from the annual mine production schedule.

Figure 16.1 Concentrate Production

OreWin, 2021

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16.3 QP Opinion

Data and assumptions for macroeconomic trends, taxes, royalties, interest rates, marketing information and plans, legal matters such as statutory and regulatory interpretations affecting the mine plan, and environmental matters are outside the expertise of the QPs and are within the control of the registrant (see Section 25).

Additionally, the project has been in operation for a number of years, and following a review of the current supplied information, the opinion of the Puna21TRS QPs is that he current plans and input parameters appear adequate for use as inputs to the Puna21TRS.

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17 ENVIRONMENTAL STUDIES, PERMITTING AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS

Significant environmental and social analysis has been conducted for both mines. A summary of key physical, chemical, and biological information is provided in the following sub-sections.

17.1 Chinchillas

17.1.1 Surface Hydrology and Water Quality

The Chinchillas property is located in a small, contained valley near the headwaters of the Colquimayo and Orosmayo rivers. Drainage from small ephemeral streams into the Project area collect in the valley bottom in the Arroyo Uquillayoc, which drains to the east into Rio Colquimayo.

Flows in the small tributaries that drain the Project area are governed primarily by rainfall, which is typically highest between December and March. Typical flows in the Arroyo Uquillayoc near the Project site are low, ranging from 0–1.5 litres per second (L/s) during the dry period, and between 0.3 and 20 L/s during the rainy season.

Surface water quality samples were obtained and analysed from 22 sites between 2011 and 2016, from within the Project area in the Arroyo Uquillayoc as well as far-afield sites in Quebrada San Pedro, the Rio Colquimayo and Rio Cincel, as well as the Rio Orosmayo.

Both surface and groundwater baseline sampling show the influence of native mineralisation in the host rock. While surface water chemistry is generally circumneutral, Arroyo Uquillayoc near the Project site seasonally shows variation from acidic (pH 5.9) during higher flows to basic (maximum pH 8.0) during lower flows. Annual average pH at these sites was neutral, between 6.8 and 7.2.

In Argentina, the Environmental Protection for Mining Activity Law (Ley de Protección Ambiental para la Actividad Minera in Spanish) specifies limits of parameter concentrations in water quality in the absence of site-specific data for various end uses, including drinking water, aquatic life, irrigation, and livestock watering. Metals such as aluminium, antimony, arsenic, barium, boron, cadmium, copper, chromium, iron, lead, manganese, nickel, vanadium, and zinc occasionally are at, or exceed, these concentrations in the baseline water sampling.

Surface water parameters in the Quebrada de San Pedro exhibited generally more neutral pH, but with similar metal concentrations.

The sampling location in the Arroyo Uquillayoc as it exits the Chinchillas valley will be used during operations as a point of control to monitor water quality during operations. In the baseline condition, samples from the Arroyo Uquillayoc at the outlet of the Chinchillas valley exhibited exceedances for a number of the limits set by the Environmental Protection for Mining Activity Law. This suggests that some metal parameters occur naturally in higher concentrations in Project area waters, which would be expected, as they are draining the valley that contains the mineralised zone. Mitigation and management programmes are part of the Project permitting. These programmes consider the naturally elevated baseline parameters.

Currently MPSA monitors 17 of the original 22 sites given the expansion of the mine. The monitoring programme includes 1 control point and two compliance points. MPSA monitoring to indicates that the water quality values are between the maximum and minimum baseline parameters.

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17.1.2 Hydrogeology

The Chinchillas site is located in a caldera or bowl-like feature in the side of the mountain range, resulting in some flow towards the bowl from the north and south as well as from the east. The bowl is somewhat like a shallow open pit.

Groundwater discharges to topographic lows, such as the local drainage in the deposit area depression and to the regional low elevation at the base of the range to the east and west of the Project area. Elevations are highest along the south–south-west / north–north-east divide of the Sierras and decrease towards the east and west. Groundwater gradients are therefore steepest towards the east and west, and groundwater is expected to generally flow in these directions following topography.

Hydrogeological data were collected during a 2015 site investigation consisting of drillhole logs, hydraulic conductivity testing (packer tests and open-hole tests), water level observations, and drilling circulation records. Sixteen packer tests and nine open hole falling head tests were completed in three geotechnical drillholes in the deposit area. Hydraulic conductivity values estimated from the packer tests range from less than 1 x 10–8 m/s to 1 x 10–5 m/s (Figure 17.1).

Figure 17.1 Response Test Hydraulic Conductivity by Lithology

MPSA, 2021

The metasediments outside the caldera feature are expected to have a relatively low hydraulic conductivity. Storage values are expected to be low, provided almost entirely by joints, fractures, bedding planes and faults. Within approximately 300 m from the contact margins with the overlying tertiary pyroclastics, the permeability of the metasediments increases due to the strongly fractured nature of the rock.

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North-west trending faults likely provide partial barriers to groundwater flow across the faults and enhanced flow parallel to faults. The fractured zone adjacent to the metasediments has relatively high hydraulic conductivity, likely in excess of 1 x 10–6 m/s.

Groundwater discharges occur primarily in topographic lows, often into stream beds. The indications from the available surface flow measurements are that groundwater discharge contributes from 1.5 L/s to upwards of 4 L/s to stream flows at the eastern extent of the Chinchillas valley. The groundwater reporting to the pit area is estimated to be 1.8 L/s.

Arid climatic conditions result in relatively high evapotranspiration rates that ultimately minimise the amount of precipitation available for groundwater recharge. The variation in annual precipitation impacts the precipitation available for groundwater recharge from one year to the next.

Recharge could vary from insignificant to about 50 mmpa, depending on climatic conditions and surface materials. This is expected to result in water level increases of a few metres in wet years, which would decrease over drier years. Smaller variations can be expected on a seasonal basis.

Currently the dewatering system consists of sumps located on the base of the pit and discharged through a pump to a contact water pool near the facility (‘A’ Pond – contact water). This water is used for dust control.

MPSA is evaluating the necessity of a dewatering system consisting of wells containing submersible pumps located in the perimeter of the pit.

Groundwater quality samples from monitoring wells immediately adjacent to the Project area were collected in 2015 and 2016. Similar water quality parameters were observed in the groundwater to those identified in the surface water samples discussed above.

Sample results were compared to limits specified in the Environmental Protection for Mining Activity Law. As was noted in the surface water, exceedances were noted in the baseline condition for some metals parameters. These variably included exceedances of the drinking water, aquatic life, irrigation, and livestock watering limits. However, these exceedances are considered natural and represent water that drains from within and around the mineralised zone and are carefully documented as part of the baseline monitoring programme.

The current monitoring programme includes one well located downstream of the Chinchillas mine. The most recent shows water quality values are between the maximum and minimum baseline parameters.

17.1.3 Geochemistry

Geochemical investigations were undertaken in order to assess the potential for net acid generation and the potential for metal leaching. As described above, both surface water samples and groundwater samples in the area of the mineralisation show circum-neutral pH values. Water samples exhibited slightly elevated sulfates (ranging from <25 mg/L SO4 to 100 mg/L SO4), alkalinity up to 100 mg/L and a range of dissolved and total metals. There are no strongly acidic seepages found in the Project area, either in the surface drainage or the groundwater. Of particular interest in the prediction of water chemistry from the Project, there are slightly elevated values of Al, Zn, Cd, Fe, Mn, and Sb found in some baseline samples. These metals are consistent with the mineralisation of the Project area and the Chinchillas deposit.

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The regional geology comprises a package of sediments overlain by volcanics. Within this region, the deposit was formed by a major east–west trending fault structure along which volcanic intrusions and mineralising events have resulted a zone of pyroclastic rocks (breccias, tuffs and ash) forming a roughly elliptical deposit. The deposit has undergone several different types of alteration, primarily clay alteration with lesser sericitisation, silicification, and carbonate alteration. The deposit lithology is therefore broadly grouped by lithology into (meta)sediments and volcanics, and further by alteration.

