HBM 6-K, filed Mar 30, 2021, 0001062993-21-003156

UNITED STATES SECURITIES AND EXCHANGE COMMISSION Washington, D.C. 20549

FORM 6-K

REPORT OF FOREIGN PRIVATE ISSUER PURSUANT TO RULE 13A-16 OR 15D-16 OF THE SECURITIES EXCHANGE ACT OF 1934

For the month of March 2021

Commission File Number: 001-34244

HUDBAY MINERALS INC. (Translation of registrant’s name into English)

25 York Street, Suite 800 Toronto, Ontario M5J 2V5, Canada (Address of principal executive offices)

Indicate by check mark whether the registrant files or will file annual reports under cover of Form 20-F or Form 40-F.

Form 20-F [ ] Form 40-F [X]

Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(1): [ ]

Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(7): [ ]

Indicate by check mark whether the registrant by furnishing the information contained in this Form is also thereby furnishing the information to the Commission pursuant to Rule 12g3-2(b) under the Securities Exchange Act of 1934.

Yes [ ] No [X]

If “Yes” is marked, indicate below the file number assigned to the registrant in connection with Rule 12g3-2(b): 82- _____________________________

EXPLANATORY NOTE

On March 29, 2021, Hudbay Minerals Inc. (“Hudbay”) filed on the Canadian Securities Administrators’ System for Electronic Document Analysis and Retrieval (SEDAR) website at www.sedar.com the following documents: (1) NI 43-101 Technical Report - Lalor and Snow Lake Manitoba, Canada, (2) Certificate of Qualification - Olivier Tavchandjian - Snow Lake Operations, (3) Consent of Qualified Person - Olivier Tavchandjian - Snow Lake Operations, (4) NI 43-101 Technical Report - Constancia Mine, Cuzco, Peru, (5) Certificate of Qualification - Olivier Tavchandjian - Constancia Mine, Cuzco and (6) Consent of Qualified Person - Olivier Tavchandjian - Constancia Mine, Cuzco.

Copies of the filings are attached to this Form 6-K and incorporated herein by reference, as follows:

Exhibit 99.1 — NI 43-101 Technical Report - Lalor & Snow Lake Manitoba, Canada
Exhibit 99.2 — Certificate of Qualification - Olivier Tavchandjian - Snow Lake Operations
Exhibit 99.3 — Consent of Qualified Person - Olivier Tavchandjian - Snow Lake Operations
Exhibit 99.4 — NI 43-101 Technical Report - Constancia Mine, Cuzco, Peru
Exhibit 99.5 — Certificate of Qualification - Olivier Tavchandjian - Constancia Mine, Cuzco
Exhibit 99.6 — Consent of Qualified Person - Olivier Tavchandjian - Constancia Mine, Cuzco

SIGNATURE

Pursuant to the requirements of the Securities Exchange Act of 1934, the registrant has duly caused this report to be signed on its behalf by the undersigned, thereunto duly authorized.

HUDBAY MINERALS INC.
(registrant)
By:	/s/ Patrick Donnelly
Name:	Patrick Donnelly
Title:	Vice President and General Counsel

Date: March 30, 2021

EXHIBIT INDEX

The following exhibits are furnished as part of this Form 6-K:

Exhibit	Description
99.1	NI 43-101 Technical Report - Lalor & Snow Lake Manitoba, Canada
99.2	Certificate of Qualification - Olivier Tavchandjian - Snow Lake Operations
99.3	Consent of Qualified Person - Olivier Tavchandjian - Snow Lake Operations
99.4	NI 43-101 Technical Report - Constancia Mine, Cuzco, Peru
99.5	Certificate of Qualification - Olivier Tavchandjian - Constancia Mine, Cuzco
99.6	Consent of Qualified Person - Olivier Tavchandjian - Constancia Mine, Cuzco

 Hudbay Minerals Inc.: Exhibit 99.1 - Filed by newsfilecorp.com

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

CAUTIONARY NOTE REGARDING FORWARD LOOKING INFORMATION

This Technical Report contains "forward-looking statements" and "forward-looking information" (collectively, "forward-looking information") within the meaning of applicable Canadian and United States securities legislation. All information contained in this Technical Report, other than statements of current and historical fact, is forward-looking information. Often, but not always, forward-looking information can be identified by the use of words such as "plans", "expects", "budget", "guidance", "scheduled", "estimates", "forecasts", "strategy", "target", "intends", "objective", "goal", "understands", "anticipates" and "believes" (and variations of these or similar words) and statements that certain actions, events or results "may", "could", "would", "should", "might" "occur" or "be achieved" or "will be taken" (and variations of these or similar expressions). All of the forward-looking information in this Technical Report is qualified by this cautionary note.

Forward-looking information includes, but is not limited to, our objectives, strategies, intentions and expectations regarding Hudbay's Lalor mine and Snow Lake operations, including the anticipated commissioning and ramp-up of the New Britannia mill, forecasted production and projected capital and operating costs the based on the life of mine plans for Lalor and the Snow Lake operations, the forecasted cash costs of production, planned and potential optimization opportunities, including the Stall mill recovery improvement project and the potential to increase the future production rate at Lalor, the potential to convert mineral resources to mineral reserves and to identify additional mineral resources through exploration, the anticipated effect of external factors on revenue, such as commodity prices, estimation of mineral reserves and resources, mine life projections, reclamation costs, economic outlook, government regulation of mining operations, and expectations regarding labour and community relations. Forward-looking information is not, and cannot be, a guarantee of future results or events. Forward-looking information is based on, among other things, opinions, assumptions, estimates and analyses that, while considered reasonable by us at the date the forward-looking information is provided, inherently are subject to significant risks, uncertainties, contingencies and other factors that may cause actual results and events to be materially different from those expressed or implied by the forward-looking information.

The material factors or assumptions that we identified and were applied by us in drawing conclusions or making forecasts or projections set out in the forward-looking information include, but are not limited to:

the schedule for the refurbishment, commissioning and ramp-up of the New Britannia mill;
the success of mining, processing, exploration and development activities;
the scheduled maintenance and availability of our Snow Lake processing facilities;
the accuracy of geological, mining and metallurgical estimates;
anticipated metals prices and the costs of capital projects and production;
the supply and demand for metals we produce;
the supply and availability of third party processing facilities for our concentrate;
the supply and availability of all forms of energy and fuels at reasonable prices;
the availability of transportation services at reasonable prices;
no significant unanticipated operational or technical difficulties;
the availability of personnel for our exploration, development and operational projects and ongoing employee relations;
maintaining good relations with the communities surrounding Lalor mine and no labour disruptions;
no significant unanticipated challenges with stakeholders at our various projects;
no significant unanticipated events or changes relating to regulatory, environmental, health and safety matters;
no contests over title to our properties, including as a result of rights or claimed rights of aboriginal peoples;

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

no significant unanticipated litigation; and
no significant and continuing adverse changes in general economic conditions or conditions in the financial markets (including commodity prices and foreign exchange rates).

The risks, uncertainties, contingencies and other factors that may cause actual results to differ materially from those expressed or implied by the forward-looking information may include, but are not limited to, risks generally associated with the mining industry, such as economic factors (including future commodity prices, currency fluctuations, energy prices and general cost escalation), risks related to the new Lalor and Snow Lake mine plan including the schedule for the refurbishment, commissioning and ramp-up of the New Britannia mill and the Stall mill recovery improvement project, risks related to the transition of our work force from the 777 mine to Snow Lake following the closure of the 777 mine, risks related to the negotiation of new collective bargaining agreement with the labour unions representing our employees in Manitoba, uncertainties related to the development and operation of our projects, dependence on key personnel and employee and union relations, risks related to political or social unrest or change, risks in respect of aboriginal and community relations, rights and title claims, operational risks and hazards, including unanticipated environmental, industrial and geological events and developments and the inability to insure against all risks, failure of plant, equipment, processes, transportation and other infrastructure to operate as anticipated, depletion of Hudbay's reserves, volatile financial markets that may affect our ability to obtain additional financing on acceptable terms, the failure to obtain required approvals or clearances from government authorities on a timely basis, uncertainties related to the geology, continuity, grade and estimates of mineral reserves and resources, and the potential for variations in grade and recovery rates, uncertain costs of reclamation activities, Hudbay's ability to comply with its pension and other post-retirement obligations, our ability to abide by the covenants in our debt instruments and other material contracts, tax refunds, hedging transactions, as well as the risks discussed under the heading "Risk Factors" in our most recent Annual Information Form dated March 29, 2021

Should one or more risk, uncertainty, contingency or other factor materialize, or should any factor or assumption prove incorrect, actual results could vary materially from those expressed or implied in the forward-looking information. Accordingly, you should not place undue reliance on forward-looking information. We do not assume any obligation to update or revise any forward-looking information after the date of this Technical Report or to explain any material difference between subsequent actual events and any forward-looking information, except as required by applicable law.

Hudbay uses certain non-IFRS financial performance measures in its financial reports and in this 43-101 Technical Report. Cash cost and sustaining cash cost per pound of copper and ounce of gold produced are shown because the company believes they help investors and management assess the performance of its operations, including the margin generated by the operations and the company. Unit operating costs are shown because these measures are used by the company as a key performance indicator to assess the performance of its mining and processing operations. These measures do not have a meaning prescribed by IFRS and are therefore unlikely to be comparable to similar measures presented by other issuers. These measures should not be considered in isolation or as a substitute for measures prepared in accordance with IFRS and are not necessarily indicative of operating profit or cash flow from operations as determined under IFRS. Other companies may calculate these measures differently. For further details on these measures, including reconciliations of historical unit operating costs and cash costs per pound of copper produced to the most comparable IFRS measures, please refer to page 53 of Hudbay's management's discussion and analysis for the three and twelve months ended December 31, 2020 available on SEDAR at www.sedar.com.

This 43-101 Technical Report contains references to both United States dollars and Canadian dollars. All dollar amounts referenced, unless otherwise indicated, are expressed in constant Canadian dollars.

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

SIGNATURE PAGE

This Technical Report titled "NI 43-101 Technical Report, Lalor and Snow Lake Operations, Manitoba, Canada", dated March 29^th^ 2021 and effective as of January 1^st^ 2021, was prepared under the supervision and signed by the following author:

Dated this 29^th^ day of March, 2021.

/s/ Olivier Tavchandjian Signature of Qualified Person

Olivier Tavchandjian, P. Geo.

Vice President, Exploration and Geology

Hudbay Minerals Inc

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

TABLE OF CONTENTS

1. Summary	1-1
1.1. Qualified Person (QP)	1-1
1.2. Property Location, Ownership and Description	1-1
1.3. History	1-2
1.4. Geology and Mineralization	1-2
1.5. Exploration	1-3
1.6. Drilling, Sampling, Assaying, QAQC and Validation	1-3
1.7. Mineral Processing and Metallurgical Testing	1-4
1.8. Mineral Resource Estimates	1-5
Lalor, 1901, Watts and Pen II	1-5
WIM and 3 Zone	1-6
New Britannia Mine	1-7
1.9. Mineral Reserve Estimates	1-8
1.10. Mining Methods and Production Schedule	1-9
1.11. Recovery Methods	1-10
1.12. Market Studies and Contracts	1-11
1.13. Environmental Studies, Permitting and Social or Community Impact	1-11
1.14. Capital And Operating Costs	1-12
1.15. Interpretation and Conclusion	1-13
1.16. Recommendations	1-14
2. Introduction & Terms of Reference	2-15
2.1. Qualified Person (QP)	2-16
2.2. Sources of Information	2-16
2.3. Unit Abbreviations	2-16
2.4. Acronyms and Abbreviations	2-17
2.5. Other Contributions	2-18
3. Reliance on Other Experts	3-19
4. Property Location and Description	4-20
4.1. Lalor Mine, 1901 and Pen II Deposits	4-20
Land Use Permitting	4-21
4.2. New Britannia Mine and Other Snow Lake Gold Properties	4-22
4.3. Watts Deposit	4-24
4.4. WIM Deposit	4-25
5. Accessibility, Climate, Local Resources, Infrastructure and Physiography	5-27
5.1. Accessibility	5-27
5.2. Climate	5-27
5.3. Local Resources	5-29
5.4. Existing Infrastructure	5-29

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
5.5. Physiography	5-31
---	---
6. History	6-32
6.1. Lalor & The Chisel Basin Area	6-32
6.2. New Britannia and Snow Lake Deposits	6-32
6.3. WIM Deposit	6-33
6.4. Pen II	6-33
6.5. Watts	6-33
7. Geological Setting & Mineralization	7-34
7.1. Regional Geology	7-34
VMS Deposits	7-36
Snow Lake Gold Deposits	7-37
7.2. Property Geology	7-38
Lalor	7-38
Snow Lake Gold Properties	7-39
1901 Deposit	7-40
Pen II Deposit	7-40
Watts Deposit	7-41
Wim Deposit	7-42
8. Deposit Types	8-44
8.1. VMS Deposits	8-44
Lalor Mine and 1901 Deposit	8-45
Watts Deposit	8-45
Pen II Deposit	8-46
Wim Deposit	8-46
8.2. Snow Lake Gold Properties	8-46
9. Exploration	9-49
9.1. Lalor Mine	9-49
9.2. 1901 Deposit	9-49
9.3. Snow Lake Gold Properties	9-49
9.4. Pen II Deposit	9-50
9.5. WIM Deposit	9-50
9.6. Watts	9-50
9.7. Region Exploration	9-51
10. Drilling	10-52
10.1. Lalor Mine	10-52
Surface Drilling	10-52
Underground Drilling	10-52
10.2. 1901 Deposit	10-54
10.3. Snow Lake Gold Properties	10-54
10.4. Watts	10-55

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
10.5. Pen II	10-56
---	---
10.6. WIM	10-56
11. Sample Preparation, Analyses & Security	11-56
11.1. Lalor, 1901 & Pen II	11-56
Sample Preparation	11-56
ACME/Bureau Veritas	11-57
SGS	11-57
Specific Gravity via gas pycnometer (Bureau Veritas, Vancouver)	11-58
Quality Control and Quality Assurance	11-58
Standards	11-58
Blanks	11-58
Coarse Duplicates	11-59
Umpire Laboratory	11-59
11.2. Watts Deposit	11-59
Sample Preparation	11-59
Specific Gravity	11-60
Sample Analysis	11-60
Hudbay's Flin Flon Laboratory	11-60
Bureau Veritas	11-60
SGS	11-60
Quality Assurance and Quality Control	11-61
Standards	11-61
Blanks	11-61
Coarse Duplicates	11-61
Umpire Laboratory	11-61
11.3. WIM	11-62
Sample Preparation	11-62
Specific Gravity	11-62
Sample Analysis	11-62
Quality Assurance and Quality Control	11-62
11.4. Snow Lake Gold Properties	11-62
Sample Preparation	11-62
Sample Analysis	11-63
Bureau Veritas	11-63
SGS	11-64
Specific Gravity	11-64
Quality Assurance and Quality Control	11-64
Standards	11-64
Blanks	11-65
Coarse Duplicates	11-65

iii

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
Umpire Laboratory	11-65
---	---
12. Data Verification	12-66
12.1. Lalor Mine, 1901 Watts, 3 Zone, WIM and Pen II Deposits	12-66
Site Visits	12-66
Database Management	12-66
Database Review	12-66
12.2. Snow Lake Gold Properties	12-66
Site Visits	12-66
Database Review	12-67
13. Mineral Processing and Metallurgical Testing	13-68
13.1. Stall Concentrator	13-68
Historical Metallurgical Performance	13-68
Historical Metallurgical Testwork (2009 to 2018)	13-69
Metallurgical Testwork (2019 to 2020) - Lalor Base Metal Ore	13-69
Mineralogical Investigation	13-71
Jameson Cell Pilot Plant Testwork	13-71
Future Ore Testwork	13-72
13.2. New Britannia Concentrator	13-73
Metallurgical Testwork - SGS Lakefield (2015)	13-74
Sample Selection	13-74
Mineralogical Investigation	13-75
Communication Testwork	13-75
Flotation Testwork	13-76
Leach Testwork	13-77
Cyanide Destruction Testwork	13-78
Metallurgical Testwork - Blue Coast Research (2019 to 2020)	13-79
Sample Selection	13-79
Mineralogical Investigation	13-79
Flotation Testwork	13-80
13.3. Flotation Tails Leach Testwork	13-82
13.4. 1901 Base Metal Ore Metallurgical Testwork	13-83
Sample Selection	13-83
Mineralogical Investigation	13-84
Flotation Testwork	13-85
13.5. WIM Metallurgical Testwork	13-86
Sample Selection	13-86
Mineralogy	13-87
Comminution Testwork	13-88
Flotation Testwork	13-88
13.6. 3 Zone Metallurgical Testwork	13-89

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
Sample Selection	13-89
---	---
Mineralogy	13-90
Leach Testwork	13-90
Gravity Concentration Testwork	13-91
13.7. Pen II Metallurgical Testwork	13-92
14. Mineral Resource Estimates	14-93
14.1. Lalor Mine	14-93
Modeling Database	14-93
Modeling of the Mineralized Envelopes	14-93
Density	14-94
Compositing	14-94
Exploratory Data Analysis	14-94
Grade Capping	14-95
Variography	14-96
Grade Estimation and Interpolation Methods	14-96
Grade Estimation Validation	14-97
Global Bias Checks	14-98
Smoothing Assessment	14-98
Smoothing Correction	14-98
Classification of Mineral Resources	14-99
Reasonable Prospects of Economies Extraction and Mineral Resource Estimates	14-100
14.2. 1901 Deposit	14-101
Modeling Database	14-101
Modeling of the Mineralized Envelopes	14-101
Density	14-102
Compositing	14-102
Exploratory Data Analysis	14-102
Grade Capping	14-103
Variography	14-103
Grade Estimation and Interpolation Methods	14-103
Grade Estimation Validation	14-103
Visual Inspection	14-104
Global Bias Checks	14-104
Smoothing Correction	14-104
Classification of Mineral Resource	14-105
14.3. Watts River	14-106
Modeling Database	14-106
Modeling of the Mineralized Envelope	14-106
Density	14-107
Compositing	14-107

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
Exploratory Data Analysis	14-107
---	---
Grade Capping	14-108
Variography	14-108
Grade Estimation and Interpolation Methods	14-108
Grade Estimation Validation	14-108
Visual Inspection	14-109
Global Bias Checks	14-109
Smoothing Assessment	14-109
Smoothing Correction	14-109
Classification of Mineral Resource	14-110
Reasonable Prospects of Economics Extraction and Mineral Resource Estimates	14-110
14.4. Pen Zone	14-111
Modeling Database	14-111
Modeling of the Mineralized Envelope	14-111
Density Assignation	14-111
Compositing	14-111
Variography	14-111
Grade Estimation and Interpolation Methods	14-112
Grade Estimation Validation	14-112
Classification of Mineral Resources	14-112
Reasonable Prospects of Economics Extraction and Mineral Resource Estimates	14-112
14.5. WIM & 3Zone	14-113
14.6. New Britannia Mine and Birch Deposit	14-113
15. Mineral Reserve Estimates	15-116
15.1. Mining Strategy and Reserve Estimations Process	15-117
Mining Method Selection & Mining Strategy	15-117
Mining Method Selection & Mining Strategy for the Future Operation at the 1901, WIM and 3 Zone Deposits	15-118
Reserve Methodology	15-119
15.2. Dilution and Recovery	15-120
Dilution	15-120
Mining Recovery	15-121
15.3. Deswik Stope Optimzer	15-121
15.4. Stope Net Smelter Return Calculation	15-122
15.5. Ore Development	15-123
15.6. Reconciliation of Reserves To Production	15-123
16. Mining methods	16-125
16.1. Lateral Development	16-125
16.2. Vertical Development	16-125
16.3. Stope Mining	16-126

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
Cut and Fill Mining	16-126
---	---
Longhole Open Stope Mining	16-127
Teleremote and Autonomous Mucking	16-129
16.4. Backfill	16-129
16.5. Ore Handling	16-129
16.6. Geotechnical Design	16-129
16.7. Ground Support Systems	16-130
Primary Support	16-130
16.8. Underground Development	16-130
Vertical Development	16-131
Raises	16-131
Paste Backfill, Drain and Service Holes	16-131
16.9. Diamond Drilling	16-131
16.10. Mine Drainage and Dewatering	16-131
16.11. Mining Operations	16-131
Mining Fleet	16-131
Mine Power	16-132
Production Schedule	16-132
Mine Ventilation	16-134
16.12. Workforce	16-135
16.13. Mine Safety and Health	16-136
Refuge Stations	16-136
Personal Evacuation Infrastructures	16-136
17. Recovery Methods	17-137
17.1. Introduction	17-137
17.2. Stall Mill	17-137
Flowsheet	17-137
Crushing	17-139
Grinding	17-139
Flotation	17-140
Talc Flotation	17-140
Copper Flotation	17-140
Zinc Flotation	17-140
Concentrate Dewatering	17-141
Tailings	17-141
Water	17-141
Stall Mill Recovery Improvement Project	17-141
Future Stall Projects	17-143
17.3. New Britannia Mill	17-143
Flowsheet	17-143

vii

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
Ore Receiving	17-145
---	---
Grinding	17-145
Flotation	17-145
Concentrate Dewatering	17-146
Leaching	17-146
Tailings	17-146
Water	17-146
Future New Britannia Projects	17-147
18. Project Infrastructure	18-148
18.1. Area Surface Infrastructure	18-148
Chisel North Mine Site	18-148
Chisel Lake Mine Site	18-149
Anderson Tailings Impoundment Area Site	18-149
Mine Access Corridor	18-149
WIM and 3 Zone Project Sites	18-149
18.2. Mine Site Surface Infrastructure	18-150
18.3. Mine Underground Infrastructure	18-153
Electrical Distribution and Equipment	18-153
Mine Underground Power Supply	18-154
Emergency Generators	18-154
1901 Underground Infrastructure	18-155
WIM and 3 Zone Underground Infrastructure	18-155
18.4. Paste Plant and Anderson Pump Station	18-156
18.5. Stall Concentrator and Anderson TIA	18-157
18.6. New Britannia Mill	18-158
18.7. Snow Lake Camp	18-159
19. Market Studies and Contracts	19-161
19.1. Metal Marketing	19-161
Zinc Concentrates and Zinc Metal	19-161
Copper Concentrates	19-161
Gold Doré	19-161
19.2. Contracts	19-161
20. Environmental Studies, Permitting and Social or Community IMpact	20-163
20.1. Environmental Studies and Planning	20-164
Lalor Mine Stall Mill and Paste Plant	20-164
New Britannia Mill	20-164
Snow Lake Satellite deposits	20-164
20.2. Waste and Tailings Disposal and Water Management	20-164
Lalor Mine, Stall Mill and Paste Plant	20-164
New Britannia Mill	20-165

viii

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
Satellite deposits	20-165
---	---
20.3. Permitting Requirements	20-165
Lalor Mine, Stall Mill and Paste Plant	20-165
New Britannia Mine Mill	20-165
Satellite deposits	20-166
20.4. Community Support	20-166
20.5. First Nations	20-166
20.6. Heritage Resources	20-166
20.7. Mine Closure	20-166
Satellite Deposits	20-167
Timing of Closure	20-167
21. Capital and Operating Costs	21-168
21.1. Capital Costs	21-168
New Britannia Mill Refurbishment and Pipeline Corridor	21-168
Sustaining Capital and Capitalized Development	21-169
21.2. Operating Costs	21-170
Mining Operating Costs	21-170
Milling Operating Costs	21-171
Refining Costs	21-172
General and Administration Costs	21-172
21.3. Unit Cash And Sustaining Cash Costs	21-173
22. Economic Analysis	22-175
23. Adjacent Properties	23-176
24. Other Relevant Data and Information	24-177
25. Interpretations and Conclusions	25-178
26. Recommendations	26-180
27. References	27-182

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

LIST OF FIGURES

Figure 4-1: Location of the Lalor, 1901 and Pen II Deposits Mineral Leases and Mineral Claims	4-21
Figure 4-2: Location of the Snow Lake and Squall Lake Mineral Leases and Mineral Claims	4-24
Figure 4-3: Location of the Watts Deposit Mineral Claims	4-25
Figure 4-4: Location of the WIM Deposit Mineral Claims	4-26
Figure 5-1: Snow Lake Regional Map	5-28
Figure 7-1: Geology of Manitoba	7-35
Figure 7-2: Snow Lake Area (After Bailes and Galley, 2011)	7-36
Figure 7-3: Lalor Mine Cross Section with Drill Hole Traces	7-38
Figure 7-4: New Britannia Representative Cross Section with Drill Hole Traces	7-39
Figure 7-5 1901 Representative Cross Section with Drill Hole Traces	7-40
Figure 7-6 Pen II Representative Cross Section with Drill Hole Traces	7-41
Figure 7-7 Watts Representative Cross Section with Drill Hole Traces	7-42
Figure 7-8: Wim Deposit Cross Section with Drill Hole Traces	7-43
Figure 8-1: Conceptual Model of a Volcanogenic Massive Sulfide (VMS) Deposit	8-44
Figure 8-2: Schematic Diagram Showing the Setting and Nature of Orogenic Lode-Gold Deposits	8-47
Figure 9-1: Mineral Occurrences in the Snow Lake Camp	9-51
Figure 13-1: Stall Concentrator Flotation Flowsheet April 2019	13-70
Figure 13-2: Jameson Cell Pilot Testwork Results for the Copper Circuit	13-72
Figure 13-3: Jameson Pilot Testwork Results for the Zinc Circuit	13-72
Figure 13-4: Head Grade Recovery Relationship For Copper, Gold and Zinc in Future Lalor Base Metal Ores	13-73
Figure 13-5: New Britannia Process Flowsheet	13-74
Figure 13-6: Composite Sample Mineralogy	13-80
Figure 13-7: Talc Prefoliation Rougher Testwork Results	13-82
Figure 13-8: Metallurgical Testwork Composite Sample Drill Core Location	13-84
Figure 13-9: 1901 Composite Sample Mineralogy	13-85
Figure 13-10: WIM Composite Sample Drill Hole Locations	13-87
Figure 13-11: WIM Composite Sample Minerology	13-88
Figure 13-12: WIM Composite Sample Cleaner Flotation Testwork Grade Recovery Curves	13-89
Figure 13-13: 3 Zone Master Composite Samples Sized Minerology	13-90
Figure 13-14: 3 Zone Arsenic Head Grade Correlation with Leach Residue Gold Grade	13-91
Figure 14-1: 3D View of Mineral Envelope Wireframes, Looking West	14-94
Figure 14-2: SUB-Domain in Mineralized Envelope 25	14-95
Figure 14-3: Visual Check on Model for ZN Grade in Lens 20	14-97
Figure 14-4: Resource Classification in Mineralized Envelope 25	14-99
Figure 14-5: View of Mineral Envelope Wireframes	14-102
Figure 14-6: Plan view of the Resource Classification	14-105
Figure 14-7: View of Mineral Envelope Wireframe	14-107
Figure 15-1: 840 Level 10 Lens Plan View at Lalor	15-117

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
Figure 15-2: 910 Level Post Pillar C&F at Lalor	15-118
---	---
Figure 15-3: Access to 1901 Deposit Isometric View	15-119
Figure 16-1: Typical Post Pillar Cut and Fill Mining Plan View	16-126
Figure 16-2: Typical Inclined ROM and Pillar Mining Plan View	16-127
Figure 16-3: Typical Isometric View - Transverse Longhole Open Stopping	16-128
Figure 16-4: Typical Long Section - Longitudinal Retreat LH Open Stoping	16-128
Figure 16-5: Current Lalor Mine Ventilation Long Section	16-135
Figure 17-1: Stall Mill Flowsheet	17-139
Figure 17-2: new britannia Mill Flowsheet	17-145
Figure 18-1: Map of Snow Lake Mining, Processing and Tailings Facilities	18-148
Figure 18-2: Lalor Mine Site General Arrangement	18-151
Figure 18-3: WIM Site Layout	18-152
Figure 18-4: 3 Zone Site Layout	18-152
Figure 18-5: New Britannia Site Layout	18-158

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

LIST OF TABLES

Table 1-1: Summary of Mineral Resource Estimates for the Lalor, 1901, Watts and Pen II deposits (as of January 1st, 2021)	1-6
Table 1-2: Gold Resource Estimate for the New Britannia Mine and Birch Deposit ^(1)(2)(3)^	1-7
Table 1-3: Summary of Lalor, 1901, WIM and 3 Zone Mineral Reserves	1-9
Table 1-4: LOM Production Schedule	1-10
Table 1-5: Cash and Sustaining Unit Cash Costs	1-13
Table 2-1: Unit Abbreviations	2-17
Table 2-2: Acronyms and Abbreviation	2-17
Table 2-3: Contributions for this Report	2-18
Table 4-1: List of Mineral Leases, OIC and Mineral Claims for the Lalor Mine, 1901 Deposit and Pen II Deposit	4-20
Table 4-2 General Permits and Quarry Lease for Lalor	4-22
Table 4-3: List of Mineral Leases and Mineral Claims for the Snow Lake Gold and Squall Lake Properties	4-23
Table 4-4: List of Mineral Claims for the Watts Deposit	4-24
Table 4-5: List of Mineral Claims for the WIM Deposit	4-26
Table 6-1: Lalor Mine Production	6-32
Table 10-1: Surface Drilling Programs by Year (In Meters Drilled)	10-52
Table 10-2: Underground drilling programs by year (In Meters Drilled)	10-53
Table 10-3: Drilling Data by Year	10-53
Table 10-4: Summary of the 1901 Deposit Surface Diamond Drilling	10-54
Table 10-5: Summary of Surface and Underground Diamond Drill Holes (As of January 1, 2021)	10-55
Table 11-1: Watts Samples Preparation And Analysis by Year	11-60
Table 11-2: Density Measurements At Watts by Year	11-60
Table 13-1: Historical Stall Concentrator Metallurgical Performance	13-68
Table 13-2: Historical Stall Concentrator Concentrate Production	13-69
Table 13-3: Flotation Circuit Survey Mass Balance	13-70
Table 13-4: LOM Forecast Metal Recovery Range	13-74
Table 13-5: SGS Lakefield Composite Sample Head Grades	13-75
Table 13-6: Overall Composite Mineralogy	13-75
Table 13-7: Communication Testwork Results	13-76
Table 13-8: Rougher and Cleaner Flotation Teamwork Results	13-77
Table 13-9: Flotation Tails Cyanidation Leach Testwork Results	13-78
Table 13-10: Blue Coast Research Composite Sample Head Grades	13-79
Table 13-11: Rougher Flotation Testwork Results	13-81
Table 13-12: Lock Cycle Test Results	13-82
Table 13-13: Flotation Tails Leach Testwork Gold Extraction Results	13-83
Table 13-14: 1901 Base Metal Lense Composite Assays	13-84
Table 13-15: 1901 Floatation Testwork Copper Circuit Results	13-86
Table 13-16: WIM Composite Sample Head Grades	13-87

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report
Table 13-17: WIM Master Composite Rougher Floatation Testwork Results	13-88
---	---
Table 13-18: 3 Zone Composite Sample Head Grades	13-90
Table 13-19: 3 Zone Master Composite Leach Testwork Results	13-91
Table 13-20: 3 Zone Knelson Concentrator Testwork Results	13-92
Table 14-1: 2021 Capping Thresholds by Lens of Gold and Silver (G/T)	14-96
Table 14-2: Search Ellipse Parameters	14-97
Table 14-3: Summary of the Applied Smoothing Correction	14-98
Table 14-4: Lalor Mineral Resource Estimates	14-100
Table 14-5: Search Ellipse Parameters	14-103
Table 14-6: Summary of Smoothing Corrections - 1901 Deposit	14-104
Table 14-7: 1901 Mineral Resource Estimates	14-106
Table 14-8: Search Ellipse Parameters	14-108
Table 14-9: Summary of Smoothing Corrections - Watts River	14-110
Table 14-10: Watts Mineral Resource Estimates	14-110
Table 14-11: Pen II Mineral Resource Estimates	14-112
Table 14-12: Cost and Recovery Price Assumption	14-114
Table 14-13: Mineral Resource Estimate for the New Britannia Mine and Birch Deposit ^(1)(2)(3)^	14-114
Table 15-1: Lalor Mine and Snow Lake Deposits Mineral Reserve Estimates (as of January 01, 2021)	15-116
Table 15-2: Stoping Parameters by Mining Method	15-120
Table 15-3: Dilution Incurred by Mining Method	15-121
Table 15-4: Recovery Factors by Mining Method	15-121
Table 15-5: Average Recovery Factors by Mining Method	15-121
Table 15-6: Cut-Off Costs	15-122
Table 15-7: Reserve Model Depletion Reconciled to Mine's Credited Production by the Mill	15-124
Table 16-1: Mine Mobile Equipment	16-132
Table 16-2: Mine Production Schedule	16-133
Table 16-3: Mill Production Schedule	16-133
Table 16-4: Mill Recovery Schedule	16-133
Table 16-5: Metal Production Schedule	16-134
Table 16-6: Mine Operations Workforce	16-136
Table 17-1: Forecast Metal Recovery for the Current Flowsheet and an Upgraded Flowsheet	17-142
Table 17-2: Stall Recovery Project Class 3 CAPEX	17-142
Table 18-1: New Britannia Mill Refurbishment - Project Development Schedule	18-159
Table 21-1: New Britannia Mill Refurbishment Capital Costs	21-168
Table 21-2: Snow Lake Capital Expenditures	21-169
Table 21-3: Mine Operating Costs	21-171
Table 21-4: Milling Operating Costs	21-172
Table 21-5: Refining Operating Costs	21-172
Table 21-6: General and Administration Costs	21-173
Table 21-7: Cash and Sustaining Unit Cash Costs	21-174

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

1. SUMMARY

Hudbay is a Canadian integrated mining company with assets in North and South America. Hudbay operates multiple properties in the Province of Manitoba. Current operations near Snow Lake include the Lalor underground mine, the Stall ore concentrator, a tailings impoundment area and other ancillary facilities that support the operation including a paste plant at Lalor mine. The New Britannia mill which will produce a copper concentrate as well as gold-silver doré is currently being refurbished with a planned start of production in the second half of 2021.