Silver, lead, and zinc bearing minerals include silver sulfosalts, boulangerite, tetrahedrite, freibergite, sphalerite, and galena. Associated mineral assemblages include chalcopyrite, pyrite, siderite, limonites, manganese oxides, and malachite. The mineralisation occurs as disseminated within the breccias but primarily along structure within the volcanics and basement rocks. Considering the environmental geochemistry, this deposit would be considered a low sulfide system and a low carbonate (alkalinity) system.

A suite of 34 samples were selected for geochemical testing to provide spatial coverage of the expected mine areas and to evaluate the characteristics of the various lithological and alteration units, and mineralisation within ore and waste for the deposit. The extensive exploration ICP database was evaluated before selecting the samples in order to ensure that representatives of low grade ore and waste rock were selected.

Testing included both standard elemental analyses (by ICP) and acid base accounting to characterise the range of sulfide content (and therefore potential for acid generation) and carbonate content (and therefore potential for neutralisation).

The static test results are consistent with those expected from the deposit geology; relatively low sulfur content and low carbonate content, and mineralisation concentrated in the breccias. The metal contents reflect the main minerals in the deposit, with Zn and Cd associated with the sphalerite, Al associated with the clay alteration, and Cu occurring in the freibergite and chalcopyrite.

The key findings with respect to the potential for net acid generation are:

•Paste pH of the samples range from neutral to slightly acidic, with the majority of the samples between paste pH of 5.7 to 8.1.

•Total sulfide content of the samples is low, ranging from <0.01% to 4% S, with one sample of breccia at 7% S. This is consistent with the statistical analysis of the entire exploration IPC database of the deposit (including ore) which shows sulfide concentrations range from <0.1 wt% to >10 wt% with an average of 0.75 wt% for the deposit.

•Carbonate concentrations are relatively low, ranging from less than detection to 4.3 kg/t CaCO3 equivalent.

•Sulfate sulfur concentrations are low in the rock samples, indicating minimal in situ oxidation of the sulfides. This is consistent with the geological model of a shallow oxidation front.

•The ratio of neutralisation potential to acid potential (NP:AP) is used to indicate the potential for net acid generation from a static test. Approximately 75% of the samples are considered non-acid generating based on the NP:AP ratio or the low sulfur content. Approximately 25% of the samples could be considered potentially weakly acid generating, however given the relatively low sulfide content this may represent only local zones of potential net acidity.

This is consistent with the baseline observations of generally circum-neutral water quality in the project area.

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Selected samples were tested using various short-term leach extraction tests to provide an indication of potential metal leaching from these samples. These tests are designed as ‘batch’ or instantaneous tests to maximise dissolution of metals from a sample; these tests can overestimate actual drainage water chemistry in the longer term. The short-term filter extraction tests were used to indicate the potential for metal leaching for the range of rock samples encountered in both waste rock and low grade materials. Sample results indicated that certain units of waste rock may have leachable aluminium, cadmium, copper, lead, and zinc where lower pH values occur.

The static tests and the evaluation of the ICP database confirm that the samples selected cover the range of expected sulfide concentrations in the mining material. On-site materials have a low neutralisation potential. Therefore, the classification of materials is primarily a function of the content of sulfur and metal. These results indicate that most of the waste rock has low potential for acid drainage and metal leaching, mainly due to relatively low sulfide and mineralisation outside the ore zones.

A combination of sulfide, zinc, and paste pH are used to identify waste rock that is a potential source of metals leaching or acid drainage. These parameters are included in the mine block model and are used for the design of the waste rock handling.

The mine block model is used to manage the waste rock according to the net acid generation potential and/or metal leaching potential in the waste rock storage areas. This is accomplished through segregation of potentially reactive waste rock (Class ‘A’) placement in the dumps with contained drainage. These waste rock storage areas have controlled drainage and, in the long term, can be directed to the open pit if necessary. Non-reactive waste rock (Class ‘B’ or ‘C’) are placed separately further downstream in the catchment.

A rock sampling programme is projected on 2022 in order to update the geochemical information. The programme will include kinetic acid rock drainage testing of the rock.

17.1.4 Water Management

During the Project life, water quantity and quality is managed to maximise diversions and maintain ‘non-contact’ water. The site water management plan is designed to ‘keep clean water clean’ as much as possible. Diversion ditches have been designed around the dumps, pit and stockpiles to convey clean or non-contact freshwater around these disturbed areas, where it is physically practical. The ‘Class A’ Rock Storage Area stores potentially reactive rock and is located such that it can drain into a contact water pond (A Pond – Contact Water), to allow monitoring and batch treatment if required before discharge. Currently the contact water is used for dust control given that the quality is appropriate for this use.

Water that accumulates on Project infrastructure is collected for settling and testing prior to any discharge. No water will be discharged to the environment that would have adverse environmental impact.

The dewatering and water management plan is comprised of three systems:

•Diversion ditches

•Pit groundwater dewatering sumps (non-contact water)

•Surface contact water ru-noff dewatering

Water collected within the catchments of the open pit and each waste rock dump area are directed to two ponds constructed at the low point of each area. The water of both ponds in used for dust control. A general arrangement of this system is included in Figure 17.2. The contact water diversion channels water is currently used for dust control. The non-contact water diversion channels water flows are drained to Uquillayoc river.

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Figure 17.2 Project General Arrangement and Water Management Features

MPSA, 2021

The contact water diversion channels are shown in red; this water is currently used for dust control. The non-contact water diversion channels are shown in green; these water flows are drained to Uquillayoc river.

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17.1.5 Flora and Fauna

The Chinchillas property area is a mix of high Andean plains and Puna landscape, characterised by grassy steppes and low-growing shrub land (Figure 17.3 and Figure 17.4), interspersed with bare soil and alkaline wetlands (peladares). Where standing water is encountered, such as at ponds and streams, surrounding wetland vegetation are collectively known as ‘vegas’, dominated by the families Juncaceae, Cyperaceae, Poaceae, Oxalidáceas and Scrofulariaceas (Figure 17.5). In upland drier zones, cactus such as Maihueniopsis and Lobivia can be found.

Figure 17.3 Grassland Steppes on the Western Edge of the Project Area

MPSA, 2021

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Figure 17.4 Shrub Land on the Northern Edge of the Project Area

MPSA, 2021

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Figure 17.5 Vega Habitat

MPSA, 2021

The effects of the high-altitude environment include increased solar radiation, constant winds, and large temperature fluctuations. Soils are typically young with low levels of organic material. These conditions have influenced the development of plant species in this area, where species of different families often show similar morphologies. Grasses typically have a high proportion of cellulose and lignin for added rigidity, and extra layers of cutin or suberin to restrict water loss. Woody plants are typically found as shrubs, with almost no tree layers evident.

Fauna of the Project area are highly correlated to wetter and humid areas, including the vegas. Several species of insects have been recorded, along with three species of amphibians. Three species of reptiles (two lizards and one snake) have also been documented in the area.

There are at least 72 species of birds known to be present for at least part of the year in the Project area. The most abundant of these are the Ash-breasted Sierra Finch (Phygilus atriceps) and the Bright-rumped Yellow Finch (Sicalis uropygialis). Other birds in the area of note include the Andean Condor (Vultur gryphus), the Ornate Tinamou (Nothoprocta ornate), the Puna Rhea (Pterocnemia tarapacensis), the Mountain Parakeet (Bulborhynchus aurifrons), and the Bare-faced Ground Dove (Metriopelia ceciliae).

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Studies completed in 2015 identified nine native and one exotic mammalian species. Numerous domestic species (e.g., llamas) were also noted in the area. The most common native mammals were the Vicuña (Vicugna vicugna) and the Vizcacha (Lagidium viscacia).

Some displacement of vegetation communities and attendant wildlife habitat has occurred within and adjacent to the Project footprint as a result of project development. These impacts have been assessed and approved in the Resolution No. 014/2017 (DIA Mina Chinchillas). As a result of the approval of the EISA of Chinchillas Mine, since 2018 MPSA developed a biannual Community Environmental Monitoring Program that includes fixed monitoring stations of flora, fauna, and limnology, as well as water, air, and soil quality.