This Technical Report has been prepared to support the public disclosure of Mineral Resources and Mineral Reserves at the Lalor Mine and nearby Snow Lake satellite deposits and to provide an updated mine plan that achieves up to 5,300 tpd of base metal, gold and copper-gold ore feed to the Stall and New Britannia concentrators. Refurbishing the New Britannia mill will significantly increase gold production from Lalor, extend the life of the Snow Lake operations through the processing of gold production from the WIM and 3 Zone deposits and enable new gold and copper-gold exploration opportunities in the Snow Lake region by having an operating processing facility with substantially higher gold, silver and copper recoveries. In addition, the recently discovered 1901 deposit, located between the former producing Chisel mine and Lalor mine and within trucking distance to surface will provide supplementary feed for several years in order to maximize the use of the processing capacity at both the Stall and New Britannia mills.

The Snow Lake satellite deposits that host additional mineral reserve estimates and are included in the consolidated mine plan of the Snow lake operations include the 1901, WIM and 3 Zone deposits. Additionally, mineral resources have also been estimated for the Pen II, Watts and New Britannia properties. These satellite deposits are not currently part of the mine plan and may provide an opportunity in addition to the Lalor and 1901 inferred mineral resource estimates to increase Hudbay's mineral reserves in the area and further expand production and/or extend the operating lives of the Stall and New Britannia processing facilities.

1.1. QUALIFIED PERSON (QP)

This Technical Report has been prepared in accordance with National Instrument Form 43-101F1. The qualified person ("QP") who supervised the preparation of this Technical Report is Olivier Tavchandjian, P. Geo., Hudbay's Vice-President, Exploration and Geology.

1.2. PROPERTY LOCATION, OWNERSHIP AND DESCRIPTION

For all the properties described in this Technical Report, Hudbay owns a 100% interest and there are no royalties payable other than those potentially payable to the province. Surface rights are held under general permits and are sufficient for purposes of our development plans. These properties and their associated infrastructures are located within 60 km of the town of Snow Lake, approximately 200 km by road east of the city of Flin Flon where the nearest full service commercial airport is located.

All these properties are accessible by road from the town of Snow Lake. The Snow Lake area has a typical mid-continental climate, with short summers and long, cold winters. Climate generally has only a minor effect on local exploration and mining activities. The area is surrounded by water bodies. Topography shows gentle relief that rarely exceeds 10 m.

The town of Snow Lake is a full-service community with available housing, hospital, police, fire department, potable water system, restaurants and stores. To house non-local employees during their work rotations, the company provides a camp located in town which services Hudbay employees and contractors for the mine and mill operations. Other infrastructure in the area includes provincial roads, a 115 kV Manitoba Hydro power grid within four kilometres of Lalor and Manitoba Telecom land line and cellular phone service.

Hudbay operates two concentrators in the Snow Lake / Flin Flon area: the Stall mill located approximately 16 km from Lalor and the Flin Flon mill located 200 km to the west. While the Flin Flon mill provides an opportunity to truck additional production from Lalor that the Stall mill cannot accommodate, the recent operating performance of the Stall mill through 2020 has resulted in minimum tonnage being trucked to Flin Flon from Snow Lake over the past year as the Stall mill has been operating close to the 3,900tpd produced by Lalor mine

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

Hudbay also operates a zinc metallurgical plant in Flin Flon with a capacity of 115,000 tonnes per annum of refined zinc. However, Hudbay intends to cease operations and put the zinc plant and the Flin Flon mill on care and maintenance at the end of Q2 2022, when Hudbay's 777 mine closes. After the zinc plant closure, the concentrates produced from the Stall and Flin Flon mills will be sold to market.

The New Britannia gold mill is located approximately 16 km east of the Lalor mine and it is currently being refurbished with a new copper flotation plant under construction Hudbay plans to complete the refurbishment, commissioning and ramp-up of the New Britannia mill by the third quarter of 2021, adding a new copper flotation circuit with a nominal throughput capacity of 1,500tpd.

1.3. HISTORY

The Snow Lake area has a long exploration and mining history. Exploration in the Lalor-Chisel area has been occurring since the 1950s and the Chisel Basin area has hosted four past producing mines. This basin is also the host of the Lalor deposit. Lalor commenced initial ore production from the ventilation shaft in August 2012, only five years after its initial discovery hole and achieved commercial production from the main shaft in the third quarter of 2014. Since 2012, the mine has produced 8.8 million tonnes.

Gold was first discovered in 1914 approximately 20 km to the southeast of Snow Lake and in 1917, the Moose Horn-Ballast claims produced the first gold in Manitoba. First mine construction at the New Britannia site started in 1945 and in March 1949, the mine was opened as the Nor-Acme mine. Production continued until 1958. 4.9 Mt were mined at an average gold grade of 4.4 g/t and Nor-Acme mill recovered approximately 610,000 Oz of gold during this production period. TVX and High River formed a joint venture to reopen the mine and TVX became the operator. Full production from the main shaft was achieved in August 1996. Through various transactions, Kinross became the operator of the New Britannia mine-mill complex. Production ceased at the end of September 2004 and the mill was put on care and maintenance in 2005 due to a low gold price environment after producing 1.6 million ounces of gold. The New Britannia complex was subsequently acquired by QMX Gold Corporation.

Hudbay acquired a 100% interest in the New Britannia mill, and related mineral properties, from QMX Gold Corporation in 2015.

1.4. GEOLOGY AND MINERALIZATION

The properties of interest for this Technical Report are all located within the Trans-Hudson Orogen of the Flin Flon Greenstone Belt. The volcanic assemblages consist of mafic to felsic volcanic rocks with intercalated volcanogenic sedimentary rocks.

The volcanogenic massive sulphide (VMS) deposits located near the town of Snow Lake have been subdivided into two different groups: Cu-Zn-rich (Cu-Zn, Cu-Zn-Au) and Zn-Cu-rich (Zn-Pb-Cu-Ag) types. The Cu-Zn-rich deposits mainly occur in the Anderson sequence and the Zn-Cu-rich deposits occur in the Chisel sequence. The Watts River Deposit, located 52km east-southeast of the town of Snow Lake lies in the eastern portion of the Flin Flon-Snow Lake Greenstone belt and is a stratabound accumulation of sulphides that precipitated in a depositional environment similar to the base metal deposits of the Snow Lake mining camp.

Mineralization of the lode-gold vein-type deposits are hosted in the Amisk group mafic and felsics volcanic rocks which are structurally controlled and associated with shear zones, faults, fold hinges and axial planes that host simple to complex vein systems. The mineralization is associated with lithological contacts of contrasting properties in the sequence of interlayered volcanic and volcaniclastic rocks.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

1.5. EXPLORATION

Since the last NI 43-101 Technical Report published in 2019, exploration drilling at Lalor has focused on adding and converting inferred mineral resource estimates with a strong emphasis on confirming the continuity of the gold mineralization. In parallel, surface drilling has been successful in discovering and defining mineral resources for both base metal and gold mineralization at the 1901 deposit and also confirming mineral resources at the Watts deposit.

Exploration drilling will continue in 2021 both at Lalor and at the 1901 deposit in order to confirm and expand the know mineral resource and mineral reserve estimates. Exploration drilling will also follow-up on some initial successes obtained between Lalor and the former producing Chisel mine while efforts continue to discover new base metal and gold deposits in the Snow Lake region.

1.6. DRILLING, SAMPLING, ASSAYING, QAQC AND VALIDATION

At Lalor, over 3,300 drill holes totaling more than 580,000 m were included in the Lalor database to support the mineral resource and mineral reserve estimates while drilling supporting the 1901, Watts, Pen II and Wim mineral resource and mineral reserve estimates totals 51,000m, 25,000m, 2,000 m and 43,000 m respectively. For the New Britannia mineral resource estimates including the 3 Zone and Birch zones, over 730,000 m of drilling completed after 1995 were used. Drilling at all properties is a combination of NQ and BQ diamond drill holes, surveyed with either Reflex downhole tools or Gyro for deeper/longer holes.

At Lalor, 1901, Watts, Pen II and Wim, the core was photographed before samples were split and bagged for shipment before dispatch to the laboratories, QAQC samples were inserted into the sample stream. Hudbay's practice involves insertion of the following every 100 samples, 5 blanks, 5 duplicates, 5 base metal standards and 2 gold standards.

Sample preparation has been conducted at three different laboratories over time. Since September 2016, nearly all samples are prepared and assayed at Bureau Veritas in Vancouver. All drill core samples analyzed since the last technical report (NI 43-101 published on March 28, 2019) have been sent for analysis at Bureau Veritas while the SGS laboratory in Vancouver was used as the umpire laboratory for quality control purposes.

Results from the QA/QC program for standards, blanks, duplicates and external checks show that the program has been working effectively for the Lalor, 1901, Watts, Pen II and Wim properties, meeting industry standards and the data used provides a representative and unbiased basis for resource modeling purposes.

Most of the drill cores and chips assays from the New Britannia mine were tested by the previous owners from 1995 to 2003 and were completed at the on-site mill laboratory using a fire assay/atomic absorption finish (FA/AA) method. Standard, blank and duplicate assay samples were added to each batch of 21 samples for drill core and to each batch of 24 samples for chip samples. The sampling and analytical procedures conformed to the industry standards at the time, and these were adequate to ensure a representative determination for the type of gold mineralization identified on the property. In 2019, 6 holes drilled by Hudbay at 3 Zone confirmed previous drilling results.

It is the QP's opinion that the data is adequate and acceptable for use in the estimation of mineral resources at all the properties covered by this Technical Report.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

1.7. MINERAL PROCESSING AND METALLURGICAL TESTING

The Stall concentrator is an operating plant running at steady state and, as a result, several of the initial metallurgical test results and assumptions have been revised to reflect the operating experience and performance of the plant over the past six years of operation in processing the ore produced from the Lalor mine. The Stall concentrator is producing a copper concentrate grade of 18 to 20% copper at 83 to 85% recovery and a zinc concentrate grade of 51% zinc at 90 to 93% recovery. 55 to 62% of the gold and silver are recovered in the copper concentrate as co-products. Over the life of Lalor mine, copper, gold and silver grade will increase and the average zinc grade will decrease from Lalor. This trend will partially be offset in 2026, when the 1901 deposit enters production and feed Stall with zinc rich mineralization.

Extensive metallurgical testing was conducted in 2019 and 2020 to demonstrate the technical viability and economic benefits of some changes to the process flowsheet of the Stall mill to improve metal recoveries and/or concentrate grade from historical performance. The main changes to the flowsheet are planned to be implemented by the end of 2022 and include the addition of Jameson cells to increase copper rougher and cleaner capacity, the addition of a talc preflotation circuit, an increase in the zinc circuit cleaning capacity and froth washing and an increase in recovery of free gold through the addition of a Knelson gravity concentrator on the copper regrind cyclone underflow. The total cost to implement these upgrades is estimated to be $19 million (C$24 million) and is expected to increase Stall's copper recoveries to between 91% and 95%, gold recoveries to between 64% and 70%, and silver recoveries to between 65% and 74%.

In addition, further testwork is underway to include a lead recovery stage into the sequential flotation circuit. This testwork will consider the benefits of a copper-lead-zinc sequential flowsheet utilizing TNC 312 as the copper circuit collector and 3418A as the lead circuit collector. These further changes to the flowsheet are expected to improve metal recovery and the benefits of this addition have not yet been incorporated into the present life of mine plan, Hudbay anticipates a short payback on this additional limited investment at Stall.

In 2020 a metallurgical testwork program was conducted by Blue Coast Research to cover composites representing low grade, medium grade and high grade of the two zinc rich lenses of the 1901 deposit. A subsample of each of the composite samples was ground to a p80 of 100µm and submitted for mineralogical analysis. Mineralogical analysis, and flotation tests were completed on each of the six composites and confirmed that the metallurgical performance of the Stall concentrator for the Lalor base metal lenses was applicable to the 1901 deposit, including the potential benefit of a lead recovery stage in the flotation circuit.

Should the Watts or Pen II Zones be developed, it is expected that they could be milled at the Stall Concentrator and blended with ore from the Lalor Mine.

To establish the future performance of the New Britannia mill, two major metallurgical test work programs were undertaken between 2015 and 2020 by SGS Lakefield and Blue Coast Research. The first laboratory program was conducted at SGS in 2015-2016 to develop a flowsheet, design criteria, and metallurgical forecast information for the gold and copper-gold rich ore types. All sample composites were submitted for mineralogical analysis by QEMSCAN to identify minerals and their liberation as well as for Bond rod and Bond ball mill work tests, gravity concentration test, rougher and cleaner flotation tests, cyanidation tests, and rolling bottle leach tests. In addition, CIP modelling was completed to predict the gold extraction performance of a CIP or CIL circuit. Cyanide destruction tests using the SO2/Air process were also carried out following standard SGS procedure of completing batch tests first to confirm applicability and to optimize retention times and reagent requirements. These results were confirmed and further refined in 2019-2020 at Blue Coast Research facilities. The testwork program was conducted on 13 composites and included additional mineralogy, flotation, cyanidation, and cyanide destruction tests.

Metallurgical testwork conducted in 2019 on WIM and 3 zone has confirmed that this mineralization is also amenable to successful beneficiation at the New Britannia mill. Four composites were created for each deposit and submitted for mineralogical, comminution and flotation as well as leach test work and gravity concentration in the case of 3 Zone. These tests have been used to confirm the copper, gold and silver recoveries applied in the life of mine plan for these two satellite deposits.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

It is anticipated that any mineral resource from the 1901 gold zones and New Britannia property would also be processed at the New Britannia mill, by-passing the copper flotation circuit in the case of New Britannia zones and achieving similar gold recoveries as per historical performance of the mine.

1.8. MINERAL RESOURCE ESTIMATES

Lalor, 1901, Watts and Pen II

For the Lalor, 1901, Watts and Pen II properties, the construction of the 3D resource model and the estimation of mineral resources were performed by Hudbay personnel following Hudbay procedures in compliance with best industry standards and the CIM guidelines, using Leapfrog® version 4.2.3 and MineSight® version 15, two industry standard commercial geological and mining software.

The construction of the mineralized envelopes was solely based on the type of mineralization intersected. To this effect, assay intervals with high Au grade but no zinc or copper were excluded from base metal lenses and vice versa. Sample coding was checked to ensure proper tagging of the solids to actual drill hole locations. The mineral envelopes were used as hard boundaries in all cases for grade interpolation purposes.

Density values were generated from robust multi-regression formulas where no actual measurement was taken. Assay intervals were regularized by compositing drill hole data within each interpreted mineralized envelope. Although most of the drill holes were assayed in 1 m intervals, a composite length of 2.5 m was selected as more appropriate to conduct interpolation into the 5mx5mx5m block size selected by the mine engineering team as the optimum selective mining unit (SMU).

Exploratory data analysis (EDA) includes basic statistical evaluation of the composites for zinc (Zn), gold (Au), copper (Cu), silver (Ag) and density (SG). The EDA was conducted separately for each mineralized envelope and aimed to identify sub-domains that did not support the use of hard contacts for grade interpolation purposes but that justified being used for grade capping, block model validation, smoothing assessment and correction and resource classification.

The deciles^1^ analysis (Parish) method was used to define high-grade outliers and to confirm the need for grade capping. This analysis was conducted on the 2.5 m composites for each sub-domain independently. Down-hole and directional variograms for Zn, Au, Cu, Ag, and SG were created for each individual mineral envelope. A linear combination of a nugget and two nested spherical models were adjusted in all cases. Once generated, a systematic visual check was conducted to ensure that the search ellipsoid would be correctly oriented with respect to the geometry of the mineral envelopes.

The grade estimation process was validated for each envelope to ensure appropriate honouring of the input data and subsequent unbiased resource reporting and use of the model for reserve estimation. The validation steps included systematic visual investigations, checks for global bias and over-smoothing. Where the model was found to be over-smoothed, an indirect log-normal correction was applied to the kriged estimates.

The regression slope values obtained from the kriging of zinc (in the case of base metals lenses) or gold (in the case of gold zones) grade estimates was used as the primary criteria for resource classification. The block by block coding assignation was then smoothed to remove isolated blocks of one category within another.

For the present estimates, Hudbay has implemented a stringent approach to establish the potential for economic extraction of its resource reporting for underground deposits. With this approach, the potential for economic extraction of the mineral resource estimates at Lalor, 1901, Watts, and Pen II are reported within the constraint of a 'stope optimization envelope'. This excludes small isolated individual blocks above the economic cut-off criteria from the resource estimate and includes some 'geological dilution' that would need to be included in the economic envelope to maintain minimum spatial continuity requirements to define mineable shapes.

^1^ I.S. Parrish (1997) Mining Engineering Journal, Geologist's Gordon Knot: to cut or not to cut (pp. 45-49)

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

The parameters used as input to define the stope optimization envelope cover all the relevant technical and economic constraints including minimum stope and waste pillar dimensions and a NSR value calculation for each block based on anticipated metal recoveries, long-term metal price forecast and operating and capital costs based on the 2021 Lalor mine and Stall concentrator budgets. In all cases, two NSR values were calculated for each block to assess and compare the value of the blocks going to the Stall mill (no material difference between the two) or going to the new Britannia mill. The mineral resource estimates are reported to ensure that each potential stope would cover all its associated operating mining and milling costs.

The resulting mineral resource estimates for the Lalor, 1901, Watts and Pen II properties are presented in Table 1-1. All measured and indicated mineral resource estimates at Lalor and 1901 were either converted to mineral reserve estimates or deemed non-economic after applying the resource to reserve appropriate conversion factors.

TABLE 1-1: SUMMARY OF MINERAL RESOURCE ESTIMATES FOR THE LALOR, 1901, WATTS AND PEN II DEPOSITS (AS OF JANUARY 1ST, 2021)

Notes:

Totals may not add up correctly due to rounding.
Mineral resources that are not mineral reserves do not have demonstrated economic viability.
Mineral resources in the above tables do not include mining dilution or recovery factors.
Base metal mineral resources are estimated based on the assumption that they would be processed at the Stall concentrator while gold mineral resources are estimated based on the assumption that they would be processed at the New Britannia concentrator, which is currently being refurbished.
Metal prices of US$1.10 per pound zinc, US$1,500 per ounce gold, US$3.10 per pound copper, and US$18.00 per ounce silver with an exchange rate of 1.30 C$/US$ were used to estimate mineral resources at Lalor.
1901, Watts, and Pen II mineral resources were initially estimated using metal price assumptions that vary marginally over the assumptions used to estimate mineral resources at Lalor. In the QP's opinion, the combined impact of these small variations does not have any impact on the mineral resource estimates.
Lalor mineral resources are estimated using a minimum NSR cut-off of $105 per tonne for waste filled mining areas and a minimum of $116 per tonne for paste filled mining areas.
1901 mineral resources are estimated using a minimum NSR cut-off of C$140 per tonne.
Watts mineral resources are estimated using a minimum NSR cut-off of $150 per tonne, assuming processing recoveries of 90% for copper, 80% for zinc, 70% for gold and 70% for silver.
Pen II mineral resources are estimated using a minimum NSR cut-off of $75 per tonne and assume that the mineral resources would be amenable to processing at the Stall mill.

WIM and 3 Zone

Golder Associates ("Golder") prepared a resource model for the Wim property in 2015 and for the 3 Zone property in 2019 using Datamine® version 2.1.1547.0, an industry standard commercial geological and mining software. The procedures used to create and validate the resource block model are compliant with best industry standards and the CIM guidelines. A similar methodology was used in both cases.

A smooth and continuous 3D mineralized envelope was built using a copper equivalency cut-off grade in the case of WIM and a 2g/t gold grade in the case of 3 Zone to guide the construction of the mineralized envelope while including material below cut-off where necessary to maintain the continuity of the envelope. Scatter plots were used to identify assay outliers and for grade capping. Missing density values were generated using a linear regression formula from either the Fe grade or from the Cu grade for the WIM deposit while a constant density value was used for 3 Zone where there is no significant correlation between density and gold content. A composite length of 1 m was selected to conduct interpolation into the blocks in both deposits. Compositing was weighted by density at WIM.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

A block size of 15 m east-west (strike), 3 m north-south (thickness), and 5 m elevation was selected. Grade for Cu, Zn, Au, Ag and density (SG) was estimated using an inverse distance interpolant (IDW2) in the case of WIM and an ordinary kriging estimate in the case of 3 Zone in their respective unfolded space and interpolation was weighted by density in the case of WIM. A nearest neighbor (NN) interpolation was also performed to provide declustered sample grade for block model validation. The search ellipsoids were defined from pairwise variograms modeled by a linear combination of a nugget and nested spherical structures.

For WIM and 3 Zone, there are no mineral resource estimates being reported exclusive of the mineral reserve estimates. All measured and indicated mineral resource estimates that were deemed non-economic after applying the resource to reserve appropriate conversion factors were simply depleted from the resource inventory.

New Britannia Mine

For the former New Britannia, mine and its satellite gold deposits, the historical resource estimate performed by Kinross and by Alexis Minerals followed a conventional and industry standard approach and have been independently validated in 2018 by WSP Engineering ("WSP"). A block model of 15 ft. x 15 ft. x 15 ft. was created for the main mine upper and lower zones while a 10 ft. x 10 ft. x 10 ft. block model was created for the Birch Zone. The cut-off grades for the resource have been estimated over a 6-ft. minimum true width with a variable cut-off by zone as summarized in 1-6. The variation in the cut-off grade is related to new mining versus remnant mining. The New Britannia mill will have been refurbished before the mineral resource estimates would potentially be mined.

Given that WSP had to rely on historical documentation for some of the technical information supporting the estimation of the mineral resource estimates, the tonnes and grades previously estimated by Kinross and Alexis Minerals as measured and indicated resources were downgraded to an inferred category. Mineral resources that are not mineral reserves do not have demonstrated economic viability

TABLE 1-2: GOLD RESOURCE ESTIMATE FOR THE NEW BRITANNIA MINE AND BIRCH DEPOSIT ^(1)^^(2)(3)^

Totals may not add up correctly due to rounding.
Mineral resources that are not mineral reserves do not have demonstrated economic viability.
Mineral resources in the above tables do not include mining dilution or recovery factors.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

1.9. MINERAL RESERVE ESTIMATES

No mineral reserve estimates are reported for the Pen II, Watts, 1901 gold zones and New Britannia Mine deposits.

The following steps were followed in developing the reserve estimates at Lalor, 1901, WIM and 3 Zone:

Calculate two payable (NSR) values for each individual block in the resource model depending on whether processing would occur at the Stall concentrator or at the New Britannia concentrator, using long-term metal prices, concentrator recoveries, metal payability and downstream smelter treatment and refining costs assumptions.
Design stopes in the Deswik Stope Optimizer, considering depleted mineral resources, existing workings, resource categories and mine and mill operations costs. Dilution and recovery are estimated and applied at this step. Stopes are designed for both the Stall concentrator option and the New Britannia concentrator option.
Considering grades, value and location in the mine, assign stopes to either Stall or New Britannia concentrator.
Establish stope economics using a secondary NSR calculation where, along with mine and mill operations costs, mine capital, waste development and offsite administration costs are applied to each stope.
Assign whether stopes can be upgraded to mineral reserves based on resource classification.
Design ore development required for mining the reserves. Deplete development from the stopes. Interrogate grades of designed development for inclusion in mineral reserves. Sequence and schedule development and stope production for input to a financial Life of Mine (LOM) study to support mineral reserve economics.

The above methodology takes into consideration the different ore types and the milling options for the mine's future production and considers the various ore types found at these deposits.

The mineral reserve estimates exclude the mined out mineral resources, non-recoverable pillars (rib, post and sill) within mined out areas, mineral resources that are sterilized or not recoverable due to previous mining and stopes based on inferred mineral resource estimates.

A reconciliation between tonnes and grade predicted from the resource model and actual mine production credited by the mills for the first 9 months of 2020 shows a close comparison on all parameters. The reported tonnes and grade from the resource model are globally within 6% of actual production and on the conservative side for zinc and copper, i.e. under-estimate the quantity of metal recovered by the mill by approximately 10% reflecting the past focus for grade control on base metal production at Lalor which will evolve to gold going forward.

The author considers that the mineral reserves as classified and reported comply with all disclosure in accordance with the CIM guidelines requirements and definitions. The author is not aware of any mining, metallurgical, infrastructure, permitting or other relevant factors that could materially affect the mineral reserve estimate.

The Lalor mine, 1901, WIM and 3 Zone mineral reserves as of January 1, 2021 are summarized in 1-7.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

TABLE 1-3: SUMMARY OF LALOR, 1901, WIM AND 3 ZONE MINERAL RESERVES

Notes:

Totals may not add up correctly due to rounding.
Metal prices of US$1.10 per pound zinc, US$1,500 per ounce gold, US$3.10 per pound copper, and US$18.00 per ounce silver with an exchange rate of 1.30 C$/US$ were used to confirm the economic viability of the mineral reserve estimates.
Mineral reserves are estimated using a minimum NSR cut-off of $105 per tonne for waste filled mining areas and a minimum of $116 per tonne for paste filled mining areas.
For Lalor, individual stope gold grades were capped at 10 g/t, as a prudent estimate until reserves to mill reconciliations can establish that the high-grade gold can indeed be entirely recovered. This capping method resulted in the reduction of the global gold reserve grade by approximately 3%.
WIM mineral reserves are estimated assuming processing recoveries of 98% for copper, 88% for gold, and 70% for silver based on processing through New Britannia's flotation and tails leach circuits.
3 Zone mineral reserves are estimated assuming processing recoveries of 85% for gold based on processing through New Britannia's leach circuit.

1.10. MINING METHODS AND PRODUCTION SCHEDULE

Geology, geotechnical information, orebody geometry, productivity and mining experience are the key drivers to select the best suited mining methods at Lalor mine and the Snow Lake satellite deposits including considerations for lateral development, production, backfilling and ore transportation.

At Lalor, the hanging wall and footwall rocks are in general of good quality allowing the use of mechanized drilling and blasting techniques. The mineralized lenses dip at an average of 30°, but locally varying from 10 to 55°. Mining methods currently in use include cut and fill and longhole open stope. Paste backfill is used to increase recovery and accelerate the mining cycle. Lower grade areas will be filled with rock from waste development as all rock produced during the development cycle is disposed as backfill in the mining cycle.

Ore is mucked using Load Haul and Dump (LHD) loaders which are operated remotely in inaccessible areas. Ore is loaded into underground haul trucks at the stoping level by LHD's or from ore passes and transported to the ore handling system at the production shaft for hoisting to surface. Approximately 600 t0 1000 tpd of material is planned to be trucked to surface during the life of mine production schedule. Ore delivered to the production shaft is sized to less than 0.55 m by one of two rock breakers. Ore is hoisted from the mine by two 16 tonne capacity bottom dump skips in balance. Gold ore and base Metal ore is hoisted and stockpiled separately. On surface, ore is truck hauled to a primary crusher at the Chisel North mine site, crushed to less than 0.15 m, and will then be trucked to either the Stall, Flin Flon or New Britannia concentrator for processing. An external primary crusher is being commissioned at Stall mill that will service all ore types and once commissioned, will deem the primary crusher at Chisel North mine redundant and it will be removed from property.

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Mine development includes sub-horizontal tunneling and a ramp system to provide access between levels.

Main levels are developed parallel to and in the footwall of the ore zones. Where possible, main levels are located to provide access to multiple ore zones and are connected to haul ramps to allow mechanized equipment to travel from level to level. Main ventilation and ore pass raises are developed using a raisebore and/or Alimak climber. Drain, paste backfill and electrical cable holes are drilled from level to level using longhole or raisebore drills and are reamed to designed diameter.

Paste backfill is currently used at Lalor to increase recovery and accelerate the mining cycle and was a contributing factor to achieve a steady state of 4,500 tpd since the first quarter of 2019. Lalor mine production is approximately 20% by cut and fill miningand 80% by longhole bulk mining methods.

The Lalor mine production rate is expected to ramp-up to 5,300 tpd during the second half of 2022, following the closure of the 777 mine and the transition of personnel and equipment from the 777 mine to Lalor.

1901, WIM, and 3 Zone mining is based on development and stope designs used at 777 and Lalor mines, and the combined Life of Mine (LOM) production schedule is shown in Table 1-4. The Deswik software was used to assist with the LOM sequencing and scheduling to generate the production schedule. Historic mining rates and systems, equipment or crew's capabilities were applied and levelled. From this output, adjustments were made to further balance the capabilities to create the final plan.

TABLE 1-4: LOM PRODUCTION SCHEDULE

1.11. RECOVERY METHODS

The Stall concentrator complex is located approximately 16 km east of the Lalor Mine. Conventional crushing, grinding and flotation operations are used to process the ore. The nominal throughput rate of 3,500 tpd has been consistently exceeded in 2020 with production reaching 3,800t/d in a sustainable manner. The mill operates 24 hours per day, 365 days per year, with scheduled downtime for maintenance as required. The concentrator produces a copper concentrate with gold and silver credits and a zinc concentrate. Both concentrates are shipped by truck to Flin Flon. From there the copper concentrate is loaded onto rail cars and shipped to third party smelters. The zinc concentrate will continue to be trucked to Hudbay's zinc plant in Flin Flon which produces pure zinc metal until its planned closure in 2022. This refinery was refitted with the world's first two-stage pressure leach operation for recovering zinc from zinc sulphide concentrates in 1993. No roasters are used in this zinc extraction process and no sulfur dioxide (SO2) gas is produced. After 2022, the zinc concentrate will be loaded onto rail cars and shipped to third party smelters Tailings from the flotation circuit are utilized to produce a cemented paste backfill for use underground. Tailings not required for paste backfill will continue to be deposited into the existing Anderson TIA.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

The New Britannia mill is located approximately 16 km north of the Lalor mine. Conventional crushing, grinding, flotation, cyanide leach and recovery methods will be used to process the ore. The nominal throughput rate when the refurbishment program will be completed is 1,500 tpd starting in mid 2021. The plant will operate 24 hours per day for 362 days per year, with an availability of 92%. The concentrator will produce a copper concentrate with gold and silver credits and doré gold-silver bars. The copper concentrate will be shipped by truck to Flin Flon and will then be delivered to third party smelters. It is anticipated that, once New Britannia is operational, the copper concentrate produced at the Stall mill will be dewatered at New Britannia. The doré bars will be shipped offsite using conventional third-party transportation.

1.12. MARKET STUDIES AND CONTRACTS

Lalor produces the majority of its zinc concentrate and copper concentrate with gold and silver credits from its Stall mill. The refurbished New Britannia mill will process ore from the Lalor gold zone to produce a copper concentrate and gold-silver doré bar.

Currently, the Lalor zinc concentrate is delivered by truck to Hudbay's operations in Flin Flon and processed at the Flin Flon zinc plant into refined zinc metal and sold to customers in North America. The economic assumptions presume the same into 2022. During or beyond 2022, it is anticipated that Lalor's zinc concentrate will be sold to third party North American refineries.

The Lalor copper concentrate produced at the Stall mill is sold directly to a copper smelter in North America under a long-term agreement. The Lalor copper concentrate is delivered by rail. Lalor copper concentrate expected to be produced from the refurbished New Britannia mill has not been committed. It is anticipated that all sales contracts for the sale of copper concentrate produced from the refurbished New Britannia mill will be at standard industry terms.

The refurbished New Britannia mill will produce a gold doré commencing in 2022. It is forecast to contain approximately 69% silver, 26% gold and 5% other elements. The doré bar will be delivered and sold to refineries using conventional third-party transportation at standard industry terms.

Engineering, supply and construction contracts are initiated, managed and administrated by Hudbay's Manitoba Business Unit. Hudbay has a marketing division that is responsible for administering the company's marketing and sales of concentrates and metals. As well, Hudbay conducts ongoing research on metal prices and sales terms as part of normal business and long-range planning process to achieve market terms. Contract terms used in the Lalor financial evaluation are based on this research.

In 2020, Hudbay entered into a gold forward sale and prepay arrangement and received an upfront payment of $115 million to fund the refurbishment of the New Britannia gold mill. As a result, Hudbay is obligated to deliver a total of 79,954 gold ounces in 2022 and 2023 or approximately 3,331 gold ounces per month over a 24-month period from January 2022 to December 2023. The upfront payment was based on gold forward curve prices averaging approximately $1,682 per ounce.

1.13. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

The permits required for the current Lalor operation, including the Lalor mine, Stall concentrator and Anderson tailings facility have all been issued and remain valid.

At this time, there are no known environmental concerns which could adversely affect Hudbay's ability to operate the Lalor mine. Since the mine site is nearby existing facilities in the Snow Lake area, the Lalor mine was able to utilize infrastructure, services, and previously disturbed land associated with permitted, pre-existing and current mining operations in the Snow Lake area. The Lalor mine and associated projects are designed to minimize the potential impact on the surrounding environment by keeping the footprint of the operations as small as possible and by using existing licensed facilities for the withdrawal of water and disposal of wastes.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

Initial proposals for baseline work at Wim have been prepared by AECOM. Once complete these environmental studies will form the basis of the required approvals needed to advance this project should it be deemed viable.

3 Zone is part of the New Britannia site. Significant environmental studies of the area are available, and additional environmental assessments would be utilized to augment our understanding of property and any potential offsite impacts. Approvals to advance this project would be through Provincial regulators as part of an alteration of the existing Environment Act Licence for the property.

The 1901 deposit would leverage all existing surface and underground development near Lalor operations. Significant environmental baseline work has recently been conducted by AECOM and in conjunction with the significant number of past studies will be used to gain approvals for this development should it prove viable.

The re-commissioning of the New Britannia concentrator involves the placement of the Lalor gold ore tailings in the fully licensed Anderson tailings facility or the Lalor Paste fill plant via a pipeline connecting Stall concentrator to Lalor mine. All environmental permits are in place for the refurbishment of the New Britannia mill and associated infrastructure.

Initial environmental reviews and assessments will need to be completed at Watts. Permitting efforts will be completed prior to the project advancement. Hudbay plans to engage qualified consulting firms to carry out baseline studies when there is a better understanding on when the Watts potential mine production would best fit into the integrated business plan of the Snow Lake operations of Hudbay.

Hudbay engaged AECOM to carry out baseline studies for the Pen II property. Once completed, these studies will comply with the regulatory requirements and will be part of the permitting submission process should the Pen II Zone project be advanced.

Based on Hudbay's long-term (more than 50 years) mining experience in the Snow Lake region, and baseline studies to date, there is no known First Nation or Aboriginal hunting, fishing, trapping or other traditional use of the land in the zone of potential influence for the Lalor mine and associated facilities. Post closure, all water quality and earthen structures will be monitored and inspected in order to ensure the sites' conditions meet the applicable regulatory requirements.

1.14. CAPITAL AND OPERATING COSTS

The capital expenditures required to execute the mine plans at Lalor and 1901 amount to $598M (including $123M in growth projects) during the 2021-2025 period. Growth capital is represented by the New Britannia mill refurbishment and construction of a pipeline corridor to support the planned increase in gold production, as well as Stall recovery improvement projects. Also included in the five-year total mentioned above as sustaining capital is $54M for the development of the 1901 deposit from the Chisel-Lalor ramp.

From 2026 onwards, there is an additional $215M in sustaining capital required at Lalor and 1901 for capitalized mine development, replacement/acquisition of mining equipment and for other capitalized expenditures related to milling and environmental activities. Sustaining capital costs for the new mine plans aim at preserving mineralization with higher precious metals content for extraction following commissioning of the New Britannia mill. As a result, capitalized development costs are relatively higher during the first few years as priority is given to maximizing base metals output. Overall, this approach results in a superior economic value for the Snow Lake Operations.

The WIM mine development plan contemplates construction activities occurring in 2029, followed by commissioning in 2030 and ramp-up to maximum production rate by end of 2031. The capital expenditures required at WIM and at 3 Zone for refurbishing the existing mining infrastructures and for further developing the deposit totals $167M over the combined life of the projects.

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Operating cost estimates are based on the 2021 budget figures for Lalor and were developed by Hudbay with a bottom-up approach and using quotes from local suppliers, Manitoba operations experience, labor costs within the region, and actual 2020 costs. The operating costs include all the relevant mining, milling and refining activities as well as the attributable general and administrative costs to the Snow Lake complex production. Labour costs constitute the largest component of the mining and milling operating costs. Labour costs include both employee and contractor costs arising from mining, milling and maintenance activities. Until the closure of the zinc plant in 2022, the zinc concentrate produced from Stall will be refined into finished zinc to maximize the realizable value from zinc produced. An operating cost has been allocated to the Snow Lake operations for the refining of the zinc concentrates produced from Lalor mine. Post 2022, the zinc concentrate will be sold directly to market. General and administration costs are comprised of an allocation of shared services, administration, and other office costs to the Snow Lake operations based on relative usage versus the Flin Flon operations.

Table 1-5 illustrates that Snow Lake's significant by-product credits will reduce cash operating costs and sustaining cash costs on both a zinc and gold basis. During the first six years of operation with New Britannia (2022-2027), the Snow Lake operations will produce approximately 182,000 ounces of gold annually at a cash cost, net of by-product credits, of $US411/oz. This positions Lalor to be one of the lowest cost gold mines in Canada.

TABLE 1-5: CASH AND SUSTAINING UNIT CASH COSTS

1.15. INTERPRETATION AND CONCLUSION

Lalor Mine has been in production continuously since mid-2012 and has operated without major interruptions. As of the end of 2020, the mine has produced 8.8 million tonnes. After a continuous production ramp-up cycle, the mine has reached its steady state optimal ore production of around 4,500 tpd. More recently, during an unplanned operating shut down of 6 weeks at our 777 mine in Flin Flon, Hudbay temporarily re-allocated personnel and equipment from 777 to Lalor and achieved consistently during this period a production rate of 5,300t/d with a combination of hoisting and trucking. This performance has built sufficient confidence to include a ramp-up to 5,300t/d to the life of mine plan for Lalor starting in the second half of 2022 after the closure of the 777 mine.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

The strategy for the 2021 mine plan at Lalor is to mine base metals of higher grade zinc in 2021 while at the same time advancing the development required to start mining the gold and copper gold zones. Once the New Britannia mill becomes fully operational in the second half of 2021, the mine plan will target a production of 1,500 tonnes per day of gold ore, supplemented by 3,800 tonnes per day of base metal ore. The total production rate of Lalor is maintained at 5,300t/d from the second half of 2022 until 2026, after which the 1901 ore will supplement the base metal production to the Stall mill to allow a reduction in production at Lalor while maintaining the total Snow Lake camp production of 5,300t/d until 2029 followed by lower production rates until depletion of the mineral reserve estimates. While the Stall mill closes in 2031 based on mineral reserve estimates only, the New Britannia mill remains in operation for an additional six years until 2038 supported by the WIM and 3 zone mining production.

All the relevant capital and operating expenditures required to execute the life of mine plan of the Snow Lake operations have been accounted for in the demonstration of economic viability of the mineral reserve estimates including growth and sustaining capital costs and operating costs for the mining, milling and refining activities as well as the attributable general and administrative costs.

The author considers that the mineral resources and reserves as classified and reported comply with all disclosure in accordance with requirements and CIM Definitions. The author is not aware of any mining, metallurgical, infrastructure, permitting or other relevant factors that could materially affect the mineral reserve estimate.

The recent discovery of new zinc and gold mineralization zones at the 1901 deposit between Lalor and the former Chisel mine, the addition of mineral resource estimates at the Watts deposits and with its significant landholding position in the Flin Flon and Snow Lake belts, Hudbay remains in a strong position to continue to increase its resources and reserves in the short and medium term.

1.16. RECOMMENDATIONS

Significant value has been unlocked by leveraging Hudbay's exploration, technical and operational expertise, its existing portfolio of processing infrastructures and several inexpensive acquisitions.

On the exploration front, the author recommends to pursue in-mine exploration to incrementally add both base metal and gold rich inferred resources where lenses are still open in their along strike or down dip directions. The in-mine drill program should also continue to convert enough inferred mineral resource estimates to an indicated category so that they can be converted to mineral reserve estimates. In addition, surface exploration should continue along the Chisel-1901-Lalor trend which has the potential to define additional inferred resources in the Snow Lake camp over the next one to two years.

The Watts deposit located within 200km of the Snow Lake processing facilities offer additional opportunities to maximise the use of the Stall and New Britannia mills and/or extend their operating life.

As the Lalor mine matures and additional satellite ore sources are identified to supplement the feed blend to the Stall and new Britannia mills, several metallurgical projects are being advanced to match the configuration of the mill to the requirement of the ore including addressing the increasing Pb:Cu ratio with a traditional sequential Cu-Pb-Zn flotation circuit at Stall, increasing gold recovery at Stall, confirming the amenability of the 1901 gold lenses to processing at the New Britannia mill and improving flotation recoveries at Stall through reduction of the feed size.

Historical production at the New Britannia mill saw throughput rates of up to 2,000tpd based on the availability of ore from the underground operation. With the increase in the mineral reserves of the Lalor gold zones alongside additional satellite deposits, an opportunity exists to increase the throughput rate in the future. The study to date has identified that the tailings pipeline, pre-leach thickener and various minor pumping systems would require upgrade to accommodate the increased tonnage rate. In addition to the processing equipment upgrades the study includes the upgrade of the pipeline corridor road to accommodate ore haulage from Stall to New Britannia to remove traffic from the public access road.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

2. INTRODUCTION & TERMS OF REFERENCE

This Technical Report has been prepared for Hudbay to support the public disclosure of Mineral Resources and Mineral Reserves at the Lalor Mine and nearby Snow Lake satellite deposits and to provide an updated mine plan that contemplates 5,300 tpd of base metal, gold and copper-gold ore feed to the Stall and New Britannia concentrators.

Hudbay is a Canadian integrated mining company with assets in North and South America, principally focused on the discovery, production and marketing of base and precious metals. Hudbay's objective is to maximize shareholder value through efficient operations, organic growth and accretive acquisitions, while maintaining its financial strength.

Hudbay operates multiple properties in the Province of Manitoba. Operations near Flin Flon include the 777 Mine, an ore concentrator and a zinc plant with all required tailings disposal facilities. Current operations near Snow Lake include the Lalor underground mine as well as an ore concentrator, a tailings impoundment area and other ancillary facilities that support the operation. The property is located approximately 16 km by road west of the town of Snow Lake, Manitoba. Hudbay owns a 100% interest in the Lalor property through one Mineral Lease and eight Order in Council Leases to the south of the mine.

Refurbishing the New Britannia mill is expected to significantly increase gold production from Lalor and enable new gold and copper-gold exploration opportunities in the Snow Lake region by having an operating processing facility with substantially higher gold and copper recoveries. New Britannia was placed on care and maintenance in 2005 by its previous owner after producing 1.6 million ounces of gold. Refurbished, it will create additional value for Hudbay in the Flin Flon and Snow Lake regions as Hudbay pursues low-risk brownfield development opportunities.

The Wim deposit was acquired by Hudbay in the third quarter of 2018 for approximately C$0.5 million. Wim is a copper-gold deposit that starts from surface and is located approximately 15 km by road from the New Britannia mill. Hudbay has developed a mine plan for the joint operation of the WIM and 3 Zone deposit and conducted metallurgical testing to confirm the technical and economic viability of these two small mines accessible by underground ramp to feed the New Britannia mill after the richest portions of the Lalor reserves and resources have been depleted.

The recently discovered 1901 deposit located approximately half way between the former producing Chisel mine and Lalor mine will accessed by horizontal underground development from an existing ramp currently used for haulage of equipment and manpower between Chisel and Lalor. Prefeasibility level of engineering conducted on the measured and indicated mineral resource estimates of the 1901 deposit have confirmed the technical feasibility and economic viability of the project that will contribute to an increase in the total Snow Lake camp production from 4,500 to 5,300 tonnes per day between 2022 and 2027. The deposit hosts additional base metal and gold rich inferred resources that further enhance the production capacity and extend the life of the Snow Lake camp.

New Britannia is a former producing gold mine that produced approximately 600,000 oz between 1949 and 1958 and an additional 800,000 oz between 1995 and 2005. Significant mineral resources remain accessible at New Britannia as well as in the nearby Birch and No. 3 Zone with some investments in existing mining infrastructure. WSP was engaged in 2018 to audit and restate the historical resource estimates previously reported for these deposits. Hudbay continues to re-evaluate the technical and economic viability of the existing mineral resources and the potential to process this material at the New Britannia mill.

The Watts deposit is a medium copper-gold-zinc deposit accessible by ramp and located approximately 50 km Southeast of Snow Lake. Watts could constitute a supplemental source of feed for the Stall mill. Watts would be accessible by ramp and is 100% owned by Hudbay and free of any encumbrance. When the available capacity at the Stall mills for this deposits has been more firmly established, Hudbay will conduct in due course metallurgical testing, infill drilling and technical studies to confirm its technical and economic viability.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

Pen II is a low tonnage and high-grade zinc deposit that starts from surface and is located approximately 6 km by road from the Stall mill. Pen II could constitute a supplemental source of feed for the Stall mill.

2.1. QUALIFIED PERSON (QP)

This Technical Report has been prepared in accordance with National Instrument Form NI 43-101F1. The QP who supervised the preparation of this Technical Report is Olivier Tavchandjian, P. Geo., Hudbay's Vice-President, Exploration and Geology.

Mr. Tavchandjian is not independent of Hudbay, and this is not an independent technical report. Nevertheless, Hudbay is a "producing issuer" as defined in NI 43-101. As such, this Technical Report is not required to be prepared by or under the supervision of an independent QP.

Mr. Tavchandjian has been directly involved on a regular basis with the exploration, geology, resource-modelling, mine planning as well as with the estimation of operating and capital costs for the Lalor mine, WIM, 3 Zone and 1901 deposits as well as with the Watts and Pen II deposits. Mr. Tavchandjian has visited Lalor mine on a regular basis since September 2017 as well as the core storage facilities and the relevant internal and external laboratories and has directly overseen the mineral resource and reserve estimation process and the recent exploration drill programs conducted by Hudbay at the 1901, 3 Zone and Watts properties.

Todd McCracken, P. Geo for WSP has prepared the mineral resource estimates for the New Britannia mine site. Mr. Tavchandjian has reviewed the work conducted by WSP and is satisfied that the mineral resource estimates for the New Britannia properties were estimated in compliance with NI 43-101.

2.2. SOURCES OF INFORMATION

Sources of information for geological and mineral resources include: core drilling and sampling data, underground development and mapping, and assay and geochemistry analysis.

Sources of information for the mineral reserves include: the mineral resource block model, actual production and cost data, budget projections, life of mine inventory based on stope geometry parameters and mine development sequence.

Sources of information for the metallurgy, processing and economic analysis include: the actual operating data for the current processing facilities, extensive metallurgical test work and engineering pre-feasibility studies done by AECOM in 2018 for the refurbishment of the New Britannia mill and its associated infrastructures which were 75% completed by the end of February 2021 and by Stantec for the mine plan of the 1901 deposit.

2.3. UNIT ABBREVIATIONS

Units of measurement in this Technical Report conform to the SI (metric) system unless otherwise noted. Table 2-1 above, lists the notable unit abbreviations utilized in this Technical Report.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

TABLE 2-1: UNIT ABBREVIATIONS

2.4. ACRONYMS AND ABBREVIATIONS

Abbreviation of company names and other notable terms used in the report are as shown in Table 2-2.

TABLE 2-2: ACRONYMS AND ABBREVIATION

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2.5. OTHER CONTRIBUTIONS

TABLE 2-3: CONTRIBUTIONS FOR THIS REPORT

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3. RELIANCE ON OTHER EXPERTS

Standard professional procedures were followed when preparing the contents of this Technical Report. Data used in this Technical Report has been verified where possible and the QP has no reason to believe that the data was not collected in a professional manner and no information has been withheld that would affect the conclusions made herein.

The information, conclusions, opinions, and estimates contained herein are based on:

Information available to Hudbay at the time of preparation of this Technical Report, and
Assumptions, conditions, and qualifications as set forth in this Technical Report.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

4. PROPERTY LOCATION AND DESCRIPTION

4.1. LALOR MINE, 1901 AND PEN II DEPOSITS

The Lalor mine is located 16 km by road west of the town of Snow Lake at 54°52'24" north latitude, 100°09'01" west longitude and 303 mASL on National Topographic System (NTS) map 63K/16 (Figure 4 1).

Hudbay owns a 100% interest in the Lalor property through one Mineral Lease and eight Order in Council Leases to the south of the property totalling 946 hectares.

Hudbay holds the exclusive right to the minerals, other than quarry minerals, and the mineral access rights required for working the lands, mining and producing minerals from the Lalor mine. The surface tenure required for the mining operations have all been approved.

The 1901 deposit is located approximately 2.5km south-southeast of Lalor mine at 54°51.5'04.4" north latitude, 100°07'57.3" west longitude and 301 mASL on National Topographic System (NTS) map 63K/16. The deposit area is accessed by drill trail off the Lalor mine road to the south-southwest.

Hudbay holds the mineral rights to the claims that host the 1901 deposit and other contiguous claims in the area covering a total of 201 hectares.

The Pen II deposit is located 12 km west of the town of Snow Lake at 54°51'28" north latitude and 100°12'20" west longitude and 296 mASL on NTS map 63K/16.

Hudbay holds the mineral rights to claim CB146 which hosts the Pen II deposit. Other contiguous claims are held by Hudbay Minerals in the area and they comprise of one fraction and one regular claim for a total of 196 Ha (Table 4-1 and Figure 4-1)Table 4-1:

TABLE 4-1: LIST OF MINERAL LEASES, OIC AND MINERAL CLAIMS FOR THE LALOR MINE, 1901 DEPOSIT AND PEN II DEPOSIT

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

FIGURE 4-1: LOCATION OF THE LALOR, 1901 AND PEN II DEPOSITS MINERAL LEASES AND MINERAL CLAIMS

Land Use Permitting

General Permits are issued and administered by the Province of Manitoba Real Estate Services Division, Manitoba Department of Finance. Provided all terms and conditions of the General Permit are met, including payment of annual fees, the permit is automatically renewed for a 1-year term on an annual basis. Quarry Leases are issued and administered by the Province of Manitoba Mines Branch.

Two General Permits (GP59093 and GP63483) and one Quarry Lease are held by Hudbay and are required to carry on mining activities at Lalor. In detail, these are:

GP59093: This General Permit covers an area of 4.05Ha (4.0 km x 5m) and includes an all-weather access road, accommodates a 4 km 25 kV transmission line, a 4 km discharge line, a 4 km fresh water line, a 200m x 200m parking lot and an additional access road.

GP63483: The General Permit is related to the mine site.

QL-1928: QL-1928 was renewed November 26, 2017 for a 10-year term and provides the holder with the exclusive right to explore for, develop, and produce clay, gravel, rock or stone. QL-1928 will expire November 2027 and provided this lease is still required it can be renewed for another 10 year term.

General Permit and Quarry Lease status for the Lalor property is shown in Table 4-2

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

TABLE 4-2 GENERAL PERMITS AND QUARRY LEASE FOR LALOR

4.2. NEW BRITANNIA MINE AND OTHER SNOW LAKE GOLD PROPERTIES

Since 2015, Hudbay owns 100% interest in the Snow Lake Gold Properties through two Mineral Leases and 57 Mineral Claims covering a total area of 1,232 hectares. The New Britannia Mine-Mill Complex is located on the north side of the Town of Snow Lake at 54°53'10" north latitude and 100°01'24" west longitude and 293 mASL on NTS map 63K/16 (Table 4-2)

The two mining leases have been legally surveyed and remain in good standing providing that the annual taxes and lease fees are paid. The lease fees are CAN$ 12.00/Ha each year and are subject to a 1.38% NSR, based on production, payable to W. Bruce Dunlop.

The following summarizes the sequence of corporate transactions and mergers leading to the current ownership and consolidation of the property:

Prior to December 22, 2006, the mining leases and the Hudbay Option claims were held by 1126774 Ontario Limited (1126774 Ontario) which was held by Kinross and High River Gold Ltd. in a joint venture, with each party holding 50%.

On December 22, 2006, 1126774 Ontario was combined with Pegasus Mines Ltd. (Pegasus), with the merged entity retaining the name Pegasus.

Pegasus was wholly-owned by the New Britannia Mine Joint Venture (NBM-JV) which consisted of Piper Capital Inc. (Piper) owning 60%, and Garson Resources Ltd. (Garson Resources) owning 40%. Prior to the establishment of the NBM-JV, Garson Resources Ltd. held the Squall Lake claims.

On June 28, 2007, Piper and Garson Resources merged to form Garson Gold Corp.

On September 19, 2007, Pegasus changed its name to New Britannia Mines Ltd.

The terms of the agreement with Kinross have been described in the previous Micon Technical Report (Lewis, 2006).

At the time of the original transaction with Kinross, the claims were optioned to Kinross by Hudbay. Exercise of the option required a one-time cash payment of CAN$ 400,000 and a royalty consisting of 1.5% of the gross proceeds from the sale or disposition of all metals mined and removed from the optioned claims. However, Hudbay has retained the rights to the base metals on the mineral claims and should a base metal deposit be discovered and exploited on the claims, then Garson would retain a 1.5% royalty.

On December 31, 2007, the CAN$ 400,000 option payment was made and, subject to the 1.5% royalty, the Hudbay claims are now 100%-owned by NBM Ltd. Prior to conducting any production mining, the claims would have to be either incorporated into the existing mineral lease ML-61 or incorporated into a new mining lease. Micon indicated that due to the pre-existing royalty on ML-61, the claims should be incorporated into a new mining lease (Lewis, 2009). This would create a boundary which could be surveyed underground so that no conflicts between the two separate royalty holders would exist.

The Squall Lake group of claims north of Snow Lake comprises 14 contiguous, unpatented mining claims with an area of 899 Ha and is 100% owned by Hudbay, subject to various underlying royalties. The property contains a nets profits royalty (NPI) interest payable on commencement of commercial production of 4% to a private individual, an additional 6% NPI payable to American Barrick (now Barrick), and a 30% NPI payable to Barrick up to a maximum of $550,000. The property also contains a production royalty of $0.10 per ton payable to a private individual on products milled from 6 of the claims. No mineral resource estimates have been estimated in this Technical Report for the Squall Lake group of claims.

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The mineral claims require annual assessment work in the amount of CAN$ 25.00/Ha to keep them in good standing beyond their current expiry dates. In Manitoba, when the claim holder exceeds the minimum annual assessment expenditures for the mineral claims, surplus expenditures can be banked and utilized in future years to meet the annual assessment requirements.

Mineral lease ML-61 (NBM) and mineral lease ML-323 are located on the NTS map sheet 63K/16NE. The mineral claims currently held by Hudbay are located on NTS map sheets 63K/16NE, 63J/13NW, and 63K-16SE. The Squall Lake claims are located on NTS map sheet 63K/16NE Table 4-3 and Figure 4-2)

TABLE 4-3: LIST OF MINERAL LEASES AND MINERAL CLAIMS FOR THE SNOW LAKE GOLD AND SQUALL LAKE PROPERTIES

Project	Name	Number	Holder	Expiry	Hectares
Snow Lake Gold		ML323	Hudbay Minerals Inc.	3/28/2037	432
Snow Lake Gold		ML61	Hudbay Minerals Inc.	4/01/2034	830
Squall Area	CAM 1	CB11772	Hudbay Minerals Inc.	3/16/2022	195
Squall Area		CB4511	Hudbay Minerals Inc.	9/05/2026	43
Squall Area		CB4512	Hudbay Minerals Inc.	9/05/2026	152
Squall Area		CB4513	Hudbay Minerals Inc.	9/05/2026	119
Squall Area		CB6448	Hudbay Minerals Inc.	9/05/2026	33
Squall Area		CB7176	Hudbay Minerals Inc.	9/05/2026	33
Squall Area		CB7905	Hudbay Minerals Inc.	3/11/2022	45
Squall Area		CB7928	Hudbay Minerals Inc.	5/13/2022	42
Squall Area		CB7929	Hudbay Minerals Inc.	7/08/2026	43
Squall Area		CB8004	Hudbay Minerals Inc.	7/08/2026	43
Squall Area		CB8005	Hudbay Minerals Inc.	7/08/2026	43
Squall Area		CB8006	Hudbay Minerals Inc.	7/08/2026	43
Squall Area	KAP 1	CB8827	Hudbay Minerals Inc.	11/23/2024	48
Squall Area	DUB 11283	MB11283	Hudbay Minerals Inc.	10/13/2029	75
Squall Area	DUB 11284	MB11284	Hudbay Minerals Inc.	10/13/2029	53
Squall Area	DUB 11285	MB11285	Hudbay Minerals Inc.	10/13/2029	63
Squall Area	CAM 2	W46245	Hudbay Minerals Inc.	3/16/2026	17

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

FIGURE 4-2: LOCATION OF THE SNOW LAKE AND SQUALL LAKE MINERAL LEASES AND MINERAL CLAIMS

4.3. WATTS DEPOSIT

The Watts deposit is located approximately 52km east-southeast of the town site of Snow Lake, Manitoba and is centred on Latitude 54° 45' 51" North and Longitude 99° 15' 44" West and 260 mASL.

A group of six contiguous claims encompass the Watts River Property. The six claims cover an area of 1,282 hectares and are in good standing. Hudbay Minerals Inc. holds 100% of the mineral rights (Table 4-4).

TABLE 4-4: LIST OF MINERAL CLAIMS FOR THE WATTS DEPOSIT

Project	Name	Number	Holder	Expiry Date	Hectares
Watts Deposit	KUS 1732	P1732F	Hudbay Minerals Inc.	5/24/2024	197.7
Watts Deposit	KUS 1654	P1654F	Hudbay Minerals Inc.	4/19/2023	166
Watts Deposit	KUS 1818	P1818F	Hudbay Minerals Inc.	5/01/2022	255.2
Watts Deposit	KUS 1731	P1731F	Hudbay Minerals Inc.	5/24/2022	157.5
Watts Deposit	KUS 1816	P1816F	Hudbay Minerals Inc.	5/01/2027	227
Watts Deposit	KUS 1822	P1822F	Hudbay Minerals Inc.	5/01/2030	262.4

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

FIGURE 4-3: LOCATION OF THE WATTS DEPOSIT MINERAL CLAIMS

4.4. WIM DEPOSIT

The Wim deposit is located 16 km north of the Town of Snow Lake at 55°1'34" north latitude and 100°2'50" west longitude and 316 mASL on NTS map 63N/01(Table 4-5).

The Wim deposit is 100% owned by Hudbay. On October 31, 2018 it was announced that Alexandria had entered into a purchase agreement with Hudbay to sell the Hudvam and Wim properties. Previously, on December 22, 2014, Alexandria signed an Arrangement Agreement with Murgor Resources Inc. (Murgor) to acquire all Murgor's shares and assets (the acquisition was completed on March 9, 2015).

Hudbay holds the mineral rights to 21 claims for a total 3,149 Ha. The Wim deposit is hosted by claims W49947 and W49948 (Table 4-5).

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

TABLE 4-5: LIST OF MINERAL CLAIMS FOR THE WIM DEPOSIT

FIGURE 4-4: LOCATION OF THE WIM DEPOSIT MINERAL CLAIMS

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1. ACCESSIBILITY

The Lalor mine, its nearby satellite deposits and all the infrastructures associated with operations at Snow Lake are located approximately 200 km by road east of the city of Flin Flon and are within or close to the town of Snow Lake, Manitoba. Access to the mineral deposits is described individually below (Figure 5-1)

Access to Lalor mine is by travelling ~29.5km north along Provincial Road (PR) #392 off from PR #39 towards the Town of Snow Lake. Turn left on PR #395 for 9km till Lalor mine access road.

Access to the 1901 deposit site is by traveling to the end of PR #395 and taking a winter road across Chisel Lake or a snowmobile / quad trail for ~5km from around the past producing Chisel open pit to the site.
Access to the Pen II site is from Provincial Roads #392 and #395, located past the old Chisel mine site.
The New Britannia Mine-Mill Complex is located on the north side of the Town of Snow Lake and accessed from paved Cedar Avenue.
Access to the Wim site is from the Town of Snow Lake, travelling north from the New Britannia mine site and over a tailings dam road along the Manitoba Hydro Transmission line.
Access to the Watts River deposit is located approximately 52 km east-southeast of the town site of Snow Lake, Manitoba and is accessed by helicopter or a 17km winter road northeast from Ponton, Manitoba, where Provincial Highway #6 meets Provincial Highway #39. Alternate access can be obtained by utilizing the Manitoba Hydro powerline route that provides Snow Lake with power and passes within a few kilometres of the Property.

5.2. CLIMATE

The Snow Lake area has a typical mid-continental climate, with short summers and long, cold winters. Climate generally has only a minor effect on local exploration and mining activities. The nearest Environment Canada weather station is located near Baker's Narrows at the Flin Flon airport, approximately 16 km southeast of Flin Flon and approximately 100 km west of Snow Lake. The average annual temperature at the Baker's Narrows weather station is 0.1°C. The average summer temperature is approximately 17°C, and the average winter temperature is -14°C. The lowest monthly average temperature occurs in January at -21.1°C and the highest monthly average temperature is in July at 18.3°C. Freeze-up of small bays and lakes occurs in mid-November, with breakup occurring in mid-May. There is an average of 115 frost-free days. On average, 45.7 cm of precipitation falls annually, of which, 35% is snow. Since 1960, monthly precipitations have ranged from zero to a high of 18.1 cm, with a maximum daily precipitation of 7.82 cm. Average monthly winds for the area range from 10km/hr to 13km/hr, with 40% of the winds originating from the northwest, northeast or north.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

FIGURE 5-1: SNOW LAKE REGIONAL MAP

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5.3. LOCAL RESOURCES

The nearest community is the town of Snow Lake, Manitoba. The community of 899 (Statistics Canada, 2016 census) has 498 private dwellings. There are two cottage subdivisions located on Wekusko Lake along PR #392, as well as residences at Herb Lake Landing, located approximately 40 km south of the town. There are also a small number of seasonal remote cottages located near lakes throughout the area.