17.1.6 Protected Areas

There are 15 protected areas within the Province of Jujuy, however the majority of these are far removed from the Project area. The Laguna de Pozuelos represents the most important protected area within the Chinchillas property region.

The Laguna de Pozuelos is a large, permanent, high-altitude lake located approximately 25 km from the Project area. It is an important migratory bird stopover, particularly known as habitat for the Andean Flamingo, as well as many other species.

The Laguna is located within a National Natural Monument, protected by the ‘Administracion de Parques Nacionales’ (National Parks Administration) as well as a United Nations Educational Scientific and Cultural Organisation (UNESCO) designated Biosphere Reserve and RAMSAR Wetland of International Importance. The National Natural Monument covers a surface of approximately 16,000 ha and in this area all economic activities, including mining, are prohibited.

The National Natural Monument is surrounded by a buffer zone of approximately 380,000 ha defined as a RAMSAR Wetland of International Importance that is administered by the multi-sector organisation ‘Corporación para el Desarrollo de la Cuenca de Pozuelos’ (CODEPO: Corporation for the Development of the Pozuelos Watershed) that is responsible for promoting sustainable development in the buffer zone. This buffer zone is recognised by UNESCO, who note that one of the objectives of the Reserve buffer zone is to make development compatible with conservation (www.unesco.org).

As shown in Figure 17.6, the Jujuy Ministry of Mining GIS data indicates that the Chinchillas property is located just inside the buffer zone, while boundaries provided by the University Nacional de Jujuy (UNJ) follow the UNESCO model and divide the buffer zone into an outer transition zone, with the Chinchillas property located outside of both zones. Taking the Ministry data of the buffer outline as the most recent and correct suggests that Chinchillas falls within the Ministry buffer zone, and within the UNESCO transition zone. In either case, economic activities, including mining and exploration, are permitted in these areas.

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Figure 17.6 Laguna de los Pozuelos Buffer Zones

MPSA, 2021

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17.2 Social and Community Engagement

17.2.1 Local Communities

The Project is located in a rural area in the department of Rinconada in the province of Jujuy. The Rinconada department has an area of 6,407 km2 and a population of only 2,489 (2010 Census). The department is divided into two municipalities; Rinconada Municipality and Mina Pirquitas Municipality.

The nearest population centres to the Project include the village of Santo Domingo (approximately 6 km distant) and the larger city of Abra Pampa (approximately 75 km distant), which is located in the adjacent department of Cochinoca. Additionally, there are four villages located between the Chinchillas site and the Pirquitas Operation; Liviara (approximately 9 km distant) Orosmayo Grande and Orosmayo (approximately 14 km distance), Coyaguayma (40 km distance) and Nuevo Pirquitas (approximately 29 km distance). Each community is considered aboriginal communities, with predominant Colla ethnicity. Colla people historically occupied the high Puna regions throughout northern Argentina, western Chile, and southern Bolivia. They traditionally speak a dialect of the Quechua language.

It is estimated that 78 people live in Santo Domingo, the village most proximate to the Project. A further 60 people are estimated to live dispersed throughout the surrounding area. Similarly, an estimated 17 people live in Liviara, 12 in Orosmayo, 25 in Osormayo Grande and 116 in Nuevo Pirquitas. Abra Pampa, the largest urban area in the region, has a population of approximately 16,000.

The livelihood of the area’s population is primarily tied to small-scale livestock management, typically goats and llamas, with some limited production of sheep. Sale of livestock, meat, and wool is typically done in Abra Pampa, from where it may eventually reach markets farther afield such as San Salvador de Jujuy.

Outside of agriculture, regional inhabitants are employed by the public sector (e.g., schoolteachers), or work in the mining industry. Many local rivers are exploited for low volumes of placer gold, and several hard rock mines, including the Pirquitas mine, have operated in the area. The majority of workers from Liviara and Orosmayo are employees of the Pirquitas mine.

Currently MPSA employs or hires 149 workers that come from direct and indirect communities (ESIA Mina Chinchillas 2021).

17.2.2 Archaeology

The Puna region of Argentina has a rich history of occupation, dating from at least 10,000 years before present. Hunter gatherers roamed throughout the region, gradually domesticating llamas and moving to greater reliance on agriculture within the last 3,000 years. The Incas arrived in the region in 1475, which had a great effect on the social order and use of resources. Spanish conquistadors arrived in 1535, further altering the socio-economy of the area and ushering in the colonial era.

Mining occurred historically at the Chinchillas area on a small scale in the eighteenth century by Jesuit missionaries. In the late 1960’s, there was a period of small underground production by a local company using adits and tunnels.

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An archaeological survey was conducted at the Chinchillas property in 2015. A total of 11 archaeological sites were identified proximate to the project itself. Other sites were identified in the surrounding area totalling 31 finds.

Prior to the start of exploitation of the Chinchillas Mine, in February 2018 the archaeological clearance of 15 sites that were going to be affected by the mine facilities was carried out (approved in Resolution No. 014/2017 - DIA of the Chinchillas Mine). The final clearance was obtained by Resolution No. 453/2018 issued by Cultural and Tourism Ministry. In April 2019 an additional clearance of a historical sites carried out under the authorization of Resolution No. 151/2019. The remaining sites are being protected by the company and are subject to annual monitoring.

17.3 Project Permitting

The legal framework for mine permitting is derived mainly of the second section of the Mining Code of the Nation and its supporting National Law No. 24.585. The institutional Framework for the permitting process is driven by stipulations in Law No. 24.585, with technical support of UGAMP and the National Mining Secretariat.

The main focus of permitting is the detailed Environmental and Social Impact Assessment, which must be submitted prior to commencement of operations. Upon successful review of the ESIA, a DIA is awarded. Annex III of Law 24.585 establishes the minimum contents of the EIA, which must include:

•Description of the Environment (physical, biological, and socio-economic);

•Project Description;

•Description of Environmental Impacts;

•Environmental Management Plan (which includes measures and actions to prevent and mitigate environmental impact);

•Plan of Action on Environmental Contingencies; and

•Methodology Used.

An ESIA for the Project was developed and submitted for review in September 2016. which was subject to review by the Mining Department and UGAMP and approved on 22 December 2017 by Resolution No. 014/2017. It is subject to review by the Mining Department and UGAMP, a process that is expected to conclude with issue of a DIA in mid to late-2022 . Since then, two more ESIA Updates have been developed and submitted to Mining Authorities, ESIA Mina Chinchillas 2019 and ESIA Mina Chinchillas 2021. Both are being reviewed by Mining Authorities and it is expected to have the final DIA soon.

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The UGAMP is a multi-stakeholder group chaired by a technical appointee from the Mining Department who recommends approval or rejection of the ESIA and related work application to the provincial mining authorities. Meetings are held to allow UGAMP members to review the proposed materials with members of Golden Arrow. UGAMP representatives appurtenant to the Project include:

•Representatives from the local Communities of Santo Domingo, Orosmayo, Liviara, Orosmayo Grande, Nuevo Pirquitas and Coyaguayma;

•Mining Workers Unions;

•Provincial Department of Water Resources;

•Department of Mines and Energy;

•Provincial Secretary of Mining;

•Surface Landowners;

•Provincial Collage of Geologists;

•Provincial Department of Environment;

•Provincial Department of Human Rights and Indigenous Communities;

•National University of Jujuy;

•Jujuy Chamber Mining;

•National Parks Administration;

•Corporation for the Development of the Pozuelos River;

•Provincial Secretary of Public Health;

•Provincial Department of Agriculture and Livestock Control; and

•Provincial Department of Industry and Commerce.

Chinchillas has maintained all previous exploration activity permits in good standing, each of which required the submission of an ESIA and receipt of a DIA. As the review of the mining ESIA proceeds, precedent suggests that the DIA will also be granted.

The use of the Pirquitas pit for tailings deposition at the Pirquitas Operation is a modification to the mining activities not contemplated in MPSA’s ESIA until 2016 for the Pirquitas mine. In August 2017, MPSA issued to Mining Authorities an Addendum of the 2016 ESIA Update that included the upgrades to conduct the tailings to the pit of Mina Pirquitas. The permit was obtained on 24 September 2018 by Resolution No. 056/2018. Since then, MPSA has submitted to Mining Authorities the ESIA Update for Mina Pirquitas in September 2020, which is under review.

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17.4 Mine Closure

A conceptual closure plan and cost has been developed for the Project. There are no specific laws in Argentina that specify mine closure requirements, and there is no bonding requirement. The closure plan for the Project has been developed in consideration of best industry practice. The closure plan was designed to accommodate the following objectives:

•Health and security of the public

•Protection of the environment

•Ensure physical and chemical stability of post-closure structures

•Ensure unrestricted and unimpacted natural surface water flow

•Prevent erosion of post-closure structures from wind or water

•Safe removal of impacted surface structures and buildings

•Safety and security for people, wildlife, and livestock

17.4.1 Closure Activities

Buildings and surface structures will be cleaned of residual fuels, lubricants, reagents, and wastes prior to being deconstructed and dismantled. Recyclable wastes will be reused wherever possible. All structures will be removed to ground level, with concrete slabs or other inert foundations covered with stored topsoil. All access roads to the pit and waste rock storage areas will be blocked for safety using earthen berms accompanied by warning signs.

The water diversion systems employed during operations will be fortified for long term use in managing water post-closure. This will include maintaining all upgradient run-off as non-contact water passed downstream to the Arroyo Uquillayoc.

The pit will be allowed to flood to the phreatic level. A large safety berm accompanied by appropriate signage will be constructed around the pit rim to prevent access.

Ongoing monitoring of the closure measures will be conducted over a period of five years to ensure successful implementation. Due to the fact that the exploitation and mineral extraction stage of the Pirquitas pit has ended, some components of the Pirquitas Mine are currently in the mine closure stage, so the activities currently being carried out are those linked to the ore processing and mine closure. Closure costs for Chinchillas mine have been estimated at $30.6M. MPSA is reviewing these costs and suggested that the closure costs will be lower than this.

17.5 Pirquitas Mine

17.5.1 In-Pit Tailings Disposal

MPSA is currently using the Pirquitas pit as tailings reservoir. These tailing comes from the processing of the Mina Chinchillas ore. The use of the pit as a tailings reservoir was approved by the Authority through Resolution No. 056/2018, after submitting the Addendum to the Authority in August 2017.

Placing the tailings inside the pit involves transporting them from the process plant to the pit by means of a pumping system and a 6.3 km pipeline to the tailings box located on the edge of the pit, to be discharged to the pit. Likewise, the tailings reservoir has a water recovery system to pump the water (from the tailings, the flows that enter the pit by filtration, direct rain, and surface run-off) to the process plant for reuse. This pipe follows the same route as the pipe that transports the tails. The disposal of the tailings in the pit began in April 2019.

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Due to the fact that the exploitation and mineral extraction stage of the Pirquitas pit has ended, some components of the Pirquitas mine are currently in the mine closure stage, so the activities currently being carried out are those linked to the ore processing and mine closure.

17.5.2 Pirquitas Pit

Pirquitas pit within the same watershed as the Homonymous River. During the operation of Mina Pirquitas, the surface run-off influent to the pit was captured and conveyed along the southern perimeter of the pit to a discharge point in the Pircas River, downstream of the pit. The main tributaries to the pit correspond to the Pircas and Maray streams (contacted waters in the Pircas dump), and the Médanos stream (waters not contacted), the convergence of these three channels naturally formed the Pircas River, precisely in the pit sector.

In January 2017, with the cessation of mining in Pirquitas, the mine dewatering ceased and surface run-off from the upper basins was directed towards the pit. As a result, the pit lagoon increased in volume until it reached approximately 40% of the depth of the pit. The increase in volume led to various hydrogeological studies and models, to establish the design criteria for the tailings reservoir.

The studies concluded that the open pit acts as a sink and can contain the tailings and their processes associated with the proposed methodology for the disposal of the tailings.

As a result of the studies, a critical maximum elevation of the water level has been determined at 4,207.7 masl, where the pit would cease to be a sink. Likewise, the overflow level of the pit was established at 4,230 masl.

To maintain the required water levels for the tailings reservoir, it was necessary to construct a water management system to avoid the inflow of surface water to the pit. These works were completed in February 2020.

Based on the operational and meteorological conditions, the pit water balance is annually updated.

17.5.3 Environmental and Social description and Closure

17.5.3.1 Water Quality

In December 1998, consulting engineering firm KP completed an ESIA for Sunshine Argentina. The ESIA contained a description and evaluation of environmental conditions that existed at the time, as well as foreseeable potential effects that development of the Pirquitas mine could have on the surrounding environment. The scope of the ESIA was commensurate with the norms for environmental protection associated with Argentina’s applicable mining laws and guidelines established by international lending institutions such as the World Bank. The discussion below is either paraphrased or taken directly from the ESIA, with updates to include information about the Pirquitas mine subsequent to the date of such ESIA.

Remnants of historical mining activities at the Pirquitas mine included derelict buildings, mine structures and tin-silver jig tailings and tin placer tailings along the Río Pircas. Flotation tailings had been discharged into the Río Pircas and piles of gold placer tailings were left above the current level of the Río Pircas on paleo-river terraces near the mine camp. These areas comprise some 107 ha of surface disturbance that existed prior to Sunshine Argentina’s acquisition of the property, some of which are now associated with acid rock drainage into the Río Pircas watershed.

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Surface and ground waters are known to be acidic and metalliferous down gradient from the historical mines above the Río Pircas canyon at Tres Placas, which is located downstream from the Pirquitas pit. In addition, acidic and metalliferous ground water is present in the abandoned underground workings and some natural springs in the area, suggesting natural oxidation of sulfide mineralisation which is widespread in the rocks found on the property is also contributing to background surface water contamination.

Furthermore, the only condition the Argentina Ministry of Mines and Energy applied to its approval of Sunshine Argentina’s ESIA, apart from the mandatory two-year update to the report, was the requirement that water quality monitoring be carried out.

MPSA is currently monitoring the water quality both upstream and downstream of the mine. The monitoring programme includes 21 sites for superficial water and eight sites for groundwater monitoring. The general characterisation of the water continues the same as the original in 1998.

17.5.3.2 Flora and Fauna

In the area of influence of Mina Pirquitas, the Altoandina and Puna ecoregions are distinguished. The physiognomy of the climax vegetation corresponds to an herbaceous (High Andean) or shrubby (Puna) steppe, however, it is possible to find a mixed ecotonal community between these two ecoregions in nearby environments.

The plant physiognomy in the Puna consists of shrubby and gramineous steppes or grasslands, with low plant cover and extensive areas of bare areas. In the sectors associated with wetlands, such as lagoons and streams, there are springs, with vegetation dominated by grasses such as Festuca sp. (chillagua), sedges such as Scirpus sp., and rushes that completely cover the ground, constituting a privileged habitat for being sites where there is a high concentration of biodiversity. In the rocky areas there are species of cacti of the Maihueniopsis and Lobivia genera.

The species, both herbaceous and shrubby, have the shape of cushions (camephytes or hemicryptophytes), and settle on the ground in a scattered manner, leaving areas of bare soil. The physiognomy of the vegetation resembles a high altitude desert; however, there are endemic species and others that only appear in the rainy season (late-summer), which provides a significant richness of species. The families best represented in these environments are: Poaceae, Asteraceae and Solanaceae

Given that in the Mina Pirquitas area, there is great variability in terms of the floristic composition of the vegetation units, the area of influence was subdivided into six sub-basins. These areas are monitored twice a year since 2011.

The area of Mina Pirquitas and its immediate surroundings in Jujuy, has a high conservation value, mainly of Puno and high Andean habitats, but also in terms of its biodiversity.

The upper Pilcomayo basin, represented by the Pirquitas area, constitutes, together with the adjoining endorheic basins of Pozuelos and Vilama, a considerable area of outstanding quality of this ecosystem.