Snow Lake community services include a health facility staffed with two doctors, an ambulance and a fire truck, a grocery store, two hotels/motels, three service stations.

The nearest larger centres (5,000+ residents) are Flin Flon (208 km), The Pas (219 km) and Thompson (260 km), all accessible by paved highway. There is a 1,100m by 20m un-serviced gravel municipal airstrip located approximately 16 km from Snow Lake along PR #393. A small craft charter service operates out of the community of Snow Lake, where small planes and helicopters can be chartered. Rental vehicles are available at the Flin Flon airport. The nearest full-service commercial airport is located at Baker's Narrows near Flin Flon, approximately 185 km from Snow Lake. The nearest international airport is located in Winnipeg, approximately 700 km from Snow Lake.

There is no rail in the immediate area of Snow Lake. The nearest rail access is at Wekusko siding, approximately 65 km southeast of Snow Lake. Wekusko is accessible by an all-weather road. A gravel rail bed (ties and rail removed) connects the Stall concentrator to Chisel Lake mine and continues to a rail line at Optic Lake siding, approximately 65 km west of Chisel Lake. Optic Lake is not road accessible.

To house non-local employees during their work rotations, the company provides a camp located in the Town of Snow Lake which services Hudbay employees and contractors for the mine and mill operations.

Outland Camps in Snow Lake currently consist of an upper camp section with 7 dorms and a lower camp section with 4 dorms and a Kitchen and dining hall. The camp has a total capacity of 436 rooms. There is a recreation trainer that has recently been turned into a Covid-19 testing facility and an exercise trailer currently closed due to Covid-19.

Utilities are provided by the town of Snow Lake, including potable water, sewage and electricity. Heating is electrical and propane. The camp includes parking spots for employees.

5.4. EXISTING INFRASTRUCTURE

Hudbay operates a zinc metallurgical plant in Flin Flon, Manitoba, approximately 200 km from Snow Lake with a capacity of 115,000 tpa of refined zinc. However, Hudbay intends to cease operations and put the zinc plant on care and maintenance at the end of the second quarter of 2022.

Hudbay also operates two ore concentrators: The Stall mill located approximately 16 km from Lalor and the Flin Flon mill is located in Flin Flon. The Stall mill is currently operating seven days per week, producing at or above 3,800 tpd. During the 2021-2022 period, although 100% of Lalor's production will be processed at the Stall or New Britannia mills, any additional production could be trucked to the Flin Flon mill until this facility shuts down at the same time as the zinc plant closes. The Stall mill has two circuits with design capacities of 900 tpd and 2,180 tpd. At the Flin Flon mill, the ore from Lalor has been in the past processed separately as it is free of the encumbrances that affect the other ore processed in Flin Flon from the 777 mine, i.e. ore subject to the Silver Wheaton precious metal streaming agreement.

Both concentrators have two flotation circuits producing a zinc concentrate and a copper concentrate. The zinc concentrate produced at Stall is hauled by truck to Hudbay metallurgical facilities in Flin Flon. The copper concentrates are sold to market. The tailings from the Stall mill are deposited in the Anderson Tailings Impoundment Area or used at the paste plant. The permitted Hudbay Anderson TIA, located approximately 13 km from Lalor, is used for tailings disposal.

The Lalor paste plant was commissioned in the third quarter of 2018, is located northeast of the existing head frame complex, and is capable of delivering 165 tph solids (tails) or 93 m3/hr of paste. The paste plant is capable of varying the binder content in the paste to provide flexibility in the strength gained from the paste where higher and early strength may be required depending on mining method.

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Tails are currently pumped from the Stall concentrator to the Anderson TIA and diverted, as needed, to the Anderson booster station. Two pipelines are installed between the Anderson booster pump station and the paste plant located at Lalor mine site, which are situated approximately 13 km apart.

General area infrastructure includes provincial roads, 115 kV power to within four kilometres of Lalor provided by Manitoba Hydro and telephone land line and cellular phone service provided by Manitoba Telecom.

The Lalor mine is located 3.5 km from the Hudbay Chisel North mine. Chisel North infrastructures include a mined out open pit used for waste rock disposal, fresh (process) water sources, pumps and waterlines, 4160V and 550V power, mine discharge water lines, a 2,500 gpm water treatment plant with retention areas, and mine buildings including offices and a change house. These facilities are used for geological core processing, surface mobile equipment shop, project offices, and the crushing of Lalor ore. The Chisel site is also the location of two electrical transformers 115 kV to 25 kV that feed Lalor mine.

The infrastructure on site at Lalor includes, the main office change house, headframe, hoistroom, down cast fans, exhaust fans, main pump station, potable water treatment plant, sewage treatment plant and several other smaller buildings for purposes such as health and safety and additional change house space. The New Britannia mill is located approximately 16 km north of the Lalor mine and it is currently under construction (refurbishment) with a planned start-up of operations in the spring of 2021. The mill infrastructure consists mainly of three buildings and houses the crusher and mill/gold circuits. There are two existing crusher buildings: primary and secondary, each approximately of 108m2 and 136m2 respectively. The primary building houses the jaw crusher, and the secondary building houses the cone crusher and screen deck, each building with associated conveyors. The existing mill building is approximately 1,939m2 and contains the grinding and gold circuits, elution, EW and refining. The outside equipment includes thickener and leach tanks.

The 1901 deposit is located approximately 1,000m half-way off the ramp that connects the Chisel and Lalor mines. The deposit would be accessed by lateral development from the ramp and would use the existing surface infrastructures.

The 3 Zone deposit is located on the past producing New Britannia mine site in the town of Snow Lake and has existing infrastructure of a ventilation raise, decline and underground workings to approximately the 130mL. A total of 219,152 tonnes was mined from this area from 1995-1996. The New Britannia mill is within 2km on an all-weather road.

The WIM deposit is located approximately 16km north of the town of Snow Lake and is currently within trucking distance to either Stall or New Britannia mill. Currently no infrastructure exists on the claims but available in Snow Lake and can be brought into the property.

The Pen II deposit is located approximately 4km west southwest of Lalor mine and 12km west of the Town of Snow Lake. Currently no infrastructure exists on the claims but is available in Snow Lake or from Lalor mine site and brought into the property.

Access to the Watts property is via helicopter or a 17km winter road northeast from Ponton, Manitoba, where Provincial Highway #6 meets Provincial Highway #39. The winter trail utilizes the Manitoba Hydro powerline route that provides Snow Lake with power and passes within a few kilometres of the property.

In all cases, the surface rights are sufficient to accommodate mining activities. Given the fact that ore from all the deposits will be processed at either Stall or New Brit mills, the power and water requirements will be sufficient for the treatment of ore and the disposal of tailings and waste material. Given their relative proximity to Snow Lake, mining personal availability will not be an issue.

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5.5. PHYSIOGRAPHY

The Snow Lake mining camp is located in the Boreal Shield Ecozone, the largest ecozone in Canada, extending as a broad inverted arch from northern Saskatchewan east to Newfoundland. This area is surrounded by water bodies including the Snow, File, Woosey, Anderson and Wekusko Lakes, all located in the Churchill River Upland Ecoregion in the Wekusko Ecodistrict.

The dominant soils are well to excessively drained dystic brunisols that have developed on shallow, sandy and stony veneers of water-worked glacial till overlying bedrock. Significant areas consist of peat-filled depressions with very poorly drained typic and terric fibrisolic and mesisolic organic soils overlying loamy to clayey glaciolacustrine sediments.

Topography shows gentle relief that rarely exceeds 10 m at approximately 300 mASL, with depressional lowlands.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

6. HISTORY

6.1. LALOR & THE CHISEL BASIN AREA

Exploration in the Lalor-Chisel area began in the 1950s. The Chisel Basin area hosted four previously producing mines; Chisel Lake, Chisel Open Pit, Photo Lake and Chisel North. This basin is also the host of the Lalor deposit.

In early 2007, hole DUB168 was drilled almost vertically to test a 2003 surveyed Crone Geophysics deep penetrating pulse electromagnetic anomaly and intersected 45 m of mineralization starting at a downhole depth of 782 m. Drilling at Lalor has been continuous since the discovery of mineralization on the property and as of the date of this Technical Report, totals in excess of 500,000 drilled metres.

Lalor commenced initial ore production from the ventilation shaft in August 2012 and achieved commercial production from the main shaft in the third quarter of 2014. Table 6-1 summarizes the ore production by year at Lalor from 2012 to 2020.

TABLE 6-1: LALOR MINE PRODUCTION

In February 2019, Hudbay exploration discovered massive sulfide mineralization at the 1901 deposit. Initial inferred mineral resource estimates were declared for this deposit in August 2019 for two base metal lenses. Follow-up drilling in 2020 resulted in upgrading base metal inferred resource estimates to measured and indicated categories and adding new inferred mineral resources both for the known two base metal lenses and for two new gold rich lenses. In 2020, Hudbay also completed pre-feasibility study on the engineering and mineral processing of the 1901 measured and indicated mineral resources.

6.2. NEW BRITANNIA AND SNOW LAKE DEPOSITS

Gold was first discovered in 1914 approximately 20 km to the southeast of Snow Lake on the eastern shores of Wekusko Lake, also known as Herb Lake. In 1917, the Moose Horn-Ballast claims produced the first gold in Manitoba. First significant industrial gold production came from the Laguna mine which was in production intermittently until 1939 and produced 59k oz of gold.

Mine construction at the New Britannia site first started in 1945. In March 1949, the mine was opened as Nor-Acme mine. Production continued until 1958. 4.9Mt were mined at an average grade of 4.4g/t and the Nor-Acme mill recovered approximately 610,000 oz of gold during this production period.

Between 1958 and the early 1990s, various companies worked on the property with the intention to bring the mine back into production, however, these attempts were unsuccessful.

TVX and High River formed a joint venture in 1994 to reopen the mine. TVX became the operator of the mine in late 1994 and ore from the secondary mining areas (No. 3 Zone and Birch pit) located on mineral lease ML-323 was used to bring the mine and mill into production prior to the completion of the rehabilitation and construction of the main shaft.

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The New Britannia mill poured the first gold bullion in November 1995 which was obtained from the gold produced at No. 3 Zone. When production commenced in 1995, the name of the Nor-Acme Mine was changed to the New Britannia Mine. Production ceased at No. 3 Zone in May 1996 and production from the Birch pit was limited to the spring and summer of 1996. Full production from the main shaft was achieved in August 1996. Through various transactions, Kinross became the operator of the New Britannia mine-mill complex. Production mining ceased at the end of September 2004 and the mill was put on care and maintenance in 2005 due to a low gold price environment after producing 1.6 million oz of gold.

6.3. WIM DEPOSIT

Initial exploration on the property was comprised of airborne electromagnetic (EM) and radiometric surveys carried out by HBED in 1961, which identified several geophysical conductors. An extensive program of staking, line cutting, ground geophysical surveying and diamond drilling over high priority anomalies immediately followed.

VMS mineralization was discovered on the Wim deposit in December 1961 by diamond drilling of a ground EM anomaly delineated during the ground follow-up of an airborne EM survey. A total of one hundred and fifty-six (156) drillholes for 38,185 meters were completed from 1962 to 1991. In 2006 Murgor Resources Inc. optioned the property from Hudbay and completed a drilling program from 2007 to 2010 for sixty-four (64) diamond drill holes for a total of 19,332 metres. Most of this drilling was focused on defining the know mineralization and completing a NI 43-101 resource estimate. This historical mineral resource estimate included an indicated resource of 3.9 million tonnes @ 1.7% Cu, 0.3% Zn, 1.6g/t Au and 6.7g/t Ag and inferred mineral resource of 0.7 million tonnes @ 1%Cu, 0.4% Zn, 1.8g/t Au and 4.7g/t Ag.

In 2019, Hudbay completed seven (7) metallurgical holes for a total of 2,533 metres. Mineralization from these holes was sent off for metallurgical testwork. The WIM deposit hosts a copper-gold rich sulphide lens within amphibolite grade altered volcanic rocks.

6.4. PEN II

The first surface geophysics completed at Pen II was a Turam survey done in 1973. Early drilling in 1970's focused on this geophysical Turam survey, without significant results. Two airborne surveys were subsequently completed in 1984 and 1999. Hudbay completed Thirty-nine (39) holes between 1983 and 2007 for a total of 8,764 metres including the discovery hole which followed up an airborne and HLEM anomaly on the property. Callinan Mines Limited option the property in 1999 to 2000 and drilled three (3) holes for a total of 893.3 metres. Callinex later dropped the option. Between 2017 and 2019, Hudbay Exploration completed sixty-two (62) diamond drill holes to expand and define the resource limits for a total of 11,620 metres.

6.5. WATTS

The property was first staked in 1995 by HBED to cover anomalies located by a 1993 SPECTREM airborne EM survey. Diamond drilling of an HLEM anomaly in 1995 by HBED resulted in the discovery of the Watts River Zone. HBED drilled a total of 9,930m in 25 holes on the Watts River deposit between 1995 and 1997.

The project was optioned to Fort Knox - Berland Resources in 2000/2001. They completed seven (7) drill holes for a total of 2,476m. Only two holes were drilled on the main horizon, while most of the program focused on the newly discovered east zone located 400m in the hanging wall of the main zone. HBED implemented a new drilling program beginning in 2005 as a follow up to previous drill results and drilled a total of 9,284m in 20 holes between 2005 and 2007. In 2019 Hudbay completed six (6) holes for 2,673 metres testing the extension of the mineralized envelopes

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

7. GEOLOGICAL SETTING & MINERALIZATION

7.1. REGIONAL GEOLOGY

The properties of interest for this Technical Report are all within the Paleoproterozoic Flin Flon Greenstone Belt (Figure 7-1) and are overlain by a thin veneer of Pleistocene glacial/fluvial sediments. Located within the Trans-Hudson Orogen, the Flin Flon Greenstone Belt consists of a variety of distinct 1.92 to 1.87 Ga (billion years) tectonostratigraphic assemblages. These include juvenile arc, back-arc, ocean-floor, ocean-island and evolved volcanic arc assemblages that were amalgamated to form the Amisk accretionary collage prior to the emplacement of intermediate to granitoid plutons and subsequent deformation (Syme et al., 1998). The volcanic assemblages consist of mafic to felsic volcanic rocks with intercalated volcanogenic sedimentary rocks.

At the eastern end of the Flin Flon Greenstone Belt, the Snow Lake area includes a lithologically and structurally diverse sequence of deformed and metamorphosed volcanic, sedimentary and intrusive rocks. This sequence is dominated by 1.84-1.81 Ga fold-thrust-style tectonics and by a southwest-verging allochthon of volcanic rocks, the Snow Lake arc assemblage.

The Snow Lake arc assemblage that hosts the producing and past-producing mines in the Snow Lake area is a 20 km wide by 6 km thick section that records a temporal evolution in geodynamic setting from 'primitive arc' (Anderson sequence to the south) to 'mature arc' (Chisel sequence) to 'arc-rift' (Snow Creek sequence to the northeast, Bailes and Galley, 2007). The Anderson and Chisel sequences represent the geological units of interest (Figure 7-2).

The Anderson sequence is a bimodal primitive arc sequence dominated by basaltic andesite, which contains at least three rhyolite flow centres: Daly, Sneath and Anderson-Stall. The mafic and felsic volcanic rocks of the Anderson sequence are intruded by the Sneath Lake subvolcanic intrusive complex (Figure 7-2). The Anderson-Stall and Daly rhyolites occur directly above the Sneath Lake intrusion near its two ends.

The Watts River, deposit is located in the eastern end of the Paleoproterozoic meta-volcanic rocks of the Flin Flon Domain. Due to a thick and pervasive cover of Quaternary glacial deposits and a thin veneer of Ordovician limestone the Paleoproterozoic rocks in the vicinity of the deposit are mainly unexposed. The meta-greywacke of the Central Kisseynew subdomain was extrapolated from their exposures on Grass river beneath the Watts River area as far south as North Moose lakes (Leclaire, et al., 1997). The Watts River deposit is likely hosted by the greywacke gneisses of the Kisseynew domain.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

FIGURE 7-1: GEOLOGY OF MANITOBA

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FIGURE 7-2: SNOW LAKE AREA (AFTER BAILES AND GALLEY, 2011)

VMS Deposits

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Cu-Zn-rich deposits and occurrences in the Anderson sequence are confined almost entirely to within the Anderson-Stall and Daly rhyolite centres. All the developed deposits are within the larger Anderson-Stall rhyolite. The Sneath Lake synvolcanic tonalite in the Anderson Lake area shows a steady increase in both intensity of alteration of supracrustal rocks and of the Cu/Zn ratio (Cu/Cu+Zn) contained in the VMS deposits as the distance to the pluton decreases. The variation of alteration and metal Cu/Zn ratio is interpreted to be the result from higher temperature for pluton-proximal portions of the regional synvolcanic hydrothermal system (Bailes and Galley, 1996).

There are seven Zn-rich VMS deposits located within the mature arc Chisel sequence: Lalor, Chisel, Chisel North, 1901, Pen II, Photo and Ghost. The seven deposits generally overlie the Powderhouse dacite and local discrete rhyolite flow complexes. They have been subjected to polyphase deformation.

The Wim deposit, located north of Snow Lake, consists of one large copper-gold (zinc and silver) VMS lens occurring in an early Proterozoic island-arc assemblage of felsic and mafic volcanic rocks that stretches for an exposed length of 250 km east-west and 75 km north-south, on the western flank of the Herblet gneiss dome (Figure 7-2).

The Watts River Deposit is located 52km east-southeast of the town of Snow Lake. The deposit lies in the eastern portion of the Flin Flon-Snow Lake Greenstone Belt. The Watts River deposit is a stratabound accumulation of sulphides that precipitated at or near the seafloor in association with contemporaneous volcanism. The depositional environment is similar to the base metal deposits of the Snow Lake mining camp. The deposit appears to be distal with no associated alteration pipe recognized to date. The sulphide mineralization is interpreted as three lenses comprised of coarse-grained sulphides with a trend of azimuth of 020° and dip between 65 and 75° to the east.

Snow Lake Gold Deposits

The Amisk mafic and felsic volcanic hosted deposits lie to the South of the Missi group hosted deposits of Squall Lake. The McLeod Road Thrust (MRT) is one of the main structural features of the area. The deposits are bound to the west by the MRT which is responsible for the emplacement of ca. 1.9-1.89 Ga volcanic and volcaniclastic rocks of the Snow Lake arc assemblage (Bailes and Galley, 2007) on top of the ca. 1.85 Ga Burntwood group metasedimentary rocks (Figure 7-2).

The southern part of the property comprises a moderately north-dipping and dominantly north facing bimodal sequence of coarsely re-crystallized and deformed mafic and felsic volcanic and volcaniclastic, epiclastic rocks and gabbro. The Birch Lake Thrust fault to the east was emplaced ca. 1845 Ma and thrust the Missi Group metasediments above the volcanic and volcaniclastic rocks.

The southern volcanic sequence is intruded by pre and syn ultramafic and mafic intrusions, and syn to post-tectonic granitoids. The metamorphic assemblages are of lower to mid-amphibolite facies. Imbrication of the sequence bound by the MRT and BLT occurred during south-directed transport and resulted in the formation of a strong regional northeast plunging stretching lineation, a regional foliation and the Nor-Acme anticline. These structures were subsequently folded around the northeast-trending Threehouse synform. (Rubingh et al. 2013).

Mineralization of the lode-gold vein-type deposits hosted in the Amisk group mafic and felsics volcanic rocks are structurally controlled and associated with shear zones, faults, fold hinges and axial planes that host simple to complex vein systems. The mineralization is associated with lithological contacts of contrasting properties in the sequence of interlayered volcanic and volcaniclastic rocks. Alteration in the mineralized zones consists of quartz-carbonate-mica, with arsenopyrite as the primary sulphide accompanying the gold.

The Squall Lake area (4 km northwest of the Town of Snow Lake) is comprised of several mineralized zones that are hosted in Missi group rocks which are the product of fluvial and alluvial sedimentation. In Snow Lake, the Missi Sequence is composed of cross bedded arenites and arkose. The Missi group is commonly intruded by discontinuous gabbroic sills and dykes that are medium to coarse grained. The Missi group arenites were thrusted over the garnet staurolite schist of the Burntwood group (Beilhartz, 2006). The lowermost sill, a 20 - 40-metre-thick amphibolitic unit, is spatially associated with the known gold mineralization of the Squall lake area. This epigenetic gold mineralization is believed to be associated with structures generated by over-thrusting related to the McLeod Road Thrust Fault.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

7.2. PROPERTY GEOLOGY

Lalor

The Lalor deposit is similar to other massive sulphide bodies in the Chisel sequence (Chisel Lake, Ghost Lake, Chisel North, and Photo Lake), and lies along the same stratigraphic horizon as the Chisel Lake and Chisel North deposits. It is interpreted that the top of the zone is near a decollement contact with the overturned hanging wall rocks, the Chisel-Lalor Thrust (Figure 7-3). The most common dyke intrusion throughout these rocks is a fine-grained feldspar-phyric gabbro to diorite.

FIGURE 7-3: LALOR MINE CROSS SECTION WITH DRILL HOLE TRACES

The extensive hydrothermal alteration and metamorphic recrystallization of the footwall rocks has produced some exotic aluminous mineral assemblages. These assemblages include chlorite and sericite dominant schists and cordierite + anthophyllite gneisses. Other minerals indicative of hydrothermal alteration that occur extensively throughout these rock assemblages include quartz, feldspar, kyanite, biotite, garnet, staurolite, hornblende, and carbonate. Clinopyroxene, gahnite and anhydrite also occur locally. These assemblages are typical of metamorphosed footwall hydrothermal alteration commonly associated with VMS deposits and are similar to that at the other massive sulphide deposits in the Chisel Lake area.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

Snow Lake Gold Properties

The Snow Lake gold properties consists of the New Britannia Mine, No. 3 Zone, Birch Zone, Boundary Zone and Squall Lake deposits including the K1, K5, K7, K10, K11, Margaret (North and South), Margaret Extension, South, F1, Moon Gertie Zones, Bay and North Zones.

The New Britannia mine is a classical lode gold, quartz-carbonate vein type deposit. The mineralization is associated with shear zones, faults, and simple to complex vein systems. The auriferous zones (Toots, Dick, Ruttan and Hogg) lie along a curvilinear shear zone named the Howe Sound fault also referred to as the Nor-Acme fault (Figure 7-4). The main mineralized zones consist of sheared or "mylonitic" quartz-carbonate-mica alteration with arsenopyrite mineralization accompanying the gold emplaced in a simple intercalated sequence of altered felsic and mafic volcanics and pyroclastics of the Amisk Group. The host rocks are altered and include varying proportions of quartz and carbonate.

The gold bearing horizons in the No. 3 Zone are fault-hosted vein occurrences, consisting of a main shear vein and numerous extensional ladder veins. Gold mineralization is located within a 50 m wide fault zone at the folded contact between the coarse pyroxene mafic volcaniclastic rocks and the plagioclase-phyric pillowed basalt flows.

FIGURE 7-4: NEW BRITANNIA REPRESENTATIVE CROSS SECTION WITH DRILL HOLE TRACES

The Birch Zone is interpreted as hosted in a silicified shear zone and several parallel and sub parallel shears cut the mafic volcanic, mafic volcaniclastic and metasedimentary rocks (including some graphitic argillite) which were intruded by gabbroic dykes. Alteration and mineralization at the Birch Zone is similar to the No. 3 Zone.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

The Boundary Zone is located 0.5 km south of the No. 3 Zone and approximately 1 km north of the New Britannia Mine, and in the hanging wall of the McLeod Road Thrust fault. The lode-type gold deposit is located along the Nor-Acme anticline at the contact between mafic and felsic volcanics.

The Squall Lake area is host to several mineralized zones in Missi group rocks. Epigenetic gold mineralization on the property is believed to be associated with structures generated by over-thrusting related to the McLeod Road Thrust Fault (Beilhartz, 2006). Two economically significant gold mineralized sets of structures exist on the property. The first type of structure was generated by an initial shearing event at the contact of differing lithologies due to contrasting competencies. This type of structure accounts for the Lower Silicified Zone at the Margaret Extension and Margaret South and occur as a series of sub-parallel en-echelon mineralized lenses striking 050° to 070°. The second type of mineralized structure is a transcurrent type of shearing which developed large scale, sub-vertical shears across the property. (Beilhartz, D. 2006)

1901 Deposit

The 1901 Deposit is hosted in a shallowly NNE dipping package of highly altered volcanic rocks of calc-alkaline geochemistry (Figure 7-5). The volcanic rocks consist of altered mafic, intermediate and felsic volcanics, similar to other VMS deposits in the Chisel-Lalor area. A major flat lying fault in the hanging wall of the 1901 deposit separate the deposit from a hanging wall package of unaltered, steep to moderately dipping southwest facing open folded tholeiitic rocks of the Threehouse and North Chisel formations. The 1901 deposit is hosted in highly altered intermediate "gneiss" that structurally overlies a sequence of felsic volcanic rocks (felsic "gneiss). The deposit occurs where the intermediate "gneiss" is thinned, and the 1901 VMS type mineralization is within 30m of the felsic "gneiss". The 1901 deposit is characterized by a large zone of metamorphically recrystallized and hydrothermally altered rocks that are composed of sericite- and chlorite-rich rocks with abundant porphyroblasts of kyanite-silimanite, biotite, staurolite, anthophyllite, cordierite and garnet. The primary textures are rarely preserved, and the protolith is no longer recognizable except through (immobile) trace element geochemistry.

FIGURE 7-5 1901 REPRESENTATIVE CROSS SECTION WITH DRILL HOLE TRACES

Pen II Deposit

The Pen II deposit is located in the Snow Lake arc assemblage (Figure 7-6). The volcanic stratigraphy that hosts the Pen II deposit is on the eastern margin of the younger Ham Lake intrusion. The stratigraphy strikes north - northeast and dips moderately to steeply west. The deposit has extensive granitic to gabbro dykes throughout the hanging wall and numerous felsic to mafic volcanic flows in the footwall of the deposit. The mineralization is generally hosted within felsic volcanic or volcaniclastics or at a contact between felsic and mafic volcanics. All the rocks have undergone lower to middle amphibolite facies metamorphism and primary volcanic features have been obliterated.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

FIGURE 7-6 PEN II REPRESENTATIVE CROSS SECTION WITH DRILL HOLE TRACES

Watts Deposit

The Watts River deposit is believed to have formed in an arc-rift setting. The deposit is a stratabound accumulation of sulphide minerals that were likely precipitated in association with contemporaneous volcanism. The massive sulphide zone was interpreted to be "distal" as no alteration pipe associated with the base metal mineralization has been identified to date. The rocks hosting the Watts River deposit are coarsely recrystallized to upper amphibolite to perhaps granulite facies mineral assemblages. They are overprinted by a strong gneissic fabric and intruded by numerous small veins of granitic (quartzo-feldspathic gneiss) material. The mineralization occurs in three main lenses that extend from the bedrock-surface contact to a depth of 900m below surface (Figure 7-7).

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

FIGURE 7-7 WATTS REPRESENTATIVE CROSS SECTION WITH DRILL HOLE TRACES

Wim Deposit

The Wim deposit is hosted within a northwest-striking, northeast-dipping package of mafic and felsic metavolcanic gneisses along the southern rim of the Herblet Lake Gneiss Dome Complex (Figure 7-8).

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FIGURE 7-8: WIM DEPOSIT CROSS SECTION WITH DRILL HOLE TRACES

The meta-volcanic gneisses are interpreted to be derived from the early Proterozoic subaqueous and subaerial volcanics of the Amisk Group that is locally overlain by marine turbidites of the Burntwood Group and terrestrial sediments of the Missi Group and intruded by various dykes. Tectonic foliations in the vicinity of the Wim deposit trend northwest and dip moderately to steeply towards the northeast.

Several alteration types associated with the sulphide mineralization have been documented in outcrop and drill core. Other alteration types encountered in outcrop and drill core include medium to coarse-grained, cordierite-anthophyllite, sieve textured garnets with chalcopyrite inclusions, muscovite-sillimanite and garnet zones. There is a general coarsening of grain size with proximity to mineralization due to the metamorphic recrystallization of hydrothermally altered rock and the growth of strongly porphyroblastic minerals such as garnet, anthophyllite, cordierite and sillimanite. Altered rhyolites occur to the northwest and southeast of the deposit along the geophysical trend that coincides with the Wim deposit mineralization.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

8. DEPOSIT TYPES

8.1. VMS DEPOSITS

The deposits of interest in this Technical Report are classified as VMS (volcanogenic massive sulphides). VMS deposits are an important source of copper, zinc, lead, gold and silver (Cu, Zn, Pb. Au and Ag) and form at or near the seafloor where circulating hydrothermal fluids driven by magmatic heat that have quenched through mixing with bottom waters or porewaters in near-seafloor lithologies (Figure 8-1).

FIGURE 8-1: CONCEPTUAL MODEL OF A VOLCANOGENIC MASSIVE SULFIDE (VMS) DEPOSIT

(Modified After S. Lavoie, 2003)

Massive sulfide lenses vary widely in shape and size and may be podlike or sheetlike. They are generally stratiform and may occur as multiple lenses from small pods of less than a ton (which are commonly scattered through prospective terrains) to giant accumulations like the Rio Tinto mine (Spain), Kholodrina (Russia), Windy Craggy (Canada), Brunswick No. 12 (Canada) or Ducktown (United States). Volcanogenic massive sulfide deposits are actively forming in extensional settings on the seafloor, especially mid-ocean ridges, island arcs, and back-arc spreading basins.

The location of VMS deposits is often controlled by synvolcanic faults and fissures, which permit a focused discharge of hydrothermal fluids. As illustrated on Figure 8 1, a typical VMS deposit will include the massive mineralization located proximal to the active hydrothermal vent, footwall stockwork mineralization, and distal products, which are typically thin but extensive. Footwall, and less commonly, hanging wall semi-conformable alteration zones are produced by high temperature water-rock interactions (Franklin et al., 2005).

Massive sulphide ore in VMS deposits consists of usually more than 40 percent sulfides, dominated by pyrite, pyrrhotite, chalcopyrite, sphalerite, and galena. Non-sulfide gangue typically consists of quartz, barite, anhydrite, iron (Fe) oxides, chlorite, sericite, talc, and their metamorphosed equivalents. Sulphide ore composition may vary from Pb-Zn- to Cu-Zn-, or Pb-Cu-Zn-dominated, and some deposits are zoned vertically and laterally. VMS deposits have stringer or feeder zones beneath the massive zone that consist of crosscutting veins and veinlets of sulfides in a matrix of pervasively altered host rock.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

Lalor Mine and 1901 Deposit

The depositional environment for the mineralization at Lalor and the 1901 deposit are similar to present and past producing base metal deposits in felsic to mafic volcanic and volcaniclastic rocks in the Snow Lake mining camp. The deposits appear to have an extensive associated regional hydrothermal alteration.

The mineralized envelopes are shallow dipping, with zinc mineralization defined to date beginning at approximately 600 m from surface and extending to a depth of approximately 1,100 m. The mineralization trends about N320° to N340° azimuth and dips between 30° and 45° to the north. It has a lateral extent of about 1,400 m in the north-south direction and 800 m in the east-west direction.