The faunal indicators selected to carry out fauna monitoring did not show, at a general level, tangible changes in the richness of species in the area and in the structure of the communities, which at this scale shows that the impacts of the activities carried out in Mina Pirquitas for the groups of fauna involved, they are stable or compensated on this spatial and temporal scale.

As a result of the approval of the EISA of Pirquitas mine, since 2018 MPSA have developed a biannual Community Environmental Monitoring Program that includes fixed monitoring stations of flora, fauna, and limnology, as well as water, air, and soil quality

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17.5.3.3 Local Communities

The Area of Direct Influence is made up of the localities that have the greatest connection with the mine, either due to geographical proximity, because of the employment relationship of a large part of the active labour population, or because of the community relationship programmes that are implemented jointly, with the company.

The localities are: Nueva Pirquitas (approximately 4.5 km), Loma Blanca (approximately 50 km distance), Coyaguayma (14 km), Orosmayo- Orosmayo Grande (25 km) and Liviara (30 km).

The linkages that are carried out between the locations of the AID and the mine are fundamentally due to the proximity that exists between the operational areas and the communities, where they share communication routes and road infrastructure; Part of the workforce of the mine is from nearby towns and community investment policies are executed by the Company, which is headed by the Municipal Commission of Nueva Pirquitas.

The highest percentage of personnel working at Mina Pirquitas S.A belongs to the Province of Jujuy. That is, 163 people belong to the province of Jujuy, 96 people to the communities of the Area of Direct Influence, 17 people to the communities of the Area of Indirect Influence (AII).

17.5.3.4 Pirquitas Permitting

In 1998, the Original ESIA of the Pirquitas Project was developed by KP firm, which was prepared in accordance with the requirements of the National Law for Mining Environmental Protection, Law No. 24,585 and other substantive and formal regulations in force. The document was approved by Resolution No. 16/99. However, the mining activities provided for in the 1998 ESIA for the exploitation of the mine were not started.

In 2005, a new stage of exploration began in order to identify new mineralised areas. This year, the primary focus was a geological reconnaissance of the Oploca veins, with the aim of expanding in-depth geological and mineralogical knowledge.

The first Update of the ESIA was delivered to authorities in 2008 and was approved by Resolution No. 35/08. In June 2008, the pre-production stage began, starting the production of concentrates on April 6, 2009.In 2008, a second ESIA was completed by KP following start-up of mining activities and initiation of plant construction. While there were no observations or restrictions placed on MPSA at that time, this study began to focus on the water management plan and conceptual plans for mine waste stockpiles. A conceptual water treatment plant for neutralisation of acid waters was proposed as a contingency with a treatment capacity estimated to be as much as 150 L/s. Alternative water management measures to date have reduced the source of acidic waters, and such treatment plant has not yet been required.

A party wishing to commence or modify any exploration or mining-related activity under Argentina’s mining laws, including property abandonment or mine closure activity, must prepare and submit an ESIA, which must include a description of the nature of the proposed work, its potential risk to the environment and the measures that will be taken to mitigate that risk. The most recent update permit to MPSA’s ESIA for the Pirquitas mine, which included engineering studies for the design of water management structures and mine closure design, was submitted in December September 2016 and the addendum for in-pit disposal was submitted in August 2017. These ESIA’s were approved in September 2018 by Resolution No. 056/2018.

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The preceding update was submitted in December 2014 and formally approved in January 2016. An addendum to this ESIA regarding the closure of the Pirquitas mine was filed in December 2015, which reflected the revised mine plan projecting the completion of the Pirquitas pit, with lower grade stockpile processing expected to commence upon cessation of open pit mining activities at the Pirquitas pit. In July 2016, an updated closure plan, which included more detailed engineering of the selected closure measures and costing for both active closure and longer term care and maintenance, was submitted to the regulatory authorities. Due to the approval of Mina Chinchillas ESIA in December 2017 (Resolution No. 014/2017) the closure plan for Pirquitas mine was archived and any work proposed on this closure plan in currently submitted in the ESIA Updates every two years. and is currently under review.

The most recent ESIA submitted to authorities is the ESIA 2020 Update which is currently on revision under review by mining authorities.

The cessation of open pit mining activities at the Pirquitas pit in January 2017 has resulted in a significant reduction in workforce, as well as reduced indirect economic benefits to the surrounding and supporting communities with the start of exploitation of the Chinchillas mine, new contracts were made at Mina Pirquitas as of 2018 and currently MPSA hires 840 workers between Chinchillas and Pirquitas mines. A social impact assessment study was commissioned in 2015 and formed the basis of the social closure plan for the Pirquitas mine. The potential risks, as well as actions to reduce those risks and support the employees and the community, were developed as part of the reclamation and closure plan submitted in 2016 .

17.5.4 Closure

Argentina currently has no specific mine closure legislation other than the requirement to prepare and submit and regularly update an ESIA, including with respect to mine closure activity. However, it is expected that closure options will be proposed as part of the review of MPSA’s updated closure plan and may include passive or active neutralisation features to return discharged waters to baseline conditions (acidic at the time of baseline studies) with monitoring requirements. The closure requirements for the Pirquitas pit may change in the future and MPSA may be subject to increased obligations for both the technical and social aspects associated with such mine closure and reclamation, which would impact the closure plan and the duration of the associated closure activities.

The current closure and reclamation plan addresses a range of closure risks, design criteria and costs that are anticipated in order to comply with internationally accepted practices. It considers both the physical reclamation of the site and the social closure plan for the neighbouring communities for whom the mine provides employment and community support. The closure plan considers the short-term decommissioning and reclamation measures, as well as longer term care and maintenance activities and related costs and risks. The actual costs of reclamation and mine closure are uncertain and planned expenditures may differ from the actual expenditures required. Therefore, the amount required to be spent could be materially higher than current estimates.

MPSA is developing an update of the Puna closure plan that will includes both Chinchillas and Pirquitas mines to possible changes to the closure requirements and obligations.

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17.6 QP Opinion

Legal matters such as statutory and regulatory interpretations affecting the mine plan and environmental matters are outside the expertise of the QPs and are within the control of the registrant (see Section 25).

Following a review of the information supplied, the opinion of the QPs is that it is reasonable to rely on the information provided by SSR as outlined above for use in the Puna21TRS because a significant environmental and social analysis has been conducted for the project over an extended period, the project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

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18 CAPITAL AND OPERATING COSTS

18.1 Capital Costs

The Project utilises the existing processing facilities at the Pirquitas Operation, therefore most capital items are related to the mining equipment and infrastructure required at the mine site.

The estimated capital costs required to achieve the Mineral Reserves LOM are summarised in Table 18.1. The capital costs were estimated by the MPSA using actual costs and other information.

As the project has been in operation for a number of years, the level of project definition for the capital cost estimate is very high. Given the remaining capital scope, and the level of project definition, no contingency was included in the cost estimate. The QPs consider the capital estimate to be in the accuracy range of +/–15%.

The sustaining capital costs include:

•Surface infrastructure construction such as upgrades to the camp and kitchen, IT upgrades, and asset integrity costs.

•Mill improvements and replacement of major components.

•Tailings management facility costs.

•Mobile equipment such as new and replacement purchases and major rebuilds.

Table 18.1 Capital Costs Estimate

Cost Component	$M
Exploration and Development	21
Sustaining Capex	47
Closure and Reclamation	31
Total Capital Cost	99

Capital includes only direct project costs and does not include non-cash shareholder interest, management payments, foreign exchange gains or losses, foreign exchange movements, or tax pre-payments.

18.2 Operating Costs

Operating costs are estimated using current operating experience at Pirquitas operation, actual quotes from vendors and first principles. Operating costs are estimated by MPSA for the areas such as mining, processing, tailings and general and administration.

As the project has been in operation for a number of years, the level of project definition for the operating cost estimates is very high. Given the available project performance data and the high project definition, no contingency was included in the cost estimate. The QPs consider the operating cost estimate to be in the accuracy range of +/–15%.

The operating expenses estimated to validate the positive cash flow for the Mineral Reserves LOM. The LOM operating costs are approximately $52.67/t of ore milled, as are summarised in Table 18.2. The mining expense includes all labour, supplies / consumables, and equipment maintenance to complete mining related processes / activities, less exploration diamond drilling and capital excavations and construction. The milling expense includes all labour and supplies / consumables to complete milling related processes / activities. The administrative expense includes all labour, supplies / consumables, and equipment maintenance to complete administrative, finance, human resources, environmental, safety, supply chain, site services, camp, and kitchen, and travel related processes / activities.

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Table 18.2 Operating Costs Estimate

Cost Component	Total<br>($M)	LOM Average<br>($/t milled)
Mining	110	15.01
Processing	183	24.95
G&A	93	12.71
Total Operating Costs	387	52.67

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19 ECONOMIC ANALYSIS

19.1 Economic Assumptions

The modelling and taxation assumptions used in the Puna21TRS are discussed in detail below.

All monetary figures expressed in this report are in US Dollars ($) unless otherwise stated. Cash flows are assumed to occur evenly during each year and a mid-year discounting approach is taken. The estimates of cash flows have been prepared on a real basis as 1 January 2022 and a mid-year discounting is used to calculate net present value (NPV).

19.1.1 Pricing and Discount Rate Assumptions

Metal price assumptions are shown in Table 19.1. Other key assumptions in the economic modelling relating to product pricing are tabulated in Table 19.2. A discount rate of 5% is used for calculating net present value (NPV).

Table 19.1 Metal Price Assumptions

Commodity	Unit	2022	2023	2024	2025	Long-Term
Silver	$/oz	24.00	23.00	22.00	21.00	21.00
Lead	$/lb	1.00	0.95	0.93	0.92	0.90
Zinc	$/lb	1.30	1.20	1.20	1.20	1.20

Table 19.2 Key Economic Assumptions

Item	Unit	Lead Concentrate	Zinc Concentrate
Treatment Charge and Refinery Charge(TCRC)	$/t Conc.	1,191	724
Payability – Silver	%	95	75
Payability – Lead	%	95
Payability – Zinc	%		85
Deduction – Lead	%	3
Deduction – Zinc	%		8
Minimum Payout Factor	%	63	39
Royalty	%	3	3
Export Duty (revenue minus TCRC's)	%	4.5	4.5
Puna Credit (revenue minus TCRC's)	%	2.5	2.5

In the analysis, carry balances such as tax and working capital calculations are based on nominal dollars and outputs are then deflated for use in the integrated cash flow calculation.

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19.1.2 QP Opinion on Inputs

Data and assumptions for macroeconomic trends, taxes, royalties, interest rates, marketing information and plans, legal matters such as statutory and regulatory interpretations affecting the mine plan, and environmental matters are outside the expertise of the QPs and are within the control of the registrant (see Section 25).

Additionally, the project has been in operation for a number of years, and following a review of the current supplied information, it is the opinion of the Puna21TRS QPs that the current plans and input parameters appear adequate for use as inputs to the Puna21TRS.

19.2 Overview and Results

The estimates of cash flows have been prepared on a real basis as at 1 January 2022 and a mid-year discounting is used to calculate NPV.

The projected financial results include:

•After-tax NPV at an 5% real discount rate is $228M

•Mine life of five years

The estimated total cash costs for the LOM is $11.63/oz silver. The all-in sustaining costs (AISC), which includes infrastructure capital, capital development and reclamation, average for the LOM is $13.57 per payable ounce of silver sold. Unit costs include concentrate in stockpile. Silver provides the primary revenue for the analysis, with contributions from lead and zinc. Credits from lead and zinc are included in the cash cost.

The key results of the Puna21TRS are summarised in Table 19.3.

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Table 19.3 Puna21TRS Results Summary

Description	Unit	Total LOM
Ore Processed
Ore Tonnes Treated	kt	7,352
Ag Feed grade	g/t	160
Pb Feed grade	%	1.32
Zn Feed grade	%	0.29
Silver Recovery	%	95.5
Concentrates
Lead Concentrate – in Stockpile	kt	4
Zinc Concentrate – in Stockpile	kt	1
Lead Concentrate – Produced	kt	135
Zinc Concentrate – Produced	kt	27
Lead Concentrate – Total	kt	139
Zinc Concentrate – Total	kt	28
Metal Produced
Silver	koz	37,210
Lead	Mlb	204
Zinc	Mlb	29
Key Financial Results
Mine Site Cash Cost	$/oz payable silver	11.61
Royalties and Refining Costs1	$/oz payable silver	6.10
Credits	$/oz payable silver	–6.08
Total Cash Costs (after credits)1	$/oz payable silver	11.63
All-in Sustaining Costs (AISC)	$/oz payable silver	13.57
Site Operating Costs	$/t milled	52.67
Average Silver Price	$/oz	22.38
NPV1	$M	228
Discount Rate	%	5
Project Life	years	5

Metal produced includes current concentrate stockpiles containing 242 koz silver and 5 Mlb lead.

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19.2.1 Production and Cost Summary

The process production forecasts are shown in Table 19.4 and forecast tonnes mined are shown in Figure 19.1. The processing tonnes and metal production are summarised in Figure 19.2 and Figure 19.3 respectively.

Table 19.4 Production Forecast

Item	Unit	Total LOM
Ore Processed
Ore Tonnes Treated	kt	7,352
Ag Feed Grade	g/t	160
Pb Feed grade	%	1.32
Zn Feed grade	%	0.29
Silver Recovery	%	95.5
Concentrate Produced
Lead Concentrate – in Stockpile	kt	4
Zinc Concentrate – in Stockpile	kt	1
Lead Concentrate – Produced	kt	135
Zinc Concentrate – Produced	kt	27
Lead Concentrate – Total	kt	139
Zinc Concentrate – Total	kt	28
Metal Produced
Silver	koz	37,210
Lead	Mlb	204
Zinc	Mlb	29

Metal produced includes current concentrate stockpiles containing 242 koz silver and 5 Mlb lead.

Figure 19.1 Mining Production Profile

OreWin, 2021

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Figure 19.2 Process Feed Profile

OreWin, 2021

Figure 19.3 Silver Production

OreWin, 2021

The estimated mine site cash costs are shown in Table 19.5. The estimated total cash costs for the LOM is $11.63/oz payable silver. The AISC, which includes infrastructure capital, capital development, and reclamation, average for the LOM is $13.57/oz payable silver. Silver provides the primary revenue for the analysis, with contributions from lead and zinc. Credits from lead and zinc are included in the cash cost.

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These estimated costs include only direct operating costs of the mine site, namely:

•Mining

•Processing

•General and administrative (G&A) costs

•Government fees and charges (excluding corporate taxation)

The projected financial results include:

•After-tax net present value (NPV) at an 5% real discount rate is $228M

•Mine life five years

Table 19.5 Cash Costs

Item	LOM Average<br>($/oz Ag)
Mine Site Cash Cost	11.61
Royalties and Refining Costs	6.10
Credits	–6.08
Total Cash Costs (CC) (after credits)	11.63
All-in Sustaining Costs (AISC) (after credits)	13.57

Note: Includes concentrate in stockpile

The estimated revenues and operating costs have been presented in Table 19.6, along with the estimated net sales revenue value.

The metal prices used for the economic analysis are shown in Table 19.1. The metal prices used in this Puna21TRS are based on an SSR internal assessment of recent market prices, long-term forward curve prices, and consensus amongst analysts regarding price estimates. The metal prices selected for Puna Operations have taken into account the current project life.

Table 19.6 Operating Costs and Revenues

Description	TOTAL<br><br>($M)	LOM Average<br><br>($/t milled)
Revenue
Gross Sales Revenue	1,000	136.01
Less Realisation Costs
Treatment & Refining Charges	179	24.28
Royalties	36	4.89
Total Realisation Costs	214	29.17
Net Sales Revenue	785	106.84
Less Site Operating Costs
Mining Costs	110	15.01
Processing Costs	183	24.95
G&A Costs	93	12.71
Total	387	52.67
Operating Margin	398	54.17

Note: Includes concentrate in stockpile

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Table 19.7 Total Project Capital Costs

Item	Total<br><br>($M)
Exploration & Development	21
Sustaining Capex	47
Closure & Reclamation	31
Total Capital Cost	99

The projected financial results for undiscounted and discounted cash flows, at a range of discount rates are shown in Table 19.8.