Sulphide mineralization is dominated by pyrite and sphalerite. In the near solid (semi-massive) to solid (massive) sulphide sections, pyrite occurs as fine to coarse grained crystals ranging one to six millimetres and averages two to three millimetres in size. Sphalerite occurs interstitial to the pyrite. A crude bedding or lamination is locally discernable between these two sulphide minerals. Near solid coarse-grained sphalerite zones occur locally as bands or boudins that strongly suggest that remobilization took place during metamorphism.

Disseminated blebs and stringers of pyrrhotite and chalcopyrite occur locally within the massive sulphides, adjacent to and generally in the footwall of the massive sulphides. The hydrothermally altered rocks in the footwall commonly contain some very low concentrations of sulphide minerals. Some sections of massive pyrrhotite occur, but these tend to give way to pyrite-sphalerite-dominant zones.

The top two lenses of the stacked base metal zones at Lalor (referred to as Zone 10 and 11) and lenses 1 and 3 at the 1901 deposit have higher grade zinc and iron content. Lenses located lower in the stratigraphy and coded as Zones 20, 30, 31, 32, 40 and 42 at Lalor have moderate to high zinc grades hosted in near solid sulphides containing higher grade gold and locally appreciable amounts of copper.

The footwall gold mineralization is typical of any VMS footwall feeder zone with copper-rich, disseminated and vein style mineralization overlain by a massive, zinc-rich lens. The fact that the footwall zone is strongly enriched in gold suggests a copper-gold association which is comparable to other gold enriched VMS camps and deposits (Mercier-Langevin, 2009). Some of the footwall zones tend to be associated with silicification and the presence of gahnite. These zones are often characterized by copper-gold association and are currently interpreted as being associated with higher temperature fluids below a zone of lower temperature base-metal accumulations.

Gold and silver enriched zones occur near the margins of the zinc rich sulphide lenses and as lenses in local silicified alteration. Remobilization is illustrated in some of the gold-rich zones by late veining that is restricted to the massive lenses.

Some of the footwall zones tend to be associated with silicification and the presence of gahnite. These zones are often characterized by copper-gold association and are currently interpreted as being associated with higher temperature fluids below a zone of lower temperature base-metal accumulations.

General observations of the known gold zones indicate areas which are coarse-grained and porphyroblastic in nature are gold poor, while fine grain siliceous (± veins ± sulphide traces) and strained looking stratigraphy tend to be gold rich. However, the intensity and style of alteration can vary strongly over short distances and may suggest that the alteration was forming discordant stockwork like zones that are now strongly transposed in the main foliation (Mercier-Langevin, 2009).

Watts Deposit

The Watts River deposit trends 020° azimuth and dips between 65° and 75° to the east. The sulphide intersections range from less than 0.3m up to 23m in core length, with a lateral extent of approximately 1,200m. Diamond drilling has intersected mineralization at depths of 850m below surface .Mineralization was intersected and interpreted as three lenses; Main Lens, Main Footwall Lens, and East Lens. The Main lens mineralization is comprised of coarse-grained pyrite, pyrrhotite, chalcopyrite, sphalerite, and minor galena. Mineralization was projected from the surface bedrock contact to a depth of 900m below surface as massive sulphides, fine disseminations and clots. The Main Lens has been divided into Main A, B, and C representing zones along the lens where appreciable copper and zinc grades and widths have been identified. The Main Footwall Lens lies in the immediate footwall of the Main Lens at depths of 600 to 880m below surface and tends to be slightly richer in sphalerite. The East Lens mineralization lies 400m into the hanging wall of the Main lens and extends from the surface bedrock contact to 280m below surface. Mineralization consists of narrow massive sulphides, disseminated, blobs, splashes, stringers of coarse-grained pyrite, pyrrhotite, chalcopyrite and lesser sphalerite. The massive sulphides exhibit a durchbeweget texture, characterized by the presence of rounded clasts of clear cherty quartz and amphibole inclusions within the sulphides. This texture suggests structural milling of the wall rock and the deposition of wall rock fragments within the recrystallized sulphides. The sulphides have generally been recrystallized to a coarse grain size, but sections of finer grained sulphides do occur.

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Pen II Deposit

The Pen II deposit comprises a stratabound, semi-massive to massive sulphide lens with an adjacent stringer/disseminated sulphide zone.

Mineralization is characterized by disseminated to massive, recrystallized and medium to coarse-grained sphalerite, pyrite, pyrrhotite and minor chalcopyrite. Based on the drilling completed by Hudbay and Callinan Mines Ltd. the mineralization extends from surface to 500 m below surface. The current strike length of the deposit is 400 m with an average thickness of 4 m. The deposit is conformable to stratigraphy, trends to the northeast at a N40º azimuth, a 45-65º dip towards the northwest.

Wim Deposit

The Wim deposit comprises a stratabound, semi-massive to massive sulphide lens with an adjacent stringer/disseminated sulphide zone. Mineralization is characterized by disseminated to massive, recrystallized and medium to coarse grained pyrite, pyrrhotite, chalcopyrite and minor sphalerite. Based on the drilling completed by Murgor, the VMS mineralization extends from surface to 720 m below surface. The current strike length of the deposit is 725 m with an average thickness of 10 m. The Wim deposit is conformable to stratigraphy, trends to the northwest at a N310º azimuth, a 40-45º dip towards the northeast and a plunge of 40º to the north. In addition, disseminated and stringer sulphides occur in the stratigraphic footwall of the Wim sulphide lens and may represent the feeder zone of the stratabound sulphide lens.

8.2. SNOW LAKE GOLD PROPERTIES

The Snow Lake Gold Properties (NBM, No. 3 Zone, Boundary Zone, Birch Zone and Squall Lake Area) are considered to belong to the quartz-carbonate vein gold subtype of orogenic lode gold deposits. This subtype of gold deposits consists of simple to complex quartz carbonate vein systems associated with brittle-ductile rock behaviour, corresponding to intermediate depths within the crust, and compressive tectonic settings (Figure 8-2). Deposits of this type have been commonly referred to as mesothermal gold quartz vein deposits, but they encompass both mesothermal and hypothermal classes as initially defined by Lindgren (1933).

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

FIGURE 8-2: SCHEMATIC DIAGRAM SHOWING THE SETTING AND NATURE OF OROGENIC LODE-GOLD DEPOSITS

Note: The plate tectonic environments of formation of orogenic lode-gold deposits and other gold-rich deposit styles, adapted from Goldfarb et al. (2005) and Groves et al. (2000). (B) Depth profile of orogenic lode-gold deposits, adapted from Groves et al. (1998)

At the regional scale, the quartz-carbonate vein gold deposits occur in two contrasting geological environments: deformed clastic sedimentary terranes and deformed volcano-plutonic terranes containing diverse volcanic assemblages of island-arc and oceanic affinities. Despite lithological and structural differences, these two types of environment share the following characteristics: greenschist to locally lower amphibolite metamorphic facies, brittle-ductile nature of deformation, and geological structures recording compressional to transpressional tectonic settings. Quartz-carbonate vein gold deposits in these environments tend to occur in clusters, or districts. Both types of environments are present in several districts, in which they are separated by major fault zones. However, in such cases auriferous quartz-carbonate veins preferentially occur in the volcano-plutonic domains.

At the New Britannia mine the mineralization is associated with shear zones, faults, and simple to complex vein systems. The auriferous zones (Toots, Dick, Ruttan and Hogg from west to east respectively) within the mine lie along a curvilinear shear (fault) zone named the Howe Sound fault also referred to as the Nor-Acme fault. One or more "slip planes" accompanied by a variety of altered mylonitic zones mark the location of the fault. The mylonitic zones (quartz-carbonate-mica) are predominantly less than 0.3 m thick but can thicken locally to between 3 m to 6 m. The gold bearing quartz-carbonate rocks are almost always situated next to or astride the fault. The main mineralized zone consists of quartz-carbonate alteration zones with arsenopyrite as the main sulphide accompanying the gold, emplaced in a simple intercalated sequence of altered felsic and mafic volcanic and pyroclastic. The contacts of the mineralized zones are often gradational; however, there have been sharp shear and fault contacts noted within some parts of the deposit¸ and a biotite alteration halo occurs within a few feet of the mineralized zone.

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

Mineralization in No. 3 Zone, Boundary and Birch Zones is hosted within sheared biotite-quartz and carbonate altered volcanic and volcaniclastic rocks that host quartz-albite-iron carbonate veins with fine to coarse gold, and is associated with acicular arsenopyrite within the veins, wallrock and wallrock fragments (Galley et al., 1991). Observations of the mineralization from underground at the No. 3 Zone noted that the quartz veins varied in true width from 0-1.5m and that the mineralization was not confined to the veins and can occur in up to 2 m from the main vein system.

Ten zones of gold-arsenopyrite mineralization have been identified in the Squall Lake area through previous exploration programs from 1943 to 1987. These zones occur within a 40-60 m stratigraphic interval above the staurolite schist unit and are aligned over a 10 km strike length along the northwest flank of the McLeod Lake Synform. These zones have been named K1, K5, K7, K10, K11, Margaret (North and South), Margaret Extension, South, F1 and Moon Gertie Zones. Two other similar zones to the South are the Bay and North Zones. The gold-arsenopyrite mineralization is directly associated with tabular quartz veins that are rimmed by fine-grained arsenopyrite, sub-parallel to the enclosing stratigraphy, or as tightly folded with a near vertical attitude. (Beilhartz, 2006).

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Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

9. EXPLORATION

9.1. LALOR MINE

In 2003, a Crone Geophysics high power time-domain electromagnetic (EM) system experimental survey was conducted over the deepest portion (approximately 600 m vertical depth) of the Chisel North Mine. The survey was designed and interpreted by Hudbay and was conducted by Koop Geotechnical Services Inc. The survey provided conclusive evidence that the system could detect conductive bodies at depths greater than 500 m. It was decided to extend the survey coverage further down-dip and down-plunge of the known mineralized lenses. A double-wired transmitter loop measuring 2 km by 2 km was used to maximize the EM field strength. The survey results were interpreted using three-dimensional (3D) computer modeling. The model indicated a highly conductive, shallow-dipping zone at a vertical depth of 800 m at Lalor.

Since the last NI 43-101 Technical Report in 2019, exploration drilling at Lalor has both focused on adding and converting inferred mineral resource estimates with a strong emphasis on confirming the continuity of the gold mineralization. Due to the low angle and stacking nature of the mineralization at Lalor, holes were extended beyond the base metal target depths to explore for feeder style mineralization rich in copper - gold, which led to increases in mineral resource inventory. Approximately 30,000 m of exploration underground drilling has been conducted per year at Lalor since 2019.

9.2. 1901 DEPOSIT

In the mid 1990's Airborne EM/MAG surveys (SPECTREM) were flown and highlighted numerous anomalies in the Chisel Basin, which were subsequently followed up with gridding, ground HLEM / magnetometer surveys and diamond drilling. Airborne VTEM geophysical surveys were also flown in 2007.

In January 2019, a winter drill program was initiated to follow-up untested DPEM and BHPEM targets. The initial hole CH1901, was testing and area known as the South Bull's eye and identified VMS type sulphide mineralization. After the completion of the 1901 discovery drillhole and additional 53 drillholes were completed in 2019 yielding a total of 35,138m of NQ core. Borehole EM was conducted on several drillholes and helped to direct the drilling. In 2020, a follow-up drill program was conducted on the 1901 deposit. A total of 39 NQ drillholes were completed for a total of 26,275m.

9.3. SNOW LAKE GOLD PROPERTIES

Exploration programs and exploration history on the Snow Lake Gold Properties (SLGP) are described in the previously disclosed NI 43-101 prepared for Alexis Minerals Corp. by Genivar, in 2010. A brief outline of the most recent exploration work on these properties is provided below.

In 2011, Bailes Geoscience was contracted by Alexis Minerals Corp. to perform a geochemical study of the lithological units in this area (McLeod Road-Birch Lake Thrust (MB)) sequence. The evaluation was conducted over 649 high-quality ICP-MS litho-geochemical samples collected from 35 exploration drill holes. The highlights of the results were:

The McLeod Road-Birch Lake volcanic rocks are a thrust repetition of the Snow Lake arc assemblage to the south.
Gabbro intrusions are geochemically identical to specific supracrustal units and therefore likely to be synvolcanic feeders for the rocks.
Rhyolite flows in the MB are all very similar in geochemistry and indistinguishable from one another. They display arc tholeiite type FII rhyolite geochemistry.
Alteration within the MB is very subtle compared to rocks associated with VMS. No gold-associated major element alteration trends were identified.

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In 2017, Hudbay re-initiated a systematic exploration program for lode gold deposits in the Snow Lake area and completed an Induced Polarization (IP) survey and an unmanned aerial vehicle (UAV) magnetic survey over the northern part of the (SLGP) encompassing the Birch, No. 3 Zone and the Boundary deposit areas.

9.4. PEN II DEPOSIT

The PEN II zone was discovered in 1986 by HBED. A ground HLEM and magnetometer grid in the area had been extended to the northeast of claim CB146 and identified a previously unknown EM anomaly. The anomaly was drilled and intersected favourable mineralization, so the drilling campaign was continued for a total of 15 holes. This drilling campaign marked the discovery of what is known as the PEN II zone. Follow-up drilling was conducted by HBED between 1987 and 1991.

Between 1999 and 2000, Callinex optioned the property and drilled an additional three holes and two wedges and completed an airborne EM survey over the zone.

In 2007, HBED drilled 6 holes to determine the geological continuity of the PEN II zone and tested geophysical targets in the area. In 2017, Hudbay (previously HBED) completed an additional seventeen drillholes on the property for a total of 3,120 m.

In 2018, 23 holes were drilled, totaling 6,374 m to upgrade the majority of the mineral resource estimates at the Pen II deposit to an indicated category. During this program, two new small high-grade satellite zones were discovered within the top 400 m.

In 2019, the drilling at Pen II focused on confirming the continuity of the mineralization in the two satellite zones, helped to define the exact extent of the mineralization towards surface, and were used to collect geotechnical data and samples for metallurgical sampling.

9.5. WIM DEPOSIT

Exploration programs and exploration history at Wim is described in the previously disclosed NI 43 101 prepared for Alexandria resources by Golder, in 2015. Since acquiring the property at the end of 2018, Hudbay has initiated a drilling program to collect geotechnical and metallurgical samples in order to test the flowsheet of the future New Britannia mill. Results have confirmed that the mineralization at Wim could attain processing recoveries in copper, silver and gold similar to the ones estimated for the Lalor copper-gold zones.

9.6. WATTS

The property was first staked in 1995 by HBED to cover anomalies located by a 1993 SPECTREM airborne EM survey. Diamond drilling of a horizontal loop electromagnetic (HLEM) anomaly in 1995 by HBED resulted in the discovery of the Watts River Deposit.

A total of 19 holes from the early 1995 to1997 drilling were borehole pulse EM surveyed. The surveys determined that the region of highest conductivity appeared to be in the vicinity of drill holes KUS223, 229, and 283, which lie within the Main Lens A. Five of the seven holes drilled during the 2000 to 2001 campaign were borehole pulse EM surveyed, defining mineralization in the East Lens as predominantly disseminated and minor massive sulphides.

In 2000 - 2001 the property was optioned to Fort Knox - Berland Resources, which completed eight (8) holes focusing on the newly discovered east lens which is approximately 400m in the hangingwall of the main deposit.

Four of the five holes encountered an in-hole response and these lie within the East Lens. The borehole EM surveys during the 2005 to 2007 drilling did not reveal any off-hole anomalies.

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In 2019, Hudbay completed a diamond drill program on the Watts River deposit. Six diamond drillholes were completed for a total of 2,673m. The focus of the program was to look at possible expansion of the known mineralization.

9.7. REGION EXPLORATION

Hudbay has an extensive land package in the Snow Lake belt of 17,517 Ha that provides significant additional upside for further gold and base metals exploration (Figure 9-1).

FIGURE 9-1: MINERAL OCCURRENCES IN THE SNOW LAKE CAMP

Since 2018, Hudbay spent close to C$10 million annually on major airborne and ground geophysical surveys as well as on surface exploration drilling in the Snow Lake area. This work was instrumental in identifying several base metal and gold targets that are to be followed up in the near future with comparable annual exploration budgets.

It is the QP's opinion, that sampling to date is both representative and adequate to develop a sound interpretation of the geology and the grade distribution for all the deposits covered by this Technical Report.

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10. DRILLING

There are no known factors that could materially impact the accuracy and reliability of the drill results reported in this section.

10.1. LALOR MINE

Surface Drilling

As of January 1, 2021, a total of 305,487 m of surface drilling was completed at Lalor. Table 10-1 provides a summary by year.

TABLE 10-1: SURFACE DRILLING PROGRAMS BY YEAR (IN METERS DRILLED)

All surface holes have been surveyed using reflex downhole tools (Ez-Shot or Ez-Trac) systems at a spacing of 30 metres down the hole. The survey instruments measure magnetic azimuth, dip, temperature, earth's magnetic field and roll face of tool for wedging operations. This data is then copied onto reflex survey instruments by the drillers and provide to the geologist daily. Readings are then inputted into the drilling database and a correction is applied to the magnetic azimuth to calculate true azimuth based on yearly diurnal drift of the magnetic north pole. All down hole survey results are validated by a geologist before being entered into the drill hole database.

Underground Drilling

Underground drill holes are drilled at all dips and azimuths needed to provide adequate coverage of the orebody for interpretation and mining purposes (

Table 10-2). Holes with dips steeper than +70° are preferably avoided due to poor ergonomics and the increased risk for the drill crews.

Underground drilling at Lalor can be divided into four different categories based on the primary planned purpose of the hole:

Exploration holes for targets outside known lenses or in areas of low drill density.
Definition holes are drilled into known lenses to upgrade inferred resources to a higher category and to identify mineralization contacts for preliminary mine design purposes. Typical hole spacing is about 15 to 20 m.
Delineation holes within indicated resources establish exact ore contacts for detailed mine planning and stope design purposes. Typical hole spacing is 10 to 15 m.

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Engineering holes are drilled for mine infrastructure purposes such as drain holes, holes for electrical cables and service holes for break through rounds.

TABLE 10-2: UNDERGROUND DRILLING PROGRAMS BY YEAR (IN METERS DRILLED)

AS shown in Table 10-3, 4,042 drill holed totaling approximately 641,000 m were included in the Lalor database to support the mineral resource estimate. Amongst these samples approximately 397,300 m were analyzed for Zn, Cu, Au and Ag and density was measured for 84,142 samples.

TABLE 10-3: DRILLING DATA BY YEAR

The drill hole database was directly imported in MineSight® from acQuire® with a cut-off date for mineral resource estimate purposes of September 8th, 2020. In parallel, the drill database was also imported in Leapfrog® for additional and independent comparison and validation. No significant differences in the projection of the drill hole traces were found between the two datasets. Additional validation conducted in the drill hole database are described in section 12.

All underground surveying at Lalor mine is conducted in Universal Transverse Mercator projection using NAD83 datum Zone 14. The mine survey was tied into the Canadian Spatial Reference System grid through the use of the Global Navigation Satellite System (GNSS) and an initial reference point at surface. Using the initial reference point, a reference azimuth was established by the use of a gyro compass. Based on the initial GNSS point and gyro azimuth a traverse was conducted to carry the survey down the ramp to the 810 m level. Once the main underground development workings were completed, another traverse was done to bring the survey down to 955 m level. All surveys into the individual levels are based on the control points created by the initial survey located at 810 m level and 955 m level. All surveying is done using electronic theodolites using the resection/free station technique.

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Diamond drill lines are marked up according to layouts issued by the Mine's Geology Department. An electronic version of the layout is created for viewing on the electronic theodolite. In the field, the electronic theodolite is set up using resection or free station technique. Reflectorless Electronic Distance Measurement (EDM) is used to locate the planned drill azimuth line location according to the electronic drill hole layout. Collar location and a drill azimuth back sight is spray painted on the walls of the drift to function as front and back sights for diamond drillers to line up the drill. An anchor is installed on the painted lines to provide a permanent reference line. Once drilling is completed the collar survey is recorded using the resection setup method. Collar location is surveyed using reflectorless EDM surveying and collar location stored as single point data.

Downhole surveys were completed using a Reflex EZ-Shot®, EZ-A® or EZ-Trac® (Reflex) instrument. Surveys were completed at regular intervals of 30 m down the hole.

A down hole gyro survey was conducted at the Lalor Mine in June and July of 2016 to check the accuracy of the magnetic Reflex readings. The survey was conducted by company technicians using a Reflex TN14 gyro compass and a MEMS down hole gyro probe. Thirty-eight underground holes were surveyed by gyro method for a total of 11,358 m. No significant discrepancies were identified between azimuth values measured by magnetic Reflex survey methods and azimuth values measured by the gyro survey methods.

10.2. 1901 DEPOSIT

In January 2019, a winter drill program was initiated to follow-up untested DPEM and BHPEM targets. The initial hole CH1901, was testing and area known as the South Bull's eye and identified VMS type mineralization. After the completion of the 1901 discovery drillhole an additional 53 drillholes were drilled in the deposit area in 2019. An additional 39 drill holes were completed in 2020 for a total of 61,413m of NQ drill core between 2019 and 2020 (Table 10-4). All drill hole collars have been surveyed with a differential global positioning system (DGPS) with centimeter accuracy.

TABLE 10-4: SUMMARY OF THE 1901 DEPOSIT SURFACE DIAMOND DRILLING

All drilling was carried out following CIM Exploration Best Practices. Each surface Exploration drill hole was lined up with a Reflex TN14 Gyrocompass, the drill hole was surveyed during drilling with a Reflex EZ-Trac at 30m intervals, and subsequently surveyed with a Reflex EZ-Gyro at the end of the hole. The Reflex data is downloaded directly from the IMDEXHUB-IQ into the acQuire database to ensure that there are no transcription errors.

The Reflex EZ-Trac and Gyro surveys were compared for several holes and no significant discrepancies were identified between azimuth values measured by magnetic Reflex survey methods and azimuth values measured by the gyro survey methods.

10.3. SNOW LAKE GOLD PROPERTIES

Table 10-5 lists the historical surface and underground drilling on the Snow Lake Gold Properties since the 1940s.

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TABLE 10-5: SUMMARY OF SURFACE AND UNDERGROUND DIAMOND DRILL HOLES (AS OF JANUARY 1, 2021)

Alexis Minerals Corp. was the last operator to conduct drilling on the property. The 2010-2011 Alexis drilling campaigns focused on collecting information in areas with the potential to provide mineral resources to the mining plan and additional surface targets in the Snow Lake Gold area. the drill programs concentrated on areas around the Boundary, Birch, Mine East and near the New Britannia Mine.

This most recent drilling completed by Garson Gold Corp. and Alexis Minerals Corp was NQ diameter core utilizing a drilling company based in Alida, Saskatchewan called Dig-It Exploration. The location of all the drill holes were surveyed and tied into the established grid on the property. Down hole surveys were completed using Fordia Flexit Smart-tool System. The first reading was taken approximately 10 m down the hole with subsequent readings being taken every 50 to 60 metres down the hole.

Orientated core has been performed on some of the holes drilled at No. 3 Zone and at Boundary using the Ball Mark Orientation System. The orientated core was logged for structural measurements on veins and rock fabric.

In 2019, a drill program was completed in the 3 Zone deposit area. A winter program of 3 drill holes were completed for a total of 1,173 metres and a summer/fall metallurgical drilling program of 4 drill holes for a total of 1,160m (Table 10-6). Each drill hole was lined up with a Reflex TN14 Gyrocompass and the drill hole was surveyed with a Reflex EZ-Trac at 30m intervals downhole.

10.4. WATTS

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Four of the five holes encountered an in-hole response and these lie within the East Lens. The borehole EM surveys during the 2005 to 2007 drilling did not reveal any off-hole anomalies.

10.5. PEN II

Historical drilling at the Pen II deposit included 15 holes for a total of 2,029.26 metres, for which downhole surveys were completed sporadically using acid tests and tropari instrument. Since 2017, 9,531 metres of additional drilling have been collected with downhole survey utilizing reflex (Ez-Shot and Ez-Trac) systems. Readings were then inputted into the drilling database and a correction is applied to the magnetic azimuth to calculate true azimuth based on yearly diurnal drift of the magnetic north pole. All down hole survey results are validated by a geologist before being entered into the drill hole database. The drilling focused on improving the confidence of the mineralization between each drillhole. Hole collars were picked up utilizing a Differential Global Positioning System (DGPS) supplied by Hudbay's mine technical service group.

10.6. WIM

In total, 136 diamond drill holes and 43,285.36 m have been completed on the Wim deposit from 1962 to 2008. All drilling prior to Murgor's exploration campaigns on the Wim deposit was completed by HBED. HBED completed six drilling campaigns between 1962 and 1991, which included 75 diamond drill holes (AX, EX and BQ core sizes) for a total of 25,018.90 m of drilling of which 17 were wedge drill holes. In 2007 and 2008, Murgor completed a total of 61 NQ-sized diamond drill holes for a total of 18,266 m of drilling on the Wim deposit. Half core samples from the historical drilling programs (pre-2007) are stored at the Stall Lake Mill Complex located approximately 5 km from Snow Lake. All half core samples from Murgor's 2007 and 2008 drilling campaigns are permanently stored on the Wim project site or at the Manitoba government core storage at the past producing Centennial mine.

In winter 2019, a drill program of seven drill holes was completed on the WIM deposit. A total of 1,494.5 metres were drilled. The goal of the drill program was to complete several drill holes for metallurgical studies of the Wim deposit mineralization.

11. SAMPLE PREPARATION, ANALYSES & SECURITY

Sample preparation, analysis, and security procedures applied to Lalor and the Snow Lake satellite deposits were reviewed by Olivier Tavchandjian, P. Geo, the Qualified Person of this Technical Report. The sampling methodology, analyses and security measures have been documented in great detail in past published technical reports by Hudbay for its Manitoba operations and will only be summarized in this document. Generally, the same procedures were followed by the exploration team of Murgor Resources at Wim, who was under the supervision of a former chief geologist at Hudbay. Unless specified in the report, all the laboratories used for samples preparation, analysis and security are independent of Hudbay. The procedures followed at the New Britannia mine and other Snow Lake gold properties are summarized at the end of this section and are described in greater details in the Technical Report produced by Genivar for Alexis Resources in 2011 and available on SEDAR.

11.1. LALOR, 1901 & PEN II

Sample Preparation

At the Lalor mine, drill core is logged directly into the acQuire drill hole database using dedicated laptop computers. Lithologies, alteration, structures, mineralization core quality and recovery are described in detail throughout the core length. Sample intervals are clearly marked on the core.

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The core is photographed before samples are cut and bagged for shipment. Photographing the core is the last step of the logging process to assure exact locations of all contacts, sample locations and numbers are captured. A standard setup as well as a single camera type with standardized settings are used to ensure photographs are of consistent quality. Once photographs have been captured, they are saved on the company server and linked to the acQuire drill hole database for easy network access.

Prior to cutting the core, the database manager prints a sample list for each drill hole that includes the sample identification number, hole name, sample type and the start and end footage of each sample. The core is cut along the cut line drawn by the geologist. In gouge and rubble intervals, an aluminium sampling scoop is used to separate the gouge into two halves in the core boxes. Completed sample bags are closed using zip ties. All saws and sampling buckets are rinsed with water after cutting each sample to prevent cross-contamination.

Once logging is completed, all data from sample books including QAQC samples are entered into Hudbay's acQuire drill hole database by hole number. Before dispatch, QAQC samples are inserted into the sample stream. Hudbay's practice involves insertion of the following every 100 samples:

1) Two blanks

2) Five duplicates

3) Five base metal standards

4) Two gold standards

Sample preparation and analysis for Lalor has been conducted at three different laboratories over time. Since September 2016, nearly all samples are prepared and assayed at Bureau Veritas. All drill core samples analyzed since the last technical report (NI 43-101 published on March 30, 2019) have been sent for analysis at Bureau Veritas. SGS was used as the umpire laboratory for quality control purposes.

ACME/Bureau Veritas

All samples arriving at ACME (Bureau Veritas) were checked against the custody chain form and prepped according to the lab protocols. The sample preparation includes weighing the sample, crushing 1 kg to minimum 80% passing 2 mm and a 250 g split crushed to minimum 85% passing 75 µm.

Base metals and other elements were analyzed with aqua regia digestions coupled with ICP-ES/MS (method AQ270 and AQ370). Assays returning "over the range" values were re-assayed with higher reporting methods (GC820 for copper, GC816 for zinc, GC817 for lead and MA404 for copper and zinc when values were above 20%).

All samples were analyzed for gold using a fire assay method coupled with atomic absorption (FA430). A gravimetric finish (FA530) was used when samples return gold values above 10 PPM.

Bureau Veritas is currently registered with ISO 9001 and ISO/IEC 17025 accreditations. Fine sample pulps are kept in secure storage at the laboratory after analysis. Pulps are only released after all data is validated.

SGS

All samples arriving at SGS are checked against the custody chain form. No sample preparation was conducted at SGS since the laboratory was only used to check assays done on pulps.

Base metals and other elements were analyzed with aqua regia digestions coupled with ICP-ES (method ICP14B). Assays returning "over the range" values were re-assayed with higher reporting methods (ICP13B when values were above 1% for zinc, copper and lead).

All samples were analyzed for gold using a fire assay method coupled with atomic absorption (FAA313). A gravimetric finish (FAG303) was used when samples return gold values above 10 PPM.

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SGS is currently registered with ISO 9001 and ISO/IEC 17025 accreditations. Fine sample pulps are kept in secure storage at the laboratory after analysis. Pulps are only released after all data is validated.

Specific Gravity via gas pycnometer (Bureau Veritas, Vancouver)

A total of 16,559 density measurements were completed at Lalor since the last NI 43-101. At 1901, density measurements were performed on 13,664 samples. As for Pen II, a total of 1,306 density measurements were collected by Hudbay.

Quality Control and Quality Assurance

Data verification procedures and results since the previous technical report are included in this section based on information collected and reviewed from 2018 to 2020. During this period, all samples were assayed at the Acme/Bureau Veritas laboratory. For data verification procedures and results prior to 2018, refer to previous technical report dated March 30, 2019.

As part of Hudbay QAQC program, QAQC samples were systematically introduced in the sample stream to assess sub-sampling procedures, potential cross-contamination, precision and accuracy. Hudbay commonly includes 5% certified reference materials (CRM), 2% certified blanks, and 5% coarse duplicates. Blanks and CRMs were prepared by Ore Research and Exploration (OREAS) and Analytical Solutions Ltd. (ASL). High-grade gold standards are from Rocklabs. All QAQC samples were analyzed following the same analytical procedures as those used for the drill core samples.

Standards

A failure rate that is less than 5% is considered good, while a failure rate between 5% and 10% is considered reasonable.

At Lalor, from 2018 to 2020, a total of 8,663 OREAS and Rocklabs CRMs were analyzed at Bureau Veritas representing 5.4% of the sample stream. A failure rate that is greater than 10% would trigger additional investigations. In most cases, the analytical bias is considered adequate for resource estimate purposes. High positive bias for silver in a low grade standard A6 is attributed to the certified value of the CRM being close to detection limit of the assay method used.

At 1901, between 2019 to 2020 a total of 797 OREAS standards were analyzed at Bureau Veritas, representing 5.8% of the sample stream. The analytical bias for all metals at 1901 is considered more than adequate for resource estimate purposes. Some failures in low grade standards A6 & B6 were noted for zinc and copper. Failures were corrected by re-analysing part of the assays batch, including the standard, until all the QAQC results cleared the built-in checks in the acquire database.

At Pen II, between 2017 to 2018 a total of 77 OREAS were analyzed at Bureau Veritas, representing 4.9% of the sample stream. In most cases, the analytical bias on zinc is considered adequate for resource estimate purposes. High failure rates mostly impact copper and silver for standard values which are rarely found at Pen II. In any case, when failure of a CRM occurred, part of the assays batch including the standard was re-assayed until all the QAQC results cleared the built-in checks in the acquire database.

Blanks

A failure rate that is less than three times the detection limit (3xDL) 90% of the time is considered acceptable. A blank failure either equates to samples swap or cross contamination unless the expected value is close to the reporting limit. In any case, if a blank fails, as part of Hudbay's QAQC protocols, the full batch is re-assayed.

At Lalor, a total of 3,466 certified OREAS and ASL blanks (about 2.2% of the samples stream), were inserted into the sample stream to monitor potential cross-contamination or sample swaps at Bureau Veritas. In all cases aside from copper in the blank F7 analysed at Bureau Veritas, the cross-contamination or sample swaps are considered adequate for resource estimate purposes. The failure of copper in F7 is linked to its certified value being equal to the 3xDL criteria. In this case, 98% of the results came back within ten times the detection limit (0.01%), which is a more appropriate performance gate, given its certified value. In any case, when failure of a blank CRM occurred, part of the assays batch, including the blank, was re-assayed until all the QAQC results cleared the database acQuire built-in checks.

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At 1901, a total of 806 blanks, sourced from a certified ASL blank (F8) and a non-certified internal geological blank (GQC-granite) that has been historically used by Hudbay, were inserted into the sample stream to monitor potential cross-contamination or sample swaps at Bureau Veritas. The level of cross-contamination and/or sample swaps are very low and/or non-existent to consider data collected during the 2019 - 2020 drilling campaign and no significant issues were observed; however, a significant number of samples returned zinc, copper, and silver results that are higher than 3x detection limits. These failures are attributed to certified value of blanks used being significantly higher than the detection limits of zinc, copper, and silver in the ultra-trace AQ251-EXT method used in exploration projects. The maximum reported results for zinc, copper, and silver in the blanks are 0.02%,0.006%, & 0.2 g/t, suggesting least cross-contamination and no sample swaps occurred at 1901.

At Pen II, a total of 82 certified OREAS and ASL blanks, or 5.3% of the samples stream, were inserted into the sample stream to monitor potential cross-contamination or sample swaps at Bureau Veritas. In all cases aside from copper in the blank F7, the cross-contamination or sample swaps are considered adequate for resource estimate purposes. The failure of copper in F7 is linked to its certified value been equal to the 3xDL criteria. In this case, 100% of the results came back within five times the detection limit, which is a more appropriate performance gate given its certified value. In any case, when failure of a blank CRM occurred, part of the assays batch including the blank was re-assayed until all the QAQC results cleared the acQuire database built-in check checks.

Coarse Duplicates

At Lalor, a total of 8,540 coarse duplicates were analyzed at Bureau Veritas representing 5.3% of the sample stream. Based on the paired analysis results, the laboratory appeared to be precise.

At 1901, a total of 804 coarse duplicates were analyzed at Bureau Veritas representing 5.9% of the sample stream. Based on the paired analysis results, the laboratory appeared to be precise.

At Pen II, a total of 43 coarse duplicates were analyzed at Bureau Veritas representing 2.7% of the sample stream. Based on the paired analysis results, the laboratory appeared to be precise.

Umpire Laboratory

At Lalor, a total of 8,273 pulp rejects were sent to and analyzed at SGS to validate the reproducibility of Bureau Veritas results. These external check samples represent 5.2% of the sample stream. Based on the paired analysis results, assay data from Bureau Veritas appeared to be reproducible

At 1901, a total of 787 pulp rejects were sent to and analyzed at SGS to validate the reproducibility of Bureau Veritas results. This external check represents 5.8% of the sample stream. Based on the paired analysis results, the Bureau Veritas results appeared to be reproducible.

At Pen II, a total of 67 pulps rejects were sent to and analyzed at SGS to validate the reproducibility of Bureau Veritas results. This external check represents 4.3% of the sample stream. Based on the paired analysis results, the Bureau Veritas results appeared to be reproducible.

11.2. WATTS DEPOSIT

Sample Preparation

The sample preparation methods performed on samples collected at the Watts deposit during the 2019 drill program are the same as the ones used for Lalor mine and the 1901 deposit. Table 11-1 presents the number of samples prepared and analyzed by year.

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TABLE 11-1: WATTS SAMPLES PREPARATION AND ANALYSIS BY YEAR

Specific Gravity

Density measurements were performed for 699 samples at Hudbay's logging facility for samples collected between 1995 and 2007, and Bureau Veritas laboratory in Vancouver for samples collected in 2019. Density measurements were completed using a non-wax-sealed immersion technique to measure the weight of each sample in air and in water for the 1995 - 2007 samples. Density measurements at Bureau Veritas were completed using a gas pycnometer (method SPG04). Table 11-2 presents the number of density measurements by year

TABLE 11-2: DENSITY MEASUREMENTS AT WATTS BY YEAR

Sample Analysis

The laboratory analysis for samples collected at the Watts deposit between 1995 and 2007 were Hudbay's internal laboratory in Flin Flon and TSL laboratory in Saskatoon. All samples collected during the 2019 program were analysed at the Bureau Veritas laboratory in Vancouver with SGS laboratory conducting external checks.

Hudbay's Flin Flon Laboratory

Base metals and silver analysis are completed by aqua-regia digestion and read by inductively coupled plasma (ICP-AES). Gold analysis are completed on each sample by atomic absorption spectrometry (AAS) after fire assay lead collection. Gravimetric method is used for gold if the range is above the upper limit (7 PPM) of the AAS.

Bureau Veritas

The methods used by Bureau Veritas on samples collected at the Watts deposit during the 2019 drill program are the same as the ones used for Lalor, 1901 and Pen II deposits.

SGS

The methods used by SGS on pulp rejects sent by Bureau Veritas for samples collected at the Watts deposit during the 2019 drill program are the same as Lalor, 1901 and Pen II deposits.

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Quality Assurance and Quality Control

Data verification procedures and results are based on information collected and reviewed from the 1995 - 2007, and 2019 drilling campaigns. QAQC samples, i.e., standards, blanks, and duplicates were introduced into the sampling stream for samples collected between 2005 and 2019. No QAQC samples were taken for samples collected between 1995 and 2001.

As part of Hudbay's QAQC program, for samples collected from 2005 to 2019, QAQC samples were systematically introduced into the sample stream to assess sub-sampling procedures, potential cross-contamination, precision and accuracy. Hudbay commonly includes 5% certified reference materials (CRM), 2% certified blanks, and 5% coarse duplicates in the sample stream. Blanks and CRMs were prepared by Ore Research and Exploration (OREAS) and Analytical Solutions Ltd. (ASL). All QAQC samples were analyzed following the same analytical procedures as those used for the drill core samples.

Standards

A total of 73 standards that represent 2.7% of total samples taken at Watts between 1995 and 2019 were introduced into the sampling stream between 2005 and 2019. A total of 53 OREAS standards A4, B4, C4, D4, & E4 were analysed at Hudbay's Flin Flon Laboratory in the 2005 - 2007 program, representing 5.5% of total samples taken during that time frame. In 2019, a total of 20 OREAS standards, A6, B6, C6, & D6 were analyzed at Bureau Veritas, representing 5.8% of the sample stream.

The analytical bias for all metals at Watts is considered more than adequate to validate the accuracy of the laboratory assay results. Quality control failures related to analytical results coming below lower limit in low to medium grade standards were noted for zinc and copper in the 2005 - 2007 program. Failures were corrected by re-analysing part of the assays batch, including the standard, until all the QAQC results cleared the built-in checks in the acquire database.

Blanks

A total of 78 blanks, sourced from a certified ASL blank (F8) and a non-certified internal geological blank (GQC-granite) that has been historically used by Hudbay, were inserted into the sample stream to monitor potential cross-contamination or sample swaps at Hudbay's Flin Flon lab and Bureau Veritas.

The level of cross-contamination and/or sample swaps are very low and/or non-existent to consider data collected during the 2005 - 2007 and 2019 drilling campaigns confirming the laboratory processed the samples in an adequate fashion; however, significant number of samples that returned zinc, copper, and silver results higher than 3 x detection limits were noted in the 2019 program. These failures are attributed to certified value of blanks used being significantly higher than the detection limits of zinc, copper, and silver in the ultra-trace AQ251-EXT method used in exploration projects. The maximum reported results for zinc, copper, and silver in the blanks are 0.005%, 0.006%, & 0.2 g/t, suggesting least cross-contamination and no sample swaps occurred at Watts.

Coarse Duplicates

Twenty coarse duplicates were analyzed from the 2019 program at Bureau Veritas representing 5.8% of the sample stream collected in 2019. Basic statistics, scatter plots, quantile-quantile plots, probability plots and absolute difference plots were used to assess the precision of the laboratory. Based on the paired analysis results, the laboratory appeared to be precise.

Umpire Laboratory

Twenty pulp rejects were sent to and analyzed at SGS to validate the reproducibility of Bureau Veritas results in the 2019 program. This external check represents 5.8% of the sample stream collected in 2019. Basic statistics, scatter plots, quantile-quantile plots, probability plots and absolute difference plots were used to assess the precision of the laboratory. Based on the paired analysis results, the Bureau Veritas results appeared to be reproducible.

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11.3. WIM

Hudbay has not performed any exploration work on the Wim property since 1991. The sampling methodology, analyses and security measures used by the previous owners (i.e. Murgor Resources and Alexandria Minerals) have been documented in Alexandria Minerals' 2015 technical report and will only be summarised in this document.

Sample Preparation

The samples preparation methods performed on the Wim samples are the same as the ones used for the Lalor mine samples. The logs were entered in a Gemcom Access® database.

Specific Gravity

572 density measurements were collected by Murgor Resources in 2007 and 2008. These measurements were performed at TSL Laboratories in Saskatoon, using a non-wax-sealed immersion technique to measure the weight of each sample in air and in water.

Sample Analysis

"All drill cores assayed from the 2007 and 2008 drilling programs were completed by TSL Laboratories, using multi-acid digestion and atomic absorption spectrometry for the base metals and fire assay, atomic absorption and gravimetric finish for gold. All samples selected for analysis were prepared and analyzed for Cu, Zn, Au, Ag, Fe, As and Pb.". TSL Laboratories is currently registered with ISO/IEC 17025 accreditation

Quality Assurance and Quality Control

Data verification procedures and results are based on information collected and reviewed from the 2007 and 2008 drilling campaign. During this period, all samples were assayed at the TSL laboratory in Saskatoon while whole rock geochemistry pulp samples were sent to ALS Laboratory in Vancouver. As part of Murgor's QAQC program, QAQC samples were systematically introduced in the sample stream to assess sub-sampling procedures, potential cross-contamination, precision, and accuracy, typically using a 10% insertion rate (i.e. one CRM, blank or duplicate every 10 samples). All QAQC samples were analyzed following the same analytical procedures as those used for the drill core samples at Wim.

The results of the review done by Golder indicated that Murgor's QAQC standard sampling program was working effectively, and the program met industry standards.

11.4. SNOW LAKE GOLD PROPERTIES

The sampling methodology, analyses and security measures used by the previous owners (i.e. Kinross and Alexis Resources) have been documented in the Technical Report produced by Genivar for Alexis Resources in 2011, available on SEDAR and will only be summarized in this document.

Limited confirmatory drilling was conducted by Hudbay in 2019 on the 3 Zone deposit. The sample preparation methods used for the samples collected during this confirmatory campaign are the same as the ones used for the Lalor mine, 1901 and Watts deposits.

Sample Preparation

Details of the sampling methods conducted by earlier operators of the New Britannia mine prior to 1995 have been lost. Between 1995 and 2003, the New Britannia sampling method consisted in core being logged on paper at the mine site by geologists. The written logs were later transferred into a Gemcom Access database. Technicians would lay out the core, split and sample the core according to the geologists' instructions. Chip samples were also collected during the mining operation by the geologist either taken from the back of the drift if the muck pile had not been removed or taken 1.40 m to 1.52 m above the floor of the drift if the blast muck has been removed. The geologist was responsible for bringing the samples to the preparation area of the mill laboratory and recording the number of each sample into a record book.

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In the 2019 drill program at the 3 Zone deposit, drill core is logged directly into the acQuire drill hole database using dedicated laptop computers. Lithologies, alteration, structures, mineralization core quality and recovery are described in detail throughout the core length. Sample intervals are clearly marked on the core.

The core is photographed before samples are cut and bagged for shipment. Photographing the core is the last step of the logging process to assure exact locations of all contacts, sample locations and numbers are captured. A standard setup as well as a single camera type with standardized settings is used to ensure photographs are of consistent quality. Once photographs have been captured, they are saved on the company server and linked to the acQuire drill hole database for easy network access.

Prior to cutting the core, the database manager prints a sample list for each drill hole that includes the sample identification number, hole name, sample type and the start and end footage of each sample. The core is cut along the cut line drawn by the geologist. In gouge and rubble intervals, an aluminium sampling scoop is used to separate the gouge into two halves in the core boxes. Completed sample bags are closed and secured at the neck using two zip ties. All saws and sampling buckets are rinsed with water after cutting each sample to prevent cross-contamination.

1) Five blanks

2) Five duplicates

3) Five standards (gold + base metals)

Sample Analysis

Most of the drill cores and chips assays from 1995 to 2003 from the New Britannia mine were completed at the on-site mill laboratory using a fire assay/atomic absorption finish (FA/AA) method.

At the other Snow Lake properties, sample analysis, including determination of gold and arsenic on all samples and specific gravity and multi-element geochemistry on selected samples, was completed at TSL Laboratories Inc. (TSL) in Saskatoon, SK. TSL was established in 1981 and certified as ISO/IEC 17025. At TSL, the samples are dried and crushed and a 250 g riffle split of the crushed sample is pulverized to 95% passing 150 mesh (106 microns). Gold concentration is determined on all samples from a 30-g split of the pulverized sample by fire assay with atomic absorption finish. All samples with a gold concentration greater than 1,000 ppb (1 gram/tonne) were re-assayed using fire assay with a gravimetric finish. Specific gravity and ICP-MS multi-element analyses were completed on selected samples as requested by Garson project geologists.

The laboratory analysis for the 2019 drill campaign on the 3 Zone deposit was conducted at the Bureau Veritas Laboratory in Vancouver with SGS laboratory conducting external checks.

Bureau Veritas

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All samples were analyzed for gold using a fire assay method coupled with atomic absorption (FA430). A gravimetric finish (FA530) was used when samples return gold values above 10 PPM.

Base metals and other elements were analyzed using the Ultra-trace by ICP-MS method (AQ251-EXT). Assays returning "over the range" values were re-assayed with higher reporting methods (AR404 for silver, copper, lead and zinc). If the range is above the upper limits of AR404, additional packages (i.e. GC816, GC817, and GC820) were used for zinc, lead, and copper respectively.

SGS

All samples arriving at SGS are checked against the custody chain form. No sample preparation was conducted at SGS since the laboratory was only used to check assays done on pulps.

All samples were analyzed for gold using a fire assay method coupled with atomic absorption (GE FAA313/GE FAA30V5). A gravimetric finish (GO FAG303/GO FAG30V) was used when samples return gold values above 10 PPM.

Base metals and other elements were analyzed with aqua regia digestions coupled with ICP-ES/MS (method GE ICP14B/GE ICM21B20). Assays returning "over the range" values were re-assayed with higher reporting methods (GO ICP13B/GO ICP21B100 when values were above 1% for zinc, copper and lead).

Specific Gravity

Density measurements were performed for 827 samples from the 3Zone deposit at Bureau Veritas laboratory in Vancouver using a gas pycnometer (method SPG04) in 2019.

Quality Assurance and Quality Control

Prior to 2014, standard, blank and duplicate assay samples were added to each batch of 21 samples for drill core and to each batch of 24 samples for chip samples.

The sampling methods and approaches and concluded that the sampling and analytical procedures conformed to the industry standards at the time, and these were adequate to ensure a representative determination for the type of gold mineralization identified on the property. The QAQC data documented in the Genivar Technical Report (2010) is not available for the QP to review. The QP is confident that the audits done by Micon in 2005 and 2009 are sufficient for the purpose of reporting inferred mineral resource estimates for the Snow Lake gold properties. The QAQC, data verification and results from the 2019 confirmatory drill program at 3 Zone were collected and reviewed and no issues was observed. During this period, all samples were assayed at the Bureau Veritas laboratory in Vancouver, British Columbia and checks conduced at SGS laboratory.

Standards

In 2019, a total of 41 OREAS standards were analyzed at Bureau Veritas, representing 5% of the sample stream. The analytical bias for gold at 3 Zone is considered more than adequate for resource estimate purposes.

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Blanks

In 2019, a total of 50 blanks non-certified internal geological blanks (GQC-granite) that has been historically used by Hudbay, were inserted into the sample stream to monitor potential cross-contamination or sample swaps at Bureau Veritas. The level of cross-contamination and/or sample swaps are very low and/or non-existent in the data collected during the 2019 drilling campaign. No significant issues were observed; however, a single blank returned a value greater than 5X the detection limit. The blank result was investigated and determined to be a result of contamination in the processing stage from a sample with visible gold (VG).

Coarse Duplicates

A total of 47 coarse duplicates were analyzed at Bureau Veritas representing 5.6% of the sample stream. Basic statistics, scatter plots, quantile-quantile plots, probability plots and absolute difference plots were used to assess the precision of the laboratory. Based on the paired analysis results, the laboratory appeared to be precise.

Umpire Laboratory

A total of 46 pulp rejects were sent to and analyzed at SGS to validate the reproducibility of Bureau Veritas results. This external check represents 5.5% of the sample stream. Basic statistics, scatter plots, quantile-quantile plots, probability plots and absolute difference plots were used to assess the precision of the laboratory. Based on the paired analysis results, the Bureau Veritas results appeared to be reproducible.

Based on the technical review of samples Preparation, Analysis and Security, the QP is satisfied that both the accuracy and precision of the laboratories used met the industry norms and formed a good basis upon which the mineral resource estimates can be reported.

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12. DATA VERIFICATION

12.1. LALOR MINE, 1901 WATTS, 3 ZONE, WIM AND PEN II DEPOSITS

Site Visits

Frequent site visits have been conducted by the Qualified Person since September 2017. These visits included the core logging facilities and the underground mine to review the geology and mineralization of the different mineralized zones. Visits of the exploration and the mine technical services offices to meet with the geologists were also part of the process. The Flin Flon Hudbay's laboratory and the Bureau Veritas laboratory in Vancouver were also inspected by the Qualified Person.

Database Management

As described in the previous sections of this Technical Report (10 and 11), drilling, logging, sampling data and assays results are captured and stored in an acQuire database. The acQuire platform provides robust data security and long-term data storage solutions. The assay database is administered by the database manager with working copies kept on the local drive of a secure computer and backups placed on a secure location on a Hudbay server. Any requests for edits to the database are made to the database manager who updates all the copies. All .pdf copies of the assay certificates and logs are available on the Hudbay's internal SharePoint website with restricted access.

Database Review

The drill holes samples used for the mineral Resource estimate included Zn, Cu, Au, Ag, Fe, Pb, As, Ni, and bulk density (SG). The drill database was directly imported into MineSight. Under the direct supervision of the Qualified Person, reviews of the drill holes to look for obvious errors were performed along with random checks of the acQuire database to ensure honouring of the original data such as the collar location, the downhole survey data and the assays results. Minor errors in the downhole surveys data of a few holes were detected and immediately corrected. It is the QP's opinion that the data is adequate and acceptable for use in the estimation of mineral resources.

12.2. SNOW LAKE GOLD PROPERTIES

The data verification used by the previous owners are summarized in this section, but more details can be found in the report from Technical Report produced by Genivar for Alexis Resources in 2011 and available on SEDAR.

Site Visits

Todd McCracken, P. Geo, for WSP, conducted a site visit in November 2018. During the site visit, the following validation tasks were completed:

Transfer of all available digital records to WSP;
Review of the state of geological/mineralization knowledge;
Review of the drilling quality assurance/control (QAQC) procedures;
Review of survey techniques for collar positions and down-hole surveys;
Review of the core handling and sampling procedures;
Review of the SG/bulk density determination procedures;
Review of database integrity and storage procedures;
Inspection of the core logging and core storage facilities.

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Database Review

WSP reviewed the process completed by Micon (Lewis, 2009) to verify completeness and reliability of electronic database files; Micon found no errors. Consequently, it can be reasonably assumed that the database is relatively free of errors.

Based on the technical review of data collection, quality control procedures, and inspection of previous studies, the QP is satisfied that both the integrity and quality of the mineral resource database met the industry norms and formed a good basis upon which the inferred mineral resource estimates can be reported.

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13. MINERAL PROCESSING AND METALLURGICAL TESTING

It is believed that all metallurgical test work presented in this section was conducted on samples that are representative of the various types and styles of mineralization and as a whole, the Lalor mine, 1901 deposit, WIM and 3 Zone deposits.

13.1. STALL CONCENTRATOR

The Stall Concentrator located 16km from the Lalor mine was commissioned in 1979. Since then, it has processed ore from many of the Snow Lake area mines. With the introduction of Lalor ore to the Stall Concentrator, an increase in the copper and precious metal content warranted the refurbishment of the concentrator to enable the production of a copper and precious metal concentrate alongside the zinc concentrate.

The Stall concentrator began processing ore from Lalor in August 2012, initially producing only a zinc concentrate and by October 2012 a copper concentrate. As Lalor increase ore production, the mill underwent an initial expansion and by the summer of 2014 was capable of processing ore at a 2,800 tpd throughput rate. Continuous improvement and debottlenecking led by the operations team has lifted the processing capacity of the Stall concentrator to above 3,800 tpd and instantaneous processing rates above 200 tph.

Historical Metallurgical Performance

The Stall Concentrator metallurgical performance processing Lalor Base Metal Ore from October 2012 to December 2020 is shown in Table 13-1 and actual concentrate produced is shown in Table 13-2.

The Stall Concentrator currently is producing a copper concentrate grading 18 to 20% copper depending on the levels of lead and talc present in the feed ore and a zinc concentrate grading 49 to 51% zinc. Metal recovery to the copper concentrate ranges from 85 to 88% and 55 to 62% for copper and precious metals, respectively. Metal recovery to the zinc concentrate ranges from 90 to 93% depending on the levels of gahnite, a zinc oxide mineral that is not responsive to flotation, in the feed ore and the amount of zinc recovered to the upstream processes.

TABLE 13-1: HISTORICAL STALL CONCENTRATOR METALLURGICAL PERFORMANCE

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TABLE 13-2: HISTORICAL STALL CONCENTRATOR CONCENTRATE PRODUCTION

Historical Metallurgical Testwork (2009 to 2018)

Historical metallurgical testwork results from the discovery of Lalor through to 2018 are outlined and can be viewed in the previous technical reports dated March 29th 2012, March 30th 2017 and March 28th 2019.

Metallurgical Testwork (2019 to 2020) - Lalor Base Metal Ore

In Q1 2019 a study program was begun to determine the recovery entitlement of the copper, gold, silver and gold within the Lalor Base Metal ore, the gap between plant performance and recovery entitlement and the main diluents in the final mineral concentrates. The study programs included:

Detailed flotation surveys (Blue Coast Research)
Detailed base metal mineralogy (ActLabs)
Detailed base metal and gold mineralogy (Advance Mineral Technology Laboratory, AMTEL)
Batch flotation tests on future ore composites (Blue Coast Research)
Jameson cell pilot plant testwork

In April 2019, Blue Coast Research with the assistance of site personnel conducted a multiple flotation circuit surveys. The flotation flowsheet and the material balance are outlined in Figure 13-1 and Table 13-3 respectively.

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FIGURE 13-1: STALL CONCENTRATOR FLOTATION FLOWSHEET APRIL 2019

TABLE 13-3: FLOTATION CIRCUIT SURVEY MASS BALANCE

In addition to the survey, Blue Coast Research conducted batch flotation tests on the cyclone overflow, copper regrind overflow and copper rougher tails to determine the recovery entitlement for the copper and zinc circuits. The flotation tests confirmed that the zinc circuit was operating at entitlement recovery, with the survey recording 93.58% zinc recovery (95.6% zinc circuit) against a laboratory results of 96.5% zinc circuit recovery. The flotation test also confirmed that the copper circuit was not operating at entitlement recovery. The laboratory tests returned a gold recovery entitlement of 67.7% (70.4% rougher recovery and 96.3% cleaner recovery) and a copper recovery entitlement of 92.3% (94.1% rougher recovery and 98.2% cleaner recovery). The absolute gap between the copper circuit recovery entitlement and the actual measured was 25.8% and 11.3% for gold and copper, respectively.

The recovery gap in the copper circuit formed the basis of the recovery improvement testwork program and all the mineralogy investigations focused on streams within the copper circuit.

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Mineralogical Investigation

AMTEL and ActLabs conducted detailed mineralogy focusing assessment on select samples from the copper circuit collected throughout the flotation survey campaign. The stream analyzed were primary cyclone overflow, rougher concentrate, rougher tails, cleaner 1 tails and cleaner 3 concentrate. Pyrite was the principal sulphide mineral in the cyclone overflow followed by sphalerite, pyrrhotite, chalcopyrite and galena. Chalcopyrite is the dominate copper bearing mineral present accounting for greater than 99% of the copper mineralisation.

The target mineral and the most abundance in the cleaner 3 concentrate is chalcopyrite. With principal diluent minerals being talc, galena, and sphalerite.

In the cyclone overflow, the chalcopyrite liberations are 89.4% with the remainder composited with sphalerite (2.83%) and pyrrhotite (2.84%). There are no measurable levels of association between galena and talc with chalcopyrite, indicating these minerals are recovered to the cleaner 3 concentrate via mechanical or chemical processes and not insufficient liberation.

Of the chalcopyrite lost to the rougher tails, 51.7% is in the free/liberated highlighting that most of the chalcopyrite that is not recovered to the copper concentrate is recoverable. A further 17.5% is associated with pyrrhotite is loss to the rougher tails and is potentially recoverable. This indicates that mechanical or chemical processes are the principal causes for the mineral loss and that the size reduction processes are operating as designed. Additional flotation capacity or the application of pneumatic flotation devices would be beneficial for the recovery of these liberation classes.

Amtel conducted additional mineralogy assessment on the primary cyclone overflow, rougher tails, cleaner 1 tails and cleaner 3 concentrate focusing on the gold deportment.

In the cyclone overflow 36.7% of the gold is present as free/liberated particles accounting for 0.91 g/t of the gold in feed. Of note is the 13.5% of gold in feed that is above 40µm in size and is amenable to gravity recovery processes. Fee/liberated gold accounts for 22.2% and 52.4% of the gold present in the rougher and cleaner 1 tails, respectively. This indicates that mechanical or chemical processes are the principal drives of recovery loss in this liberation class as these particles are highly responsive to flotation. Additional flotation capacity or the application of pneumatic flotation devices would be beneficial for the recovery of these liberation classes.

Sulphide association accounts for 29.3%, 22.9% and 35.2% of the gold in the cyclone overflow, rougher tails and cleaner 1 tails, respectively. The recovery of these gold particles is dependant on which sulphide mineral they are associated with. Gold not associated chalcopyrite has a high probability of being rejected. Sub-microscopic does occur within this fraction and is a contributor to the loss of gold to tails. Of the 0.61 g/t gold in the rougher tails, 0.17 g/t or 28% of the gold in the associated sulphide classes is sub-microscopic gold is not recoverable without diluting the copper concentrate.

Jameson Cell Pilot Plant Testwork

Based on the mineralogical investigation, Glencore Technology was contacted to supply a L500 Jameson Cell pilot plant for onsite metallurgical testing. The pilot plant testing was conducted on three streams in the copper circuit and three stream in the zinc circuit to determine if this flotation technology was suitable. The streams tested were copper rougher feed, copper rougher concentrate (post regrind), copper cleaner 1 tail, zinc rougher feed, zinc rougher concentrate and zinc cleaner 1 concentrate. The results for the copper streams tested are outlined in Figure 13-2 and in Figure 13-3 for the zinc stream tested.

In the single pilot Jameson cell treating the copper rougher feed, copper and gold recoveries tracked a typical mass recovery curve, with +90% copper and +70% gold recoveries achieved at 15% mass recovery indicating the Jameson cell is suitable for the rougher application. When treating the reground copper rougher concentrate or copper cleaner 1 concentrate, final concentrate grade (20% Cu) was achieved up to 60% recovery which is typical for Jameson Cell applications in base metal operations. This indicates that the Jameson cell is suitable for cleaning duties.

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FIGURE 13-2: JAMESON CELL PILOT TESTWORK RESULTS FOR THE COPPER CIRCUIT

When treating the zinc circuit streams it was evident that the sphalerite highly liberated and there were minimal competing minerals in the system. Final concentrate grade zinc concentrate (+50% Zn) was achieved with all stream tested with the zinc rougher feed achieving final concentrate grade at +80% recovery.

FIGURE 13-3: JAMESON PILOT TESTWORK RESULTS FOR THE ZINC CIRCUIT

Future Ore Testwork

Blue Coast Research conducted flotation testwork on nine composite samples created to simulate the future Lalor Base Metal ore from 2024 to 2030. The testwork was focused on understanding the flotation response in the ores to be treated on recovery of copper, gold, lead, and zinc. The results are outlined in Figure 13-4. The recovery of zinc in the zinc circuit is consistent with past performance. The recovery of copper and gold in the copper circuit highlights that recoveries greater than historical performances are achievable with 90 to 95% copper and 60 to 80% gold recovery possible.

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FIGURE 13-4: HEAD GRADE RECOVERY RELATIONSHIP FOR COPPER, GOLD AND ZINC IN FUTURE LALOR BASE METAL ORES

Further work was undertaken on the composites to determine the effect of a finer grind on the copper, gold, and zinc recovery. Copper and zinc recovery was not materially impacted by the decrease in primary grind, and this is reflection of the highly liberated state these metals achieved at +100µm. Gold recovery is affected by a reduction in primary grind size, with the recovery improving on average by 1.2% per 10µm reduction in primary grind size.

Mineralogy analysis on the each of the composite samples, indicates a general upward trend in the proportion of pyrite/pyrrhotite in feed. Of concern is the level of galena and talc, which will have a significant impact on the copper flotation circuit without appropriate management strategies. Additional testwork has been initiated to map the lead grade profile across the LOM and the viability of a separate lead flotation stage in the flowsheet. A talc preflotation stage has already been designed to handle the variable levels of talc in feed.

13.2. NEW BRITANNIA CONCENTRATOR

In March 1949, the Nor-Acme mine and gold concentrator was opened. Production continued until 1958, with the Nor-Acme mill recovering approximately 610,000 oz of gold during this period. Between 1958 and the early 1990s, various companies worked on the property with the intention to bring the mine back into production. These efforts were unsuccessful.

In 1994, TVX and High River formed a joint venture to reopen the mine. Using ore from secondary mining areas, the mill was brought into production while completing construction of the main shaft. When production commenced in 1995, the name Nor-Acme was changed to New Britannia. Through various transactions, Kinross became the operator of the New Britannia mine-mill complex. Production mining ceased at the end of September 2004 and the mill was put on care and maintenance in 2005 due to a low gold price environment after producing 1.6 million oz of gold. In 2015, Hudbay acquired the New Britannia mill. It is currently being refurbished to process gold and copper-gold zone ores from the Lalor Mine.

Two significant metallurgical test work programs were undertaken between 2015 and 2020 by SGS Lakefield and Blue Coast Research which form the basis of the concentrator refurbishment project.

The New Britannia is scheduled to achieve commercial production in H2 2021. The concentrator flowsheet is outlined in Figure 13-5. The LOM average metallurgical performance forecasts based on testwork completed and mine production grades are outlined in Table 13-4.

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TABLE 13-4: LOM FORECAST METAL RECOVERY RANGE

FIGURE 13-5: NEW BRITANNIA PROCESS FLOWSHEET

Metallurgical Testwork - SGS Lakefield (2015)

In 2015, SGS Lakefield conducted a metallurgical test program on nine composite samples from within the Lalor gold zones with the objective of evaluating the processing options for the gold zone ores. The testwork included program mineralogy, comminution, gravity separation, flotation, cyanidation, CIP model and cyanide destruction.