The results of NPV5% sensitivity analysis to a range of changes in silver price (primary commodity) and discount rates is shown in Table 19.9. NPV sensitivity analysis for changes to operating and capital costs are shown in Table 19.9.

A chart of the after tax cumulative cash flow is shown in Figure 19.4 and details of the cash flow is shown in Table 19.11.

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Figure 19.4 After-Tax Annual and Cumulative Cash Flow

OreWin, 2021

Table 19.8 Financial Results

Discount Rate		NPV (M)
	Before-Tax
Undiscounted		279	253
2%		268	242
5%		253	228
10%		231	206
12%		223	199

All values are in US Dollars.

Note: Includes concentrate in stockpile

Table 19.9 After-Tax NPV Sensitivity to Silver Price and Discount Rates

After Tax NPV5%		Long-Term Silver Price (/oz Ag)
	10.00
Discount Rate		M	M	M	M	M	M	M	M	$M
Undiscounted		–17	105	192	204	253	278	327	401	474
2%		–17	101	183	195	242	266	313	384	454
5%		–16	95	172	183	228	250	294	360	427
10%		–14	86	156	166	206	226	266	327	387
12%		–14	83	150	160	199	218	257	315	373

All values are in US Dollars.

Note: NPV includes concentrate in stockpile

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Table 19.10 After-Tax NPV5% Sensitivity to Operating and Capital Cost Changes

Item		Changes to Cost<br><br>(%)
	–30%		–20%		–10%		–5%		–		+5%		+10%		+20%		+30%
Operating Cost		342		304		266		247		228		208		189		151		113
Capital Cost		238		234		231		229		228		226		224		221		217

Note: NPV includes concentrate in stockpile

Table 19.11 Estimated Cash Flow

Description
	(M)	(M)	(M)	(M)	(M)	(M)	(M)	(M)
Total Gross Revenue
Total Realisation Costs
Net Revenue
Site Operating Costs
Mining
Processing
G&A
Total Operating Costs
Operating Surplus / (Deficit)
Capital Costs
Exploration & Development
Sustaining Capex
Closure & Reclamation
Total Capital
Working Capital
Pre-tax Cash Flow
Tax Payable
After-tax Cash Flow

All values are in US Dollars.

Note: Table shows $M, includes concentrate in stockpile

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20 ADJACENT PROPERTIES

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21 OTHER RELEVANT DATA AND INFORMATION

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22 INTERPRETATION AND CONCLUSIONS

Data and assumptions for macroeconomic trends, taxes, royalties, interest rates, marketing information and plans, legal matters such as statutory and regulatory interpretations affecting the mine plan, and environmental matters are outside the expertise of the QPs and are within the control of the registrant (see Section 25).

Mineral Resources and Mineral Reserves in the Puna21TRS are reported in accordance with subpart 1300 of US Regulation S-K Mining Property Disclosure Rules (S-K 1300).

Significant factors that could materially affect the Mineral Resources and Mineral Reserves are:

•Environmental, Permitting Social and Community – Argentina currently has no specific mine closure legislation other than the requirement to prepare and submit and regularly update an ESIA, including with respect to mine closure activity. MPSA is developing an update of the Puna closure plan that will includes both Chinchillas and Pirquitas mines to possible changes to the closure requirements and obligations. In order to operate the mine, MPSA must maintain appropriate relations with all the authorities and stakeholders. Social, community and government relations are managed by MPSA and include programmes and engagement with the local communities and both local and national governments.

•Mine planning to maximise the current Mineral Resources is an important activity that MPSA has identified and commenced. Expediting this work will optimise the project and has the potential expand on the current project opportunities.

•Metal price impacts – silver is the primary revenue component and is produced from lead and zinc concentrates. Zinc prices have been relatively high compared to the long-term forecasts. As the operation has a short life evaluation of the prices for all metals will needed to maximise the value of extracted metal.

•Geotechnical impacts – the mine designs for Chinchillas were revised in 2021, a review by geotechnical engineers of the updated designs should be prepared to confirm that the designs are suitable for the current slope recommendations.

•The location of the project means that the concentrates are required to be transport ed a significant distance to customers. Delays or other issues pose a risk to revenue and MPSA needs to maintain planning and strategies to provide for an efficient logistics function.

•Closure of the processing plant may limit the development options of the Pirquitas underground Mineral Resource.

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23 RECOMMENDATIONS

Key recommendations for the project are:

•Potential remains to expand the current Mineral Resources and to define new Mineral Resources on the property.

•At Chinchillas it is recommended that MPSA examine advanced grade control (using reverse circulation drilling) at a grid spacing of 20 m, to determine if it will improve prediction particularly where the grade trends are horizontal. This examination should identify the targets and cost of the programme.

•The shallow eastward dip of high grades should be carefully managed by pit mapping and advanced grade control drilling to provide appropriate levels of confidence to manage risk. A detailed review of Socavon should be undertaken to determine whether portions may be amenable to economic extraction.

•Prepare a study to re-evaluate and assess the Pirquitas Mineral Resources and determine the development horizon available prior to the completion of the Chinchillas open pit and the impact of the current operation.

•Upgrade the Pirquitas density estimation method in future modelling.

•Conduct a review by geotechnical engineers of the updated designs should be prepared to confirm that the designs are suitable for the current slope recommendations.

•Undertake a geotechnical study of the waste rock dumps.

•Further pit optimisation using a range of metal prices and cost input parameters.

•Prepare additional detailed planning and design for rock storage and the general site layout.

•Prepare a geometallurgical study and design a testwork programme.

•Continue with ongoing review of capital and operating cost estimates and performance and productivity tracking.

•Finalise the update of the Puna closure plan and associated costs for Chinchillas and Pirquitas mines including analysis of the possible changes to requirements and obligations.

Costs for this work cost included in the cash flows and as the work will be primarily undertaken by site and other SSR personnel the costs are not considered significant extra costs above the budgeted operating and capital costs.

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24 REFERENCES

Armstrong, R. and Spratt, J., 2016. EPMA study of Ag-bearing Minerals from the Pb-Zn-Ag Mineralization of the Chinchillas Project. Unpublished report prepared by the Natural History Museum for Golden Arrow Resources Corporation.

Board W., Kennedy B. and Yeomans T., 2011. NI 43-101 Technical Report on the Pirquitas Mine, Jujuy Province, Argentina. 23 December 2011. Prepared for Silver Standard Resources Inc.

Blasco, G. G., 2011. Actualización bianual estudio de impacto ambiental etapa de exploración: Golden Arrow environmental impact assessment actualization report for exploration stage, 2011, 51 p.

Caffe P., 2002. Estilos eruptivos del complejo dómico Pan de Azucar – Puna Norte. Revista de la Asociación Geológica Argentina, 57 (3) 232-250.

Caffe P., 2013. Petrografía de muestras de testigo y roca. Mina Chinchillas. Unpublished report prepared for Golden Arrow Resources.

Caffe P. and Coira B., 2008. Depósitos epitermales polimetálicos asociados a complejos volcánicos domicos: Casa Colorada, Pan de Azucar, Chinchillas y Cerro Redondo, en: Relatorio del XVII Congreso GeológicoArgentino, JUJUY, 2008.

Canadian Securities Administrators (CSA), 2011. National Instrument 43-101 Standards of Disclosure for Mineral Projects. Retrieved from: https://www.bcsc.bc.ca/Securities_Law/Policies/Policy4/43-101_Standards_of_Disclosure_for_Mineral_Projects NI 43-101

Caranza, H., and Carlson, G. G., 2012. Final report on the phase I drill program Chinchillas Ag-Pb-Zn Project. Golden Arrow internal report, 2012, 1360 p.