Sample Selection

The 2015 test program used 395 kg of core from zone 25 and 315 kg from zone 27. The zone 25 samples were segregated into four variability composite samples and a single overall composite including of waste dilution mass. The zone 27 samples were segregated into three variability composites, a single overall composite excluding waste and an overall composite including waste. A final zone 25 and zone 27 blend composition sample was generated at a ratio of 2/3 and 1/3 zone 25 to zone 27, respectively. The head assays for the test composites are outline in Table 13-5.

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TABLE 13-5: SGS LAKEFIELD COMPOSITE SAMPLE HEAD GRADES

Mineralogical Investigation

A subsample of zone 25 overall and zone 27 overall composite was ground to a p80 of 153um for zone 25 and 101um for zone 27 with the results outlined in Table 13-6. The two composites varied significantly in both non-sulphide gangue and sulphide content. The zone 25 overall composite was composed predominately of quartz, amphibole/pyroxene, and feldspar with the sulphides accounting for less than 5% of the total mass. The zone 27 overall composite was composed predominately of quartz, amphibole/pyroxene, chalcopyrite, and garnets with the sulphides accounting for more than 20% of the mass.

TABLE 13-6: OVERALL COMPOSITE MINERALOGY

Communication Testwork

RWi and BWi tests were completed on several the composites sample with the results outline in Table 13-7 The RWi values range from 9.3 to 13.3 kWh/t for the material tested. Compared to the base metal ore zones (6.0 to 10.0 kWh/t) the material tested from zone 25 and zone 27 is 30 to 50% harder. The BWi values range from 14.5 to 21.0 kWh/t for the material tested. Compared to the base metal ore zones (10.0 to 14.0 kWh/t) the material tested from zone 25 and zone 27 is 40 to 60% harder.

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TABLE 13-7: COMMUNICATION TESTWORK RESULTS

Flotation Testwork

The flotation testwork conducted by SGS Lakefield focused on the removal of copper prior to tailings cyanidation to recovery the residual gold and silver. Rougher, cleaner, and locked cycle tests were completed on the composite samples. The results from the rougher and cleaner tests are outlined in Table 13-8.

High recoveries of copper to the copper concentrate (+90%) and high concentrate grades were achieved where the lead head grade does not exceed the copper head grade. The gold and silver recovery ranged from 40.8 to 91.9%. The test work indicates that further work is required to investigate the potential to scalp a high grade concentrate from the rougher flotation stage.

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TABLE 13-8: ROUGHER AND CLEANER FLOTATION TEAMWORK RESULTS

The lock cycle tests conducted all produced copper concentrates at saleable grade and with metal recoveries consistent with the batch rougher and cleaner flotation tests. Detailed analysis of the concentrates indicates that concentrates produced by the zone 27 material is low in delirious minor elements such as F, As and Hg. The zone 25 concentrates however display elevated levels of Hg and Sb that are in the payable penalty range however as this ore is projected to be blended with zone 27 and the subsequently concentrates produced blended with Stall Concentrator copper concentrates, the impact of these penalty elements will be minimised.

Leach Testwork

Cyanidation leach tests were conducted on the tailings from the lock cycle and rougher tests. The test results are outline in Table 13-9.

From initial sighter tests the standard conditions of 40% solids, 8 hours pre-aeration, 50 g/t lead nitrate and 0.5 g/L cyanide concentrate were selected. The gold extraction values range from 72.2 to 85.6% after 48 hours with most of the extraction occurring within the first 24 hours period. Cyanide consumption to maintain a 0.5 g/L pulp concentration through the test ranged from 0.65 to 2.98 kg/t. As most of the gold extraction occurs within the first 24 hours further optimization testwork is required to leach time to reduce the cyanide consumption.

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TABLE 13-9: FLOTATION TAILS CYANIDATION LEACH TESTWORK RESULTS

Cyanide Destruction Testwork

Cyanide destruction testwork using the SO2/Air Process was conducted on the on the leach tails from tests CN 20, 21, 42 and 44. The reagent sources of used for the tests was sodium metabisulphite (SMBS) for the SO2 and copper sulphate for the Cu2+. The results are outlined in Table 13-17. The CNT and CNWAD values in the leach tails ranged from 232 to 466 mg/L.

For the batch tests conducted on all the samples, <1.00 mg/L CNWAD was achieved with SO2 additions ranging from 7.28 to 17.25 g/g CNWAD and Cu additions ranging from 0.05 to 0.12 g/g CNWAD. In CN-42 continuous tests, the removal of copper sulphate was attempted. The two tests achieved 210.00 and 38.20 CNWAD mg/L after 90min and 120min indicating that copper sulphate is required to catalyze the reaction even though 285 mg/L Cu was present in the feed solution. With the addition of 0.11 g/g copper sulphate to CNWAD and additional SO2 the CNWAD was reduced to <1.00 mg/L.

The cyanide destruction tests were successful at reducing the CNWAD to <1.00 mg/L in all tests and further optimization is required around retention time, SO2 and copper sulphate additions.

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Metallurgical Testwork - Blue Coast Research (2019 to 2020)

From 2019 to 2020, Blue Coast Research conducted a metallurgical test program on thirteen composite samples from within the Lalor gold zones to confirm the optimum processing options for the gold zone ores. The testwork program included mineralogy, flotation, cyanidation, and cyanide destruction.

Sample Selection

The samples selected for the were sourced primarily from zone 27 with a single master composite and nine variability samples. A single zone 25 sample was also included to test individually and to enable the generation of two zone 25 and zone 27 blend composites (50/50 and 75/25 zone 25 to zone 27 respectively). The head assays for the test composites are outline in Table 13-10

TABLE 13-10: BLUE COAST RESEARCH COMPOSITE SAMPLE HEAD GRADES

Mineralogical Investigation

A subsample of each of the composite samples was ground to 90um and submitted for mineralogical analysis with the result outlined in Figure 13-6. The dominant mineral species across all the composite samples where quartz, amphibole, clinochlore and chalcopyrite. The mineralogy of the composite sample was consistent with the 2015 SGS program except for Talc. The level of talc present in each of the composites, particularly the zone 25 composite are significantly higher than in the 2015 program. In the 2015 program the talc content of the zone 25 composite was 1.5%. When compared with the zone 25 composite in the Blue Coast program at ~10%, the significant increase in talc would have an impact on the flotation circuit.

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FIGURE 13-6: COMPOSITE SAMPLE MINERALOGY

Additional focused mineralogy was conducted on the zone 25 and zone 27 master composite to determine the liberation state of the chalcopyrite Results showed that most of the chalcopyrite in both zone 25 and zone 27 is free/liberated across all size fractions with only minor quantities associated with sphalerite, pyrite, and other non-sulphide minerals. This indicates that there is little benefit to the recovery of copper at primary grind sizes below 90um.

Flotation Testwork

The flotation testwork conducted by Blue Coast Research focused on the removal of copper prior to tailings cyanidation to recovery the residual gold and silver. Rougher flotation tests were completed on the zone 25 (SKT 8, 9, 11, 34) and zone 27 master composite sample with the results outlined in Table 13-11.

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TABLE 13-11: ROUGHER FLOTATION TESTWORK RESULTS

The rougher flotation tests confirmed that high recoveries are achievable for copper, gold, and silver in the rougher flotation stage. The tests on zone 25 master composite averaged a rougher concentrate copper grade of 2.0 to 2.5% Cu compared to the zone 27 composite rougher concentrate grade of 10.7 to 17.1%. The principal driver for the depressed concentrate grade in the zone 25 rougher tests was the presence of high levels of Talc in the feed. The MgO grade in the zone 25 test was 12 to 14% indicating the bulk of the mass in concentrate was Talc.

The final stage of the rougher flotation testwork was to test the zone 25 and zone 27 blend composites (Table 13-27). The blend of the high talc zone 25 ore with the cleaner zone 27 ore did not impact recovery however the concentrate grade was significantly impacted. The greater the proportion of zone 25 ore in the blend the lower the achievable concentrate grade dropped. In the 50/50 the concentrate MgO ranged from 9 to 11% with the 75/25 blend ranging from 12 to 14%.

Cleaner flotation testwork was conducted on a bulk concentrate sample generated from the zone 27 composite. The zone 27 ore displayed exceptional flotation response with recovery of copper, gold and silver all exceeding 96% and at a saleable concentrate grade.

Cleaner flotation testwork was also conducted on the blend composites. As found in the rougher flotation tests, the recovery was not adversely affected by the blending on the zone 25 and zone 27 ores. The effect of blending these ores was the depression of the concentrate grade due to the presence of talc. In the 50/50 blend the MgO in the concentrate was 8 to 10% and in the 75/25 blend it was 11 to 12%. The addition of CMC did not significantly impact the concentrate quality.

The presence of talc at these levels indicates that a dedicated preflotation stage is required to selectively recovery the talc directly to flotation tailings ahead of the copper flotation circuit to ensure a saleable concentrate grade can be produced. Additional flotation tests were conducted on the efficacy of a preflotation stage with the results outline in Figure 13-7. In the presence of only MIBC and at pH 9.5 the talc was recovered to the concentrate in a single rougher stage. The subsequent recovery of copper and gold to the talc concentrate was unacceptably high and required additional optimization. The addition of a sulphide depressant (Orform D8 or NaCN) significantly reduced the recovery of copper and halved the recovery of gold to the talc concentrate. As the flotation tailings in the New Britannia circuit reports to the leach circuit the loss of this level of gold is acceptable as it will be recovered to Doré. Additional optimization is required including investigating the effectiveness of a talc cleaning stage as is typical in base metal concentrators operating a talc preflotation circuit.

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FIGURE 13-7: TALC PREFOLIATION ROUGHER TESTWORK RESULTS

Two lock cycle tests were conducted on 50/50 blend composite without a talc preflotation stage. The major changes between each of the lock cycle test was the removal of the cleaner scavenger stage, the change in pH and the increase in CMC addition.

The results for the lock cycle tests are outline in Table 13-12. LCT-1 was operated aggressive generating an exceptional recovery at the expense of final concentrate grade. It is important to note that for LCT-1 the effort to control talc recovery was minimal with the resulting final concentrate grade of 15.4% Cu and +6.0% MgO. LCT-2 increased the effort to remove MgO from the concentrate with the CMC increase in the rougher from 150 g/t to 400 g/t and in the cleaners from 25 g/t to 50 g/t. LCT-2 achieved a 25.9% Cu and 1.0% MgO concentrate grade however at a significant discount to copper and gold recovery.

TABLE 13-12: LOCK CYCLE TEST RESULTS

13.3. FLOTATION TAILS LEACH TESTWORK

Cyanidation leach tests were conducted on the tailings from the lock cycle and rougher tests (Table 13-13). The flotation tailings from the batch flotation tests, bulk flotation tests (cleaner testwork feed generation) and lock cycle tests were used as the feed stock for the test work. The base conditions for the flotation tails leach testwork were 40% solids, 48-hour leach duration, 0.5 g/L cyanide concentration and pH at 10.5 to 11.0.

CN-1 and CN-2 were grind size variability flotation tests with a flotation feed p80 of 90µm and 120µm, respectively. CN-1 and CN-2 show that at a p80 grind of 90µm 74% gold extraction is achieved verses 68.4% gold extraction at 120µm. A 90µm grind was carried forward for all the subsequent leach tests. CN-3 to CN-5 assessed the impact of pre-aeration on the gold extraction. Minimal difference between no pre-aeration and 4 hours pre-aeration was found and unless stated, pre-aeration was not utilized in future tests.

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TABLE 13-13: FLOTATION TAILS LEACH TESTWORK GOLD EXTRACTION RESULTS

All tests achieved low cyanide consumptions except for CN-14. CN-14 feed material was the tailings from LCT-2 which reported higher residue copper in the flotation tailings at 0.18% versus the residual copper in tailings for CN-1 to CN-13 between 0.03 to 0.10%. With the correct operation of the flotation circuit ahead of the leach stage, cyanide consumption below 1.0 kg/t is achievable.

To minimize the loading of copper onto carbon in the CIP stage the CNFree/Cu molar ratio in leach liquor must be maintained above 4.0. Most tests exceeded the required CNFree/Cu molar ratio low limit. CN-14 low CNFree/Cu molar ratio was a direct result of the elevated residual copper grade in the flotation tailings. CN-9 and CN-10 were the dosage tests for lead nitrate which indicates the addition of lead nitrate enhanced the leach rate of copper. The leach liquor copper tenor was 95 ppm without the addition of lead nitrate and rose to 154 ppm with its addition. The addition of lead nitrate does not improve the extraction of gold and appears to enhance the extraction rate of copper potentially causing downstream issues with carbon loading.

13.4. 1901 BASE METAL ORE METALLURGICAL TESTWORK

1901 is a satellite deposit located between the Chisel and Lalor deposit to the east side of the decline. This deposit consists of four distinct lens, two high zinc base metal lenses and two high gold bearing lenses. reportable level.

In 2020 a metallurgical testwork program was initiated with Blue Coast Research to cover the four identified ore lenses. At the time of writing, only testwork for the two base metal lenses, Lens 1 and 3 has been completed to a

Sample Selection

Three composite samples were created for each of base metal lens. As the base metal lenses are zinc dominant the composite sample were select to represent a low, average, and high zinc grade associated with the lens. The distribution of the drill core selected to create the metallurgical testwork composites are shown in Figure 13-8 and the grades in Table 13-14.

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FIGURE 13-8: METALLURGICAL TESTWORK COMPOSITE SAMPLE DRILL CORE LOCATION

TABLE 13-14: 1901 BASE METAL LENSE COMPOSITE ASSAYS

Mineralogical Investigation

A subsample of each of the composite samples was ground to a p80 of 100µm and submitted for mineralogical analysis with the result outlined in Figure 13-9. The dominant mineral species across all the composite samples is pyrite, chlorite, sphalerite, and dolomite. Chalcopyrite and galena are the principal minerals for copper and lead, respectively.

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FIGURE 13-9: 1901 COMPOSITE SAMPLE MINERALOGY

Flotation Testwork

The flotation testwork conducted by Blue Coast Research on the 1901 composite samples used a sequential copper-zinc flotation circuit approach. Flotation tests were completed on each of the six composites with the bulk of the optimization testwork carried out on the master composite sample from both Lens 1 and Lens 3. The feed grades for each of the composite samples shows that the lead assay is higher than the copper assay indicating that lead contamination of the copper concentrate is likely.

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TABLE 13-15: 1901 FLOATATION TESTWORK COPPER CIRCUIT RESULTS

To examine the potential to selectively recovery copper and gold from lead, FKT-2 was run utilizing the collector TNC 312. The use of TNC 312 decreased the lead recovery from 95.2% to 22.5% and an improvement in the rougher concentrate grade from 2.6% Cu to 6.7% Cu. The cost of this improvement was a reduction in copper and gold recovery to 83.5% and 45.5% respectively which is an unacceptable reduction.

At the time of writing, further testwork is underway to examine the potential to include the same lead recovery stage into the sequential flotation circuit already considered for Lalor's future production. This testwork will operate a copper-lead-zinc sequential flowsheet utilizing TNC 312 as the copper circuit collector and 3418A as the lead circuit collector.

Zinc circuit flotation tests are consistent with testwork and historical performance from the Lalor Base metal ores, achieving +97% zinc circuit rougher recovery.

13.5. WIM METALLURGICAL TESTWORK

WIM is a satellite deposit located north of the Lalor deposit. This deposit consists of one ore zone that host of high-grade copper and moderate gold grade ore. In 2020 a metallurgical testwork program was completed by Blue Coast Research with the focus being the performance of the WIM ore through the New Britannia mill. No tailings cyanidation testwork was completed in the testwork campaign.

Sample Selection

Four composite samples were created for the WIM ore zone. Three variability composite samples were created based on the spatial distribution of the drill core locations outlined in Figure 13-10. A single master composite was also created from subsamples of the three variability samples. The head grades for all four composite samples are outline in Table 13-16.

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TABLE 13-16: WIM COMPOSITE SAMPLE HEAD GRADES

FIGURE 13-10: WIM COMPOSITE SAMPLE DRILL HOLE LOCATIONS

Mineralogy

A subsample of each of the composite samples was ground to a p80 of 90µm and submitted for mineralogical analysis with the result outlined in

Figure 13-11. The dominant mineral species across all the composite samples is pyrite, pyrrhotite, quartz and chalcopyrite. The level of pyrrhotite in the feed sample is significantly higher than comparable ore in the Lalor zone 25 or zone 27.

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FIGURE 13-11: WIM COMPOSITE SAMPLE MINEROLOGY

Comminution Testwork

Standard Bond ball mill work index tests were completed on each of the composite samples with a closing screen size of 106µm. The BWi for the samples ranged from 17.9 to 18.5kWh/t. This is comparable to the Lalor gold zone ores.

Flotation Testwork

The flotation testwork conducted by Blue Coast Research on the WIM composite samples used a single copper flotation circuit configuration. Flotation tests were completed on each of the four composites with the bulk of the optimization testwork carried out on the master composite sample to determine the optimal conditions for the variable composites.

TABLE 13-17: WIM MASTER COMPOSITE ROUGHER FLOATATION TESTWORK RESULTS

For tests F-1 through F-6 Aerophine 3418A was the only collector with F-7 utilizing only TNC 312. All tests achieved exceptional copper and gold recoveries at +96% and +73% respectively. The conditions from F-7 (pH 11.5, TNC 312) were selected as the optimal rougher baseline conditions due to the improved selectivity against iron sulfides when compared to the 3418A based tests (18.6% S recovery verses +34% S recovery). Additionally, as the WIM ore would be processed through the New Britannia mill, flotation gold recoveries are not as critical due to the tails leach operation.

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Following the selection of the baseline rougher conditions, two stage cleaner flotation tests were carried out without a regrind step. Final concentrate grade was achieved (+20% copper) at +96% copper recovery indicating this ore is amenable to the New Britannia flotation circuit.

FIGURE 13-12: WIM COMPOSITE SAMPLE CLEANER FLOTATION TESTWORK GRADE RECOVERY CURVES

The conditions identified in the master composite flotation testwork were repeated in for the three variability composites. Each of the variability samples achieved +99% copper and +65% gold recovery in the rougher stage. Outlined in Figure 13-12, each of the composite samples were able to achieve +20% copper concentrate grade at +95% recovery.

13.6. 3 ZONE METALLURGICAL TESTWORK

3 Zone is a satellite deposit located north of the Lalor deposit and is part of the historical New Britannia mine. This deposit consists of one ore zone that host of high-grade copper and moderate gold grade ore. In 2020 a metallurgical testwork program was completed by Blue Coast Research with the focus being the performance of the ore through the New Britannia mill. This ore is a primary gold ore with minimal sulfur content, to that extend only gravity concentration and cyanidation testwork was performed on the selected samples.

Sample Selection

Four composite samples were created for the 3 Zone ore. Three variability composite samples were created based on the spatial distribution. A single master composite was also created from subsamples of the three variability samples. The head grades for all four composite samples are outline in Table 13-18.

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TABLE 13-18: 3 ZONE COMPOSITE SAMPLE HEAD GRADES

Mineralogy

A subsample of the master composite samples was ground to a p80 of 75µm and submitted for sized mineralogical analysis with the result outlined in. The dominant mineral species were quartz, amphibole, biotite, and plagioclase. The principal sulfide mineral is arsenopyrite.

FIGURE 13-13: 3 ZONE MASTER COMPOSITE SAMPLES SIZED MINEROLOGY

Leach Testwork

Leach tests were completed on each of the four composites with the bulk of the optimization testwork carried out on the master composite sample to determine the optimal conditions for the variable composites. The results of the initial seven optimization tests are outline in Table 13-19. All tests exceeded 89% gold recovery with exceptionally low consumption of cyanide indicating amenability to direct cyanidation. CN-3 conditions were carried through as the baseline test conditions.

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TABLE 13-19: 3 ZONE MASTER COMPOSITE LEACH TESTWORK RESULTS

The tests were repeated for the three variability composite samples with gold recovery and cyanide consumption rates 85 to 95% and 0.25 to 0.62kg/t, respectively. The higher cyanide consumption and lower recoveries positively correlate to the arsenic feed grade indicating that the gold in the 3 Zone ore is refractory in the arsenopyrite and that further work is required to identify the process route to recovery the fraction of gold unrecoverable by the current New Britannia leach circuit.

FIGURE 13-14: 3 ZONE ARSENIC HEAD GRADE CORRELATION WITH LEACH RESIDUE GOLD GRADE

Gravity Concentration Testwork

Gravity recovery amenability testwork was carried out on the 3 Zone master composite. The ground sample was initial upgraded through a Knelson concentrator. The Knelson concentrator achieved 78.6% gold recovery at a 1.2% mass recovery. This indicates that the 3 Zone ore present in the master composite is amenable to gravity concentration methods and that further testwork is required to map the gravity recoverable gold potential through the 3 zone orebody.

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TABLE 13-20: 3 ZONE KNELSON CONCENTRATOR TESTWORK RESULTS

13.7. PEN II METALLURGICAL TESTWORK

Metallurgical test work conducted on composites representative of the Pen II deposit support the assumption that the zinc rich inferred mineral resource estimates could be processed at the Stall concentrator with similar performance as observed from Lalor and 1901 mineral reserve estimate. The reader is referred to the March 2019 Lalor and Snow Lake Technical report for more details on the test work conducted and results obtained.

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14. MINERAL RESOURCE ESTIMATES

No known factors exist which could materially affect the mineral resource estimates.

14.1. LALOR MINE

Hudbay prepared an update of the Lalor mine resource model using Leapfrog® version 5.1.1 and MineSight® version 15.60, two industry standard commercial geological and mining software. The construction of this 3D resource model and the estimation of mineral resources were performed by Hudbay personnel following Hudbay procedures in compliance with best industry standards and the CIM guidelines. The work was supervised and validated by Olivier Tavchandjian, P.Geo., Vice President, Exploration and Geology and Qualified Person for the purpose of the present Technical Report.

Modeling Database

4,118 drill holes totaling approximately 650,000 m were included in the Lalor database to support the mineral resource estimate. Amongst these samples approximately 397,300 m were analyzed for Zn, Cu, Au and Ag.

The drill hole database was directly imported in MineSight® from acQuire® with a cut-off date for mineral resource estimate purposes of September 8th, 2020. In parallel, the drill hole database was also imported in Leapfrog® for additional and independent comparison and validation. No significant differences in the projection of the drill hole traces were found between the two datasets.

Modeling of the Mineralized Envelopes

The Lalor mineralized envelopes trend along an azimuth of approximately N320° with a general dip of approximately 37° to the north east. Economic mineralization delineated so far at Lalor occurs in nine semi-massive to massive sulfide lenses typical of VMS deposits (lenses 10, 11, 20, 30, 31, 32, 33, 40 and 42), three conformable gold-rich disseminated to stringer zones (lenses 21, 23 and 25), two copper-gold rich feeders (lenses 17 and 27) and as small footwall and hanging wall gold rich zones associated to lens 10 (lenses 9, 13 and 16) The base metal and gold rich zones are stratigraphically stacked and are mostly aligned to the main regional deformation, averaging N320°/37°. The Lalor mine is continuous along a strike length of approximately 1,400 m in north-south direction, approximately 780 m in an east-west direction and with a vertical extent of approximately 760 m.

Geological units of sulphides horizons were created in 3D using core logging data and assay results. These broad geological envelopes were then refined to create base metal and gold rich mineralized envelopes. The distribution of gold mineralization follows two styles: the first one, which is synchronous to the base metal deposition, is directly associated and contained within the base metal mineralized envelopes, typically on their footwall. The second one, posterior to the base metal deposition, is contained outside the sulphides horizons and is typically sulphides poor with the exception of the Au-Cu rich mineralized envelopes 17 and 27 which are Au-Cu copper stringer zones.

The construction of the mineralized envelopes was based on the type of mineralization intersected. Smooth and continuous 3D solids were built using 4% zinc or 1% copper cut-off to guide the construction of the base metal mineralized envelopes, with the inclusion of any nearby gold intersections, while a 2.0 to 2.5 g/t cut-off was used to guide the construction of the gold rich mineralized envelopes.

In a final validation step, these envelopes were loaded into MineSight® in order to ensure proper tagging of the solids to actual drill hole locations and that drillhole traces matched both in Leapfrog and Minesight. The mineral envelopes were used as hard boundaries in all cases for grade interpolation purposes so as to correctly prevent spreading of sulfide mineralization into the gold zones and vice-versa. Wireframes of the mineralized envelopes are shown in Figure 14-1.

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Density

Density values generated from multi regression formulas are based on 25,182 assay measurements. The predicted density values were validated through comparisons against the measured density.

FIGURE 14-1: 3D VIEW OF MINERAL ENVELOPE WIREFRAMES, LOOKING WEST

Note: Lens 23 is located under lens 32 and not shown in this view.

Compositing

Assay intervals were regularized by compositing drill hole data within each interpreted mineralized envelope. Although most of the drill holes were assayed on 1m interval, a composite length of 2.5 m was selected as more appropriate to conduct interpolation into the 5 m x 5 m x 5 m block size selected by the mine engineering team as the optimum selective mining unit (SMU).

Grade compositing was weighted by specific gravity as there is a significant positive correlation between some of the metal grade and density. The compositing process was validated by comparing total length, density and length weighted average grade for each metal of the 2.5 m composites to the original assays.

Exploratory Data Analysis

Exploratory data analysis (EDA) includes basic statistical evaluation of the composites for gold (Au), silver (Ag), copper (Cu), zinc (Zn), and density (SG). The EDA was conducted separately for each mineralized envelope and also aimed to identify sub-domains that did not support the use of hard contacts for grade interpolation purposes but that justified being used for the purpose of block model validation, smoothing assessment and correction and resource classification. Capping analysis and smoothing assessment were also conducted separately in each sub-domain. An example of sub-domaining is illustrated for gold zone 25 on Figure 14-2 where the high-grade core of the lens is subdivided based on drilling coverage and also from its lower grade halo as defined by Leapfrog Geo interpolations.

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FIGURE 14-2: SUB-DOMAIN IN MINERALIZED ENVELOPE 25

Note: drill hole traces represented by the lines in grey

The 2.5 m composite statistics for Au, Ag, Cu, Zn and calculated SG are summarized in Table 14-5 in the grade estimation validation section. As expected, zinc shows significantly lower CV and skewness values in the base metal lenses while the opposite is observed for gold.

Grade Capping

Because the composites of the four metals of economic potential show highly skewed histograms in most of the mineralized envelopes, the deciles^2^ analysis (Parish) method was used to define high-grade outliers and to confirm the need for grade capping. This analysis was conducted on the 2.5 m composites for each sub-domain independently.

This method considers capping when the last decile of the population contains more than 40% of the metal and that the last percent contains more than 10% of the metal. Capping values were selected to limit the weight of the high-grade outliers on the overall population. Table 14-1 summarizes the capping applied to all the mineralized envelopes for gold and silver and the proportion of metal removed through this process. In all cases, it was found that no capping would be required for zinc or copper.

^2^ I.S. Parrish (1997) Mining Engineering Journal, Geologist's Gordon Knot: to cut or not to cut (pp. 45-49)

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TABLE 14-1: 2021 CAPPING THRESHOLDS BY LENS OF GOLD AND SILVER (G/T)

Note: HGO = high grade and LGO = low grade

The results obtained from the Parish method were then compared to the natural breaks in the cumulative frequency distribution in the last decile. The metal at risk was also assessed in the population of outliers by comparing the difference in metal using the median and mean values of these outliers. As mentioned in our previous NI43-101, an independent third-party review was conducted in November 2018 by Mr. Dominique Francois-Bongarçon (Agoratek) who concurred with the use of the Parrish as a valid approach for gold capping in the gold zones.

Variography

Down-hole and directional variograms for Zn, Au, Cu, Ag, and SG were created for each individual mineral envelope using the MineSight data analysis software. Pairwise relative variogram were preferred due to the skewed nature of metal distribution in most lenses. The major, semi-major and minor axis were built in order to fit the attitude of each mineral envelope. A linear combination of a nugget and two nested spherical models were adjusted in all cases. Once generated, a systematic visual check was conducted to ensure that the search ellipsoid would be correctly oriented with respect to the geometry of the mineral envelopes.

Grade Estimation and Interpolation Methods

The block model consists of regular blocks (5 m along strike by 5 m across strike by 5 m vertically). The block dimensions were selected to match the smallest mining unit (SMU) at Lalor mine. Where a block was intersected by two lenses, the mineral envelopes were used to assign the percentage of the block that belongs to each lens.

Both nearest neighbour (NN) and ordinary kriging (OK) grade interpolations were completed on the uncapped and capped grades, using a strict composite and block matching code by lens and three passes with increasing minimum information requirements (Table 14-2). The search passes were selected to ensure best local estimates recognizing that OK has a smoothing effect but making no attempt during interpolation to reduce this smoothing as it would negatively impact the quality of the local estimates. Over-smoothing is addressed through the post-processing of the model described in sub section "smoothing assessment".

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Grade Estimation Validation

The grade estimation process was validated for each envelope to ensure appropriate honouring of the input data and subsequent unbiased resource reporting and use of the model for reserve estimation.

TABLE 14-2: SEARCH ELLIPSE PARAMETERS

Visual Inspection

Visual inspection of block grade versus composited data was systematically conducted in section and plan views. This check confirmed a good reproduction of the data by the model. An example is illustrated in Figure 14-3 for the zinc grade in zone 20.

FIGURE 14-3: VISUAL CHECK ON MODEL FOR ZN GRADE IN LENS 20

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Global Bias Checks

This validation step consists in comparing the global average grade of each element (after capping for gold and silver) between the original composites, the kriged block estimates and the nearest neighbour estimates. This investigation was conducted not only by envelope but also for each individual sub-domain. The average grade obtained from the nearest neighbor interpolation is a useful benchmark but not a perfect one as it fails to incorporate the nugget effect measured by the variogram. A global check was performed to verify that the kriged mean block estimate was located between the mean of the composites and the mean of the nearest neighbour model.

Smoothing Assessment

The larger number of composites used for grade estimation in the block model significantly improves the individual block grade estimates but, at the same time, results in a much smoother model requiring a careful assessment and, in many cases, a post-processing of the OK estimates. The extent of grade 'over-smoothing' in the model was investigated separately by sub domains based on material differences in grade distribution and/or drilling density. The mean and variance of the kriged estimates were compared to the variance of the composites after declustering (obtained from a nearest neighbour interpolation). The expected true variance between SMUs can be calculated from the variogram models.

Smoothing Correction

An indirect log-normal correction was used to perform a change of support on the kriged models in order to obtain unbiased grade tonnage curves for the selected mining block size of 5mx5mx5m. This correction is only valid globally and provides poorer local estimates than the smoothed OK model but does not materially alter the global average grade within each zone and provides the correct grade-tonnage curve for the variogram models fitted on the drill hole data. This correction was applied by domain with consistent statistical properties and uniform drilling coverage. The targeted variance was reached within acceptable limits in most cases, as illustrated in Table 14-3.

TABLE 14-3: SUMMARY OF THE APPLIED SMOOTHING CORRECTION

Note: HG = high grade, LG = low grade, HDD = high drilling density & LDD = lower drilling density

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Classification of Mineral Resources

During the interpolation process, several control parameters were recorded for each block, e.g., number of samples, number of holes, the distance to the nearest sample and the average distance to all the samples used for the interpolation as well as the number of quadrants with samples, the kriging variance and the regression slope of kriging for each individual block estimate.

The regression slope values obtained from the kriging of zinc (in the case of base metals lenses) or gold (in the case of gold zones) grade estimates was used as the primary criteria for resource classification with 80% and 60% regression slope thresholds used respectively to separate measured from indicated and from inferred resources. From detailed reserves to mill reconciliations exercises conducted by Hudbay at its operating mines, including Lalor, this criteria was found to be a reliable first pass measure of quarterly and annual performance in tonnes and grade prediction.

The block by block coding assignation was then smoothed to remove isolated blocks of one category within another. Proportions of measured, indicated and inferred category blocks were not changed significantly through this process. In assigning the final categorization, careful consideration was also given to the proportion of mineralization grading above the anticipated cut-off grade. Gold lenses followed a more prudent approach in which some categories were downgraded and measured was only applied in areas of high drill density or which was silled through with development. Figure 14-4 illustrates the classification process followed for Lens 25.

FIGURE 14-4: RESOURCE CLASSIFICATION IN MINERALIZED ENVELOPE 25

Note: the insert on the left presents the block by block resource classification (Measured in green, indicated in yellow and Inferred on orange) based on the regression slope of gold in mineralized envelope 25 while the insert on the right present the final classification. Confidence was down graded based on the proportion of blocks above cut off.

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Reasonable Prospects of Economies Extraction and Mineral Resource Estimates

Since 2019, Hudbay has implemented a stringent approach to resource reporting for underground deposits. With this approach, the potential for economic extraction of the mineral resource estimates at Lalor are reported within the constraint of a mineable stope-shape, created using Stope Shape Optimizer (SSO). The SSO application mimics what a mine planner would do, generating strings on sections, linking these to create a wireframe shape and then evaluating the wireframes against a block model. The SSO provides a stope-shape that maximizes recovered resource value above a cut-off while also catering for practical mining parameters (such as minimum and maximum mining width, anticipated wall dilutions, minimum and maximum wall angles, minimum separation distances between parallel and sub-parallel stopes, minimum and maximum stope heights and widths, etc.). The mining parameters used to build the mineable shapes are detailed in Section 15.

NSR cut-off values are used at Hudbay in the SSO process, whereby a calculation of value per tonne is applied to the block model based on anticipated metal recoveries, metal prices and exchange rate, and concentrate transportation, smelting, and refining costs. The cut-off value input parameter in SSO considers operating and capital costs, as based on the current Lalor mine and mill budgets, to ensure that each potential stope created would cover all its associated operating mining and milling costs to be deemed profitable if mined. This process ensures that reported mineral resource estimates exclude small isolated individual resource blocks that may meet economic cut-off criteria on an individual basis but could not be aggregated into mineable shapes and include 'geological dilution' (representing small blocks grading below cut-off) in the economic shape to maintain minimum mining criteria.

The stope optimization process was run on all resource categories combined and as a result, a small amount of one resource category could be included in a stope shape dominated by another resource category. This usually occurred on the fringes of the zones defined for resource classification. Resource classification is based on broad smoothing of interpolation parameters used in the construction of the resource model and visual inspections confirmed that minor modifications to the resource categories were warranted so that all stope shapes would be entirely assigned to one resource category.

The resulting mineral resource estimates, as of January 1, 2021, are presented in Table 14-4. There are no measured or indicated mineral resource estimates in addition (exclusive) to those that were converted to mineral reserve estimates. All measured and indicated mineral resource estimates were either converted to mineral reserve estimates, as documented in section 15, or deemed non-economic after applying the appropriate resource to reserve conversion factors.

TABLE 14-4: LALOR MINERAL RESOURCE ESTIMATES

Notes:

Totals may not add up correctly due to rounding.
Mineral resources that are not mineral reserves do not have demonstrated economic viability.
Mineral resources in the above tables do not include mining dilution or recovery factors.
Base metal mineral resources are estimated based on the assumption that they would be processed at the Stall concentrator while gold mineral resources are estimated based on the assumption that they would be processed at the New Britannia concentrator, which is currently being refurbished.
Metal prices of US$1.10 per pound zinc, US$1,500 per ounce gold, US$3.10 per pound copper, and US$18.00 per ounce silver with an exchange rate of 1.30 C$/US$ were used to estimate mineral resources.
Mineral resources are estimated using a minimum NSR cut-off of $105 per tonne for waste filled mining areas and a minimum of $116 per tonne for paste filled mining areas.

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14.2. 1901 DEPOSIT

Hudbay prepared resource model using Leapfrog® version 5.0.3 and MineSight® version 15, two industry standard commercial geological and mining softwares. The construction of this 3D resource model and the estimation of mineral resources were performed by Hudbay personnel following Hudbay procedures in compliance with best industry standards and the CIM guidelines. The work was supervised and validated by Olivier Tavchandjian, P.Geo., Vice President, Exploration and Geology.

Modeling Database

Drill holes totaling approximately 80,600 m were included in the 1901 database to support the mineral resource estimate update. Amongst these, approximately 18,300 were analyzed for Zn, Cu, Au and Ag while density was measured for 15,150m.

The drill hole database was directly imported in Leapfrog from acQuire® with a cut-off date for mineral resource estimate purposes of June 9th, 2020. The drill hole database was then exported from Leapfrog and imported into MineSight via .csv files. The drillhole traces in MineSight were validated against the drillhole traces in Leapfrog to ensure no errors were introduced during the import process.

Modeling of the Mineralized Envelopes

The 1901 mineralized envelopes trend along an azimuth of approximately 340° with a general dip of approximately 25° to the east. Economic mineralization delineated so far occurs in two semi-massive to massive sulfide lenses typical of VMS deposits (lens 1 & 3) and two conformable gold-rich disseminated to stringer zones (lens 2 & 4). The base metal and gold rich zones are stratigraphically stacked and are mostly aligned to the main regional deformation. The mineralization is continuous along a strike length of approximately 550 m in north-south direction, approximately 230 m in an east-west direction and with a vertical extent of approximately 80 m.

Geological units of sulphides horizons were created in 3D using core logging data and assay results. These broad geological envelopes were then refined to create base metal and gold rich mineralized envelopes. The distribution of gold mineralization appears follows two styles: the first one, which is synchronous to the base metal deposition, is directly associated and contained within the base metal mineralized envelopes, typically on their footwall. The second one, posterior to the base metal deposition, is contained outside the sulphides horizons and is typically sulphides poor.

The construction of the mineralized envelopes was solely based on the type of mineralization intersected. To this effect, assay intervals with high Au grade but no zinc or copper were excluded from base metal lenses and vice versa. Instead, smooth and continuous 3D solids were built using 3% zinc or 1% copper cut-off to guide the construction of the base metal mineralized envelopes while a 2 g/t cut-off was used to guide the construction of the gold rich mineralized envelopes.

In a final validation step, the envelopes were loaded into MineSight in order to ensure proper tagging of the solids to actual drillhole locations. The mineral envelopes were used as hard boundaries in all cases for grade interpolation purposes so as to correctly prevent spreading of sulfide mineralization into the gold zones and vice-versa. Wireframes of the mineralized envelopes are shown in Figure 14-5.

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FIGURE 14-5: VIEW OF MINERAL ENVELOPE WIREFRAMES

NOTE: Lens 1 in yellow, 2 in green, 3 in red and 4 in fuchsia

Density

Since almost all the assay intervals contained within the different mineralized lenses had density measurements (only two samples with missing density in lens 3), density was only predicted for the barren material surrounding the mineralized zones. As for the two missing density measurements in lens 3, they were set to the average density measured in the lens (i.e. 3.52). Density values generated from a multi regression formula is based on 14,408 density measurements. The predicted density values were validated through comparisons against the measured density.

Compositing

Assay intervals were regularized by compositing drillhole data within each interpreted mineralized envelopes. Although most of the drillholes were assayed on 1m interval (Figure 7 4), a composite length of 2.5 m was selected as more appropriate to conduct interpolation into the 5 m x 5 m x 5 m block size selected by the project evaluation team as the optimum selective mining unit (SMU). Grade compositing was weighted by specific gravity as there is a positive correlation between some of the metal grade and density. The compositing process was validated by comparing total length, density and length weighted average grade for each metal of the 2.5m composites to the original assays.

Exploratory Data Analysis

Exploratory data analysis (EDA) includes basic statistical evaluation of the composites for zinc (Zn), gold (Au), copper (Cu), silver (Ag) and density (SG). The EDA was conducted separately for each mineralized envelope. Within each mineral envelope, the EDA also aimed to identify sub-domains that did not support the use of hard contacts for grade interpolation purposes but that justified being used for the purpose of block model validation, smoothing assessment and correction and resource classification.

The 2.5 m composite statistics for Zn, Au Cu, Ag and calculated SG are summarized in Table 7 5 in the grade estimation validation section. Overall, the statistics of the composites indicate that the kriging interpolation will as expected introduce some level of smoothing and that a thorough smoothing assessment and post-processing of the initial interpolation is warranted to produce unbiased grade-tonnage curves and mineral resource and mineral reserve estimates.

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Grade Capping

The deciles analysis (Parish) method was used to define high-grade outliers and to assess the need for grade capping. This analysis was conducted on the 2.5m composites for each sub-domain independently.

This method considers capping when the last decile of the population contains more than 40% of the metal and that the last percent contains more than 10% of the metal. Based on this method, gold was capped at 14 PPM in lens 1, 75 PPM in lens 2 and 20 PPM in lens 3, while silver was capped at 375 PPM in lens 2 and 160 PPM in lens 3. Capping was not applied to lens 4 composites given their limited numbers. Overall, capping removed the gold metal content by 2.7% in lens 1, 19.4% in lens 2 and 10.6% in lens 3, while capping removed the silver metal content by 15% in lens 2 and 2% in lens 3.

Variography

Down-hole and directional variograms for Zn, Au, Cu, Ag, Fe, Pb, As and SG were created for lens 1, 2 and 3 using the MineSight data analysis software. Given the limited number of composites and lens 4, the variogram parameters of lens 2 were adapted to match the nugget of lens 4 and its overall strike and dip.

The major, semi-major and minor axis were built in order to fit the attitude of each mineral envelope. A linear combination of a nugget and two nested spherical models were adjusted in all cases. Once generated, a systematic visual check was conducted to ensure that the search ellipsoid would be correctly oriented with respect to the geometry of the mineral envelopes.

Grade Estimation and Interpolation Methods

The block model consists of regular blocks (5 m along strike by 5 m across strike by 5 m vertically). Grade interpolation used a strict composite and block matching system based on the mineralized envelopes codes. For example, only composites coded as lens 1 were used to estimate the grade of the block that was assigned to lens 1. Both nearest neighbour (NN) and ordinary kriging (OK) grade interpolations were completed, using three passes with increasing minimum information requirements (Table 7 4). The search passes were selected to ensure best local estimates recognizing that OK has a smoothing effect but making no attempt during interpolation to reduce this smoothing as it would negatively impact the quality of the local estimates.

TABLE 14-5: SEARCH ELLIPSE PARAMETERS

The NN interpolation is solely used to validate that there is no global bias and to assess the level of grade smoothing in the OK interpolation for the selected SMU dimension.

Grade Estimation Validation

Visual checks of appropriate honouring of the input data but acknowledging that some natural smoothing should occur between samples as the grade of a sample in the middle of a block is not the average grade of the block;

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Absence of global bias by comparing the mean grade estimated by kriging to the original composite average grade and to a declustered grade obtained from a nearest neighbour interpolation.
Change of support assessment and correction.

Visual Inspection

Visual inspection of block grade versus composited data was systematically conducted in section and plan views. This check confirmed a good reproduction of the data by the model.

Global Bias Checks

This validation step consists in comparing the global average grade of each element between the original composites, the kriged block estimates and the nearest neighbour estimates. This investigation was conducted by envelope.

Smoothing Correction

An indirect log-normal correction was used to perform a change of support on the kriged models in order to obtain unbiased grade tonnage curves. For some of the elements, the correction did not fully attain the targeted variance reflecting that the log-normal model does not perfectly fit these elements. However, the targeted variance was reached within very close limits in most cases (Table 14-6).

TABLE 14-6: SUMMARY OF SMOOTHING CORRECTIONS - 1901 DEPOSIT

Note: HG = high grade & LG = low grade

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Classification of Mineral Resource

As for Lalor, the regression slope values obtained from the kriging of zinc (for lens 1 and 3) grade estimates was used as the primary criteria for resource classification. Blocks from the second and third interpolation pass combined with a 60% regression slope thresholds were separate indicated from inferred resources. The block by block coding assignation was then smoothed to remove isolated blocks of indicated within areas of mostly inferred category or isolated inferred blocks within areas of mostly indicated category blocks. Proportions of indicated and inferred category blocks were not changed significantly by this process. Figure 14-6 presents a plan view of the resource categories.

Given the lower level of confidence caused by the highly skewed distribution of gold and limited amount of drilling, all blocks contained within lens 2 and 4 were classified as inferred.

FIGURE 14-6: PLAN VIEW OF THE RESOURCE CLASSIFICATION

NOTE: MEASURED IN GREEN, INDICATED IN yellow and Inferred in orange

As for Lalor, the mineral resource estimates for 1901 are reported within the constraint of an optimized stope-shape. The parameters used as input to define the mineable shape covers all the relevant technical and economic constraints including minimum stope and waste pillar dimensions and the NSR value calculation are very similar to those used at Lalor for the base metal and gold lenses assuming beneficiation at the Stall and New Britannia mills. The mineral resource estimates were reported to ensure that each potential stope would cover all its associated operating and capital costs. The resulting mineral resource estimates are presented in Table 14-7.

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TABLE 14-7: 1901 MINERAL RESOURCE ESTIMATES

Notes:

Totals may not add up correctly due to rounding.
Mineral resources that are not mineral reserves do not have demonstrated economic viability.
Mineral resources in the above tables do not include mining dilution or recovery factors.
Base metal mineral resources are estimated based on the assumption that they would be processed at the Stall concentrator while gold mineral resources are estimated based on the assumption that they would be processed at the New Britannia concentrator, which is currently being refurbished.
Mineral resources were initially estimated using metal price assumptions that vary marginally over the assumptions used to estimate mineral resources at Lalor. In the QP's opinion, the combined impact of these small variations does not have any impact on the mineral resource estimates.
1901 mineral resources are estimated using a minimum NSR cut-off of $140 per tonne.

14.3. WATTS RIVER

Hudbay prepared resource model using Leapfrog® version 4.3 and MineSight® version 15, two industry standard commercial geological and mining software. The construction of this 3D resource model and the estimation of mineral resources were performed by Hudbay personnel following Hudbay procedures in compliance with best industry standards and the CIM guidelines. The work was supervised and validated by Olivier Tavchandjian, P.Geo., Vice President, Exploration and Geology.

Modeling Database

Drill holes totaling 24,757 m were included in the Watts River database to support the mineral resource estimate. Amongst these, 2,107 m were analyzed for, Cu, Au, Ag and Zn.

The drill hole database was imported in Leapfrog® and MineSight® from .csv files with a cut-off date for mineral resource estimate purposes of September 27, 2019.

Modeling of the Mineralized Envelope

Watts River mineralized envelope trend along an azimuth of approximately 20° with a general dip of approximately 65° to the east. Economic mineralization delineated so far occurs in two semi-massive to massive sulfide lens. These copper zones appear to be mostly aligned to the main regional deformation.

Units of overburden, barren and mineralized zones were created in 3D using core logging data and assay results. Two broad geological envelopes were then refined to create the mineralized envelopes.

The construction of the mineralized envelopes was based on the type of mineralization intersected. A smooth and continuous 3D solid was built using a 1 gram per tonne gold, a 0.5% copper or a 3% zinc cut-off to guide the construction of the mineralized envelope. Figure 14-7 present a plan view (main left), a looking North view (upper right) and a looking West view (lower right) of the mineralized envelopes.

In a final validation step, the envelope was loaded into MineSight® in order to ensure proper tagging of the solids to actual drillhole locations. The mineral envelope was used as hard boundary in all cases for grade interpolation purposes to correctly prevent spreading of sulfide mineralization into the barren zone and vice-versa.

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FIGURE 14-7: VIEW OF MINERAL ENVELOPE WIREFRAME

Density

Density values generated from two multi regression formulas are based on 197 density measurements contained within the main mineralized envelope: a formula for samples with iron assays results and a formula when samples do have iron assays results. The regression formulas were used to predict the density value where no actual measurement was taken.

In the case of the satellite mineralized envelope, only 10 samples had SG measurements. Nevertheless, a regression formula was developed to predict the density. Given the fact the coefficient of determination obtained was moderate (0.55 adjusted), it was deemed better to use the regression formula than use the average density for the zone. The predicted density values were validated through comparisons against the measured density.

Compositing

Assay intervals were regularized by compositing drillhole data within the interpreted mineralized envelope. Although most of the drillholes were assayed on lengths of less than 1 m interval (Reference Missing), a composite length of 2.5 m was selected as more appropriate to conduct interpolation into the 5mx5mx5m block size selected by the project evaluation team as the optimum selective mining unit (SMU). Grade compositing was weighted by specific gravity as there is a positive correlation between some of the metal grade and density. The compositing process was validated by comparing total length, density and length weighted average grade for each metal of the 2.5m composites to the original assays.

Exploratory Data Analysis

Exploratory data analysis (EDA) includes basic statistical evaluation of the composites for copper (Cu), gold (Au), zinc (Zn), silver (Ag) and density (SG). The EDA was conducted in the mineralized envelopes only. Within the mineral envelopes, the EDA also aimed to identify sub-domains that did not support the use of hard contacts for grade interpolation purposes but that justified being used for the purpose of block model validation, smoothing assessment and correction and resource classification.

Overall, the statistics of the composites indicate that the kriging interpolation will as expected introduce some level of smoothing and that a thorough smoothing assessment and post-processing of the initial interpolation is warranted to produce unbiased grade-tonnage curves and mineral resource and mineral reserve estimates.

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Grade Capping

The deciles analysis (Parish) method was used to define high-grade outliers and to assess the need for grade capping. This analysis was conducted on the 2.5m composites in the mineralized envelopes.

This method considers capping when the last decile of the population contains more than 40% of the metal and that the last percent contains more than 10% of the metal. In all cases, it was found that capping would not be required for the two base metal zones.

Variography

Variogram analysis was perform with MineSight using a combination of nugget and two nested structures calculated along the envelope plan.

Grade Estimation and Interpolation Methods

The block model consists of regular blocks (5 m along strike by 5 m across strike by 5 m vertically) coupled with an ore percentage model. The mineral envelopes were used to assign the percentage of the block that belongs to each unit.

Grade interpolation used a strict composite and block matching system based on the mineralized envelopes codes. For example, only composites coded as LITHO=1 (i.e. the main mineralized envelope) were used to estimate the grade of the block that was assigned to ORE1=1 and ORE2=1. Both nearest neighbour (NN) and ordinary kriging (OK) grade interpolations were completed, using three passes with increasing minimum information requirements (Table 14-8). The search passes were selected to ensure best local estimates recognizing that OK has a smoothing effect but making no attempt during interpolation to reduce this smoothing as it would negatively impact the quality of the local estimates.

TABLE 14-8: SEARCH ELLIPSE PARAMETERS

The NN interpolation is solely used to validate that there is no global bias and to assess the level of grade smoothing in the OK interpolation for the selected SMU dimension.

Grade Estimation Validation

Visual checks of appropriate honouring of the input data but acknowledging that some natural smoothing should occur between samples as the grade of a sample in the middle of a block is not the average grade of the block;
Absence of global bias by comparing the mean grade estimated by kriging to the original composite average grade and to a declustered grade obtained from a nearest neighbour interpolation;

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Volume variance correction to assess and correct if need be the smoothing effect of the grade interpolation

Visual Inspection

Visual inspection of block grade versus composited data was systematically conducted in section view. This check confirmed a good reproduction of the data by the model.

Global Bias Checks

This validation step consists in comparing the global average grade of each element between the original composites, the kriged block estimates and the nearest neighbour estimates.

A global check was performed to verify that the kriged mean block estimate did not present any bias when compared to the composites and the nearest neighbour model. Differences between the 2.5m composites, the NN and OK grades are acceptable.

Smoothing Assessment

The visual validation conducted in sections confirmed that the block grade interpolation was consistent with the supporting composite data. The larger number of composites used for grade estimation in the block model significantly improves the individual block grade estimates but, at the same time results in a much smoother model requiring a careful assessment and, in many cases, a post-processing of the OK estimates.

The extent of grade 'over-smoothing' in the model was investigated in the main mineralized envelope based on material differences in grade distribution and/or drilling density. The mean and variance of the kriged estimates were compared to the variance of the composites after declustering. The expected true variance between SMUs was calculated from the variogram models. Smoothing assessment was not conducted on the satellite given the fact that the low number of composites did not generate valid variograms.

Smoothing Correction

An indirect log-normal correction was used to perform a change of support on the kriged models in order to obtain unbiased grade tonnage curves. This correction is only valid globally and provides poorer local estimates than the smoothed OK model but does not materially alter the global average grade within each zone and provides the correct grade-tonnage curve for the variogram models fitted on the drillhole data. It is an appropriate method to predict the recoverable tonnage and grade such as the volume mined over three months of production which should be a realistic aim for a long-term reserve model based on exploration drilling. The targeted variance was reached within very close limits in most cases (Table 14-9).

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TABLE 14-9: SUMMARY OF SMOOTHING CORRECTIONS - WATTS RIVER

Classification of Mineral Resource

During the interpolation process, a number of control parameters were recorded for each block, e.g. number of samples, number of holes, the distance to the nearest sample and the average distance to all the samples used for the interpolation as well as the number of quadrants with samples, the kriging variance and the regression slope of kriging for each individual block estimate.

The regression slope values obtained from the kriging of copper grade estimates was used as the primary criteria for resource classification with 80% and 60% regression slope thresholds used respectively to separate measured from indicated and from inferred resources. From Hudbay's experience, this criteria was found to be a reliable first pass measure of quarterly and annual performance in tonnes and grade prediction.

Usually, this block by block coding assignation would have been smoothed to remove isolated blocks of one category within another. However, only a limited amount of blocks within ORE1 located at the immediate vicinity of drillhole intervals had regression slopes higher than 0.6.

Given the relative lower level of confidence caused by the limited amount of drilling, all blocks contained within the mineralized envelope were classified as inferred.

Reasonable Prospects of Economics Extraction and Mineral Resource Estimates

As for Lalor, the mineral resource estimates for Watts are reported within the constraint of an optimized stope-shape. The parameters used as input to define the mineable shape covers all the relevant technical and economic constraints including minimum stope and waste pillar dimensions and the NSR value calculation are very similar to those used at Lalor for the base metal lenses assuming beneficiation at the Stall mill. The mineral resource estimates were reported to ensure that each potential stope would cover all its associated operating and capital costs. The resulting mineral resource estimates are presented in Table 14-10.

TABLE 14-10: WATTS MINERAL RESOURCE ESTIMATES

Notes:

Totals may not add up correctly due to rounding.
Mineral resources that are not mineral reserves do not have demonstrated economic viability.
Mineral resources in the above tables do not include mining dilution or recovery factors.
Base metal mineral resources are estimated based on the assumption that they would be processed at the Stall concentrator.

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Mineral resources were initially estimated using metal price assumptions that vary marginally over the assumptions used to estimate mineral resources at Lalor. In the QP's opinion, the combined impact of these small variations does not have any impact on the mineral resource estimates.
Watts mineral resources are estimated using a minimum NSR cut-off of $150 per tonne, assuming processing recoveries of 90% for copper, 80% for zinc, 70% for gold and 70% for silver.

14.4. PEN ZONE

The following summarize the method used for the Pen II mineral resource model developed by Hudbay's in 2018. A more detailed description on how the model was built was published in Hudbay 2019 Technical Report and is available in SEDAR.

Modeling Database

95 drill holes were used to support the mineral resource estimate with a cut-off date of August 29th, 2018. The files were imported as collar, downhole survey and assay data into MineSight® and Leapfrog® for independent comparison and validation. No significant differences in the projection of the drillhole traces were found between the two datasets.

Modeling of the Mineralized Envelope

The Pen zone mineralized envelope trends along an azimuth of approximately 40° with a general dip of approximately 40° to the north east. Economic mineralization delineated occurs in a semi-massive to massive sulfide lens typical of VMS deposits.

A mineralized envelope was created in 3D using core logging data and assay results. This broad envelope of sulfide occurrence was then refined to create a narrow smooth and continuous 3D solid including most of the zinc rich mineralization.

Density Assignation

Density values were calculated using a linear regression formula from the zinc and iron content based on 166 measurements and supported by a coefficient of correlation of 97%.

Compositing

Due to the narrow nature of the mineralization, the grade interpolation process was simplified to a pseudo-2D approach. Assay intervals were regularized to a single full length composite from hanging wall to footwall. This modeling approach was adopted to prevent any lateral selectivity in the mine design that is not deemed realistic at PEN II. Grade compositing was weighted by specific gravity as there is a significant positive correlation between some of the metal grade and density in the mineral envelope. After compositing, the total length, the density and length weighted average grade for each metal of the composites was compared to those of the assays before compositing to ensure they were similar.

Variography

Down-hole and directional variograms for Zn, Au, Cu, Ag, and SG were created using the MineSight data analysis software. The major, semi-major and minor axis were built in order to fit the attitude of the mineral envelope. A linear combination of a nugget and at two nested spherical models were adjusted. Once generated, a systematic visual check was conducted to ensure that the search ellipsoid would be correctly oriented with respect to the geometry of the mineral envelope.

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Grade Estimation and Interpolation Methods

Both nearest neighbour (NN) and ordinary kriging (OK) grade interpolations were completed, using three passes with increasing minimum information requirements. The composite selection parameters for grade estimation in each domain were selected to ensure best local estimates. The percentage of the blocks contained within the mineralized envelope is recorded in the model.

The first interpolation pass uses 150% of the variogram ranges and is restricted to a minimum of one composite and a maximum of eight composites. The second interpolation pass uses 75% of the variogram ranges and is restricted to a minimum of three composites and a maximum of eight composites. The third interpolation pass uses similar constraints to the second pass but with a reduced 50% of the variogram ranges. At PEN II, all the composites have a different length and as a result the interpolation was both length and density weighted.

Grade Estimation Validation

The grade estimation process was validated to ensure appropriate honouring of the input data and unbiased resource reporting and use of the model for reserve estimation.

Visual inspection of block grade versus composited data was systematically conducted in section and plan views. This check confirmed a good reproduction of the data by the model.

A global check was performed to verify that the kriged mean block estimate was located between the mean of the composites and the mean of the nearest neighbour model.

Classification of Mineral Resources

As for Lalor, the regression slope values obtained from the kriging of zinc grade estimates was used as the primary criteria for resource classification. Blocks from the second and third interpolation pass combined with a 60% regression slope thresholds were used to separate indicated from inferred resources. There are no measured resource estimates reported at PEN II. The block by block coding assignation was then smoothed to remove isolated blocks of indicated resources within areas of mostly inferred resources and vice versa.

Reasonable Prospects of Economics Extraction and Mineral Resource Estimates

As for Lalor, the mineral resource estimates for PEN II are reported within the constraint of an optimized stope-shape. The parameters used as input to define the mineable shape covers all the relevant technical and economic constraints including realistic minimum stope and waste pillar dimensions. The NSR value calculation is based on the one used at Lalor for the base metal lenses assuming beneficiation at the Stall mill. The mineral resource estimates were reported to ensure that each potential stope would cover all its associated operating and capital costs. The resulting mineral resource estimates are presented in Table 14-11. There are no mineral reserve estimates reported for the PEN as the required engineering work has not been completed yet.

TABLE 14-11: PEN II MINERAL RESOURCE ESTIMATES

Notes:

Totals may not add up correctly due to rounding.
Mineral resources that are not mineral reserves do not have demonstrated economic viability.
Mineral resources in the above tables do not include mining dilution or recovery factors.

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Mineral resources were initially estimated using metal price assumptions that vary marginally over the assumptions used to estimate mineral resources at Lalor. In the QP's opinion, the combined impact of these small variations does not have any impact on the mineral resource estimates.
Pen II mineral resources are estimated using a minimum NSR cut-off of $75 per tonne and assume that the mineral resources would be amenable to processing at the Stall mill.

14.5. WIM & 3ZONE

In May 2015, Golder Associates ("Golder") prepared a resource model for the Wim property using Datamine® version 2.1.1547.0, an industry standard commercial geological and mining software. The same modeling approach and software were used by Golder in 2019 to construct a resource model for the 3 Zone gold deposit, a satellite ore body to the former producing New Britannia mine.

At both deposits, a smooth and continuous 3D mineralized envelope was built using a copper equivalency cut-off grade in the case of WIM and a gold cut-off grade in the case of 3 Zone to guide the construction of the mineralized envelope while including material below cut-off where necessary to maintain the continuity of the envelope. Scatter plots were used to identify assay outliers for grade capping. In the case of WIM, missing density values were generated using a linear regression formula from either the Fe grade or from the Cu grade. A composite length of 0.9 m in the case of WIM and 1m in the case of 3 Zone were selected to conduct interpolation into the blocks. Compositing was weighted by density at WIM but not at 3 Zone due to the absence of correlation in the latter case between density and gold grade.

A block size of 15 m east-west (strike), 3 m north-south (thickness), and 5 m elevation was selected at WIM while a block size of 5 m east-west, 5 m north-south, and 2m elevation was selected in the case of 3 Zone. At WIM, grade for Cu, Zn, Au, Ag and density (SG) was estimated using an inverse distance interpolant (IDW2) in the unfolded space and weighting by density while only gold grade was interpolated for 3 Zone using ordinary kriging (OK). In both cases, a nearest neighbor (NN) interpolation was also performed to provide declustered sample grade for block model validation. The search ellipsoids were defined from pairwise variograms modeled by a linear combination of a nugget and nested spherical structures.

The information provided by Golder was validated to ensure the built quality matched Hudbay technical standards.

Resource classification was assigned based on the search parameters used for grade interpolation. The block by block coding assignation was then smoothed to remove isolated blocks of indicated within areas of mostly inferred category and vice versa. At both WIM and 3 Zone, the resource models have supported pre-Feasibility studies and no resource estimates are reported in addition to those that been included in a viable mine plan and converted to reserve estimates.

14.6. NEW BRITANNIA MINE AND BIRCH DEPOSIT

The mineral resource estimate performed by Kinross followed a conventional and industry standard approach. The approach and general methods of resource estimation are summarized as follows:

1) Sectional interpretation of mineralization zones at a cutoff grade of 0.097 oz. per ton and a minimum true width of 7 ft.

2) All assays greater than 0.50 oz. per ton were cut to 0.500 oz. per ton.

3) The sectional interpretation of the mineralization was digitized and wireframed, using Gemcom Software.

4) The assay data was composited into 3-ft. intervals and basic statistics were completed on the assay and composited assay data.

5) Geostatistical analysis was completed and variogram models established.

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6) A block model with dimensions 15 ft. x 15 ft. x 15 ft. was used for grade estimation.

7) Grade interpolation was completed using OK using a number of interpolation passes.

8) Validation of the block model using graphical and statistical techniques.

9) Tonnage was established using the percent model utility in Gemcom and tabulated by rock type using the percentage of the block within the mineralized wireframe.

10) The blocks were composited, using Gemcom, into 50-ft. levels and 50-ft. sections for tabulation and reporting of tons and grade.

11) Tabulation and reporting of the resources was completed at US$400, US$450, and US$500 gold prices corresponding to cutoff grades of 0.097 oz. per ton, 0.086 oz. per ton, and 0.077 oz. per ton.

The review of the reports and the two procedures verify that the Kinross resource estimate, (audited and reported in the Micon International Ltd. NI-43-101 technical report dated October 2006), is a consistent representation of the drill data.

The cutoff grades for the resource have been estimated over a 6-ft. minimum true width and the parameters described in Table 14-12. The formula to calculate the cutoff grade is:

costs/(gold price*recovery)*31.103/exchange rate

TABLE 14-12: COST AND RECOVERY PRICE ASSUMPTION

Table 14-13 presents the resource estimate for the New Britannia mine and Birch property, as of January 1, 2021, using a variable cut off grade. Given its historical nature