CIM, 2003. Estimation of Mineral Resources and Mineral Reserves Best Practices Guidelines Retrieved from: http://web.cim.org/

CIM Standing Committee on Reserve Definitions, 2014. CIM Definition Standards - For Mineral Resources and Mineral Reserves. Retrieved from: http://web.cim.org/

Chen, S. and Redfearn, M., 2014. 2014 Project Report on Metallurgical Testing on the Chinchillas Project. Unpublished report prepared by Bureau Veritas Commodities Canada Ltd, Inspectorate Metallurgical Division, for Golden Arrow Resources Corporation.

Coira, B., Caffe P., Ramirez A., Chayle W., Diaz A., Rosas S., Perez A., Perez B., Orozco O. and Martinez M., 2004. Hoja Geológica 2366-I/ 2166/III Mina Pirquitas, Provincia de Jujuy. Servicio Geológico Minero Argentino, Boletin 269.

Coira B., Diaz A., Chayle W., Pérez A., and Ramírez A., 1993. Chinchillas, un modelo de complejo volcánico domico portador de depósitos de metales de base con Ag y Sn, en Puna Jujeña. XII Congreso Geológico Argentino y II Congreso de Exploración de Hidrocarburos. Actas Tomo IV (270-276).

Cunningham, C., McNamee J., Pinto J. and Ericksen G., 1991. A model of volcanic dome-hosted precious metal deposits. Econ Geol 86: 415-421.

Daroca, J. A., Undated. Lapacha S.R.L. report on Aranlee work at Chinchillas Project. Prepared for Aranlee Resources.

Davis, B. and Howie, K., 2013. Mineral Resource Estimate for the Chinchillas Silver-Lead-Zinc Project, Jujuy Province, Argentina. 20 June 2013. Retrieved from http://www.sedar.com/.

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Davis, B., Howie, K., and Smith, B., 2014. Mineral Resource Estimate for the Chinchillas Silver-Lead- Zinc Project, Jujuy Province, Argentina. 10 October 2014. Retrieved from http://www.sedar.com/.

Davis, B., Sim, R., and Smith, B., 2015. Mineral Resource Estimate for the Chinchillas Silver-Lead-Zinc Project, Jujuy Province, Argentina. 2 November 2015. Retrieved from http://www.sedar.com/.

Davis, B., Sim, R., and McEwen, B., 2016. Mineral Resource Estimate for the Chinchillas Silver-Lead- Zinc Project, Jujuy Province, Argentina. May 27, 2016. Retrieved from http://www.sedar.com/.

Glencore, 2015. Resources and Reserves as at 31 December 2016. Retrieved from http://www.glencore.com/assets/investors/doc/reports_and_results/2016/GLEN-2016-Resources- Reserves-Report.pdf

Golden Arrow Resources Corporation, 2015a. Silver Standard Plans up to US$12.6M to Advance to Feasibility Golden Arrow’s Chinchillas Project for a Business Combination with Pirquitas Mine [News Release 1 October 2015]. Retrieved from https://goldenarrowresources.com/news/2015

Golden Arrow Resources Corporation, 2015b. Management Information Circular. 20 November 2015. Retrieved from http://www.sedar.com/

Gorustovich, S., Monaldi C. and Salfity J., 2011. Geology and metal ore deposits in the Argentina Puna. In Cenozoic Geology of the Central Andes of Argentina, 169-187, SCS Publisher.

Jacobs Engineering Group, 1999. Feasibility Study Pirquitas Silver-Tin Project, Jujuy Province, Argentina. Prepared for Sunshine Argentina Inc.

Journel, A.G. and C.J. Huijbregts, 1978. Mining Geostatistics, Academic Press London

Kuchling, K., Davis, B., Howie, K., Embree, K. and Fox, J., 2014. Preliminary Economic Assessment for the Chinchillas Silver-Lead-Zinc Project, Jujuy Province, Argentina. 20 January 2014. Retrieved from http://www.sedar.com/

Kuchling, K., Davis, B., Embree, K., Fox, J., Howie, K., and Smith, B., 2015. Preliminary Economic Assessment Update for the Chinchillas Silver-Lead-Zinc Project, Jujuy Province, Argentina. Amended Date February 13th, 2015. Retrieved from http://www.sedar.com/.

Kulemeyer, J. A., 2011. Anexo sobre el patrimonio cultural, arqueológico, paleontológico e histórico de Santo Domingo, departamento de Rinconada, de la actualización bianual del estudio de impacto ambiental - etapa de exploración. Proyecto minero Chinchillas: Golden Arrow environmental impact assessment actualization report for exploration stage annex, 2011, 72 p.

Lorenz, V. and Kurszlaukis S., 2007. Root zone processes in the phreatomagmatic emplacement model and consequences for the evolution of maar-diatreme volcanoes. Journal of Volcanology and Geothermal Research, Vol 159:4-32.

Ma, W., and Redfearn, M., 2014. Mineralogical Assessments on Five Test Products. Unpublished report prepared by Bureau Veritas Commodities Canada Ltd, Inspectorate Metallurgical Division, for Golden Arrow Resources Corporation.

Marshall, D., and Mustard, P., 2012. Chinchillas Intermediate Sulphidation Epithermal system. Unpublished report prepared by Vancouver Petrographics Ltd for Golden Arrow Resources Corporation.

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OreWin, 2022, Puna 2021 Project Update Report

O’Brien, M. F., 2020. Technical Memorandum No. 2 on Geological Model and Grade Estimation for the Chinchillas Silver-Lead-Zinc Mine, San Salvador de Jujuy, Argentina, for Mina Pirquitas SA.

Peacey, J., 2016. Guidance on Treatment terms for Chinchillas Pb-Ag and Zn concentrates. Unpublished report prepared by Kingston Process Metallurgy for Silver Standard Resources Inc.

Puritch, E., 2020. Pirquitas Mine – Underground Mineral Resource Estimate.

Quantec Geoscience Argentina S.A., 2008. Geophysical report on: pole-dipole array, induced polarization and resistivity survey at the Chinchilla Project, Jujuy Province, Argentina, on behalf of Silex Argentina S.A.

Ramos V., 1999. Rasgos estructurales del territorio argentino, Instituto de Geologia y recursos Minerales, Anales 29 (24), Buenos Aires.

Silex Argentina S.A., 2008. Internal report on the Chinchillas Ag-Pb-Zn deposit.

Stanley, C.J., and Armstrong, R.N., 2016. The Opaque Mineralogy of 28 samples from the Pb-Zn-Ag Mineralization of the Chinchillas Project. Unpublished report by the Natural History Museum prepared for Golden Arrow Resources Corporation.

Soler M., Caffe P., Coira B., Onoe A., and Mahlburg Kay S., 2007. Geology of the Vilama caldera: A new interpretation of a large-scale explosive event in the Central Andean plateau during the Upper Miocene, in Journal of Volcanology and Geothermal Research, Vol.164, 27pp.

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25 RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT

The Puna21TRS QPs have relied on the following information provided by SSR in preparing the findings and conclusions in this Technical Report Summary regarding the following aspects of modifying factors:

•Macroeconomic trends, taxes, royalties, data, and assumptions, and interest rates.

•Used in Section 19, as described in that section. The QPs have relied exclusively on SSR for this information.

•Marketing information and plans within the control of the registrant.

•Used in Sections 16 and 19, as described in those sections. The QPs have relied exclusively on SSR for this information.

•Legal matters outside the expertise of the QPs, such as statutory and regulatory interpretations affecting the mine plan. The QPs have relied exclusively on SSR for this information.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

•Environmental matters outside the expertise of the qualified person.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

•Accommodations the registrant commits or plans to provide to local individuals or groups in connection with its mine plans.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

•Governmental factors outside the expertise of the qualified person.

•Content in Sections 3 and 17 are based exclusively on information and data supplied by SSR.

Following a review of the information supplied, the opinion of the QPs is, that it is reasonable to rely on the information provided by SSR as outlined above for use in the Puna21TRS because a significant environmental and social analysis has been conducted for the project over an extended period, the project has been in operation for a number of years, SSR employs professionals and other personnel with responsibility in these areas and these personnel have the best understanding of these areas.

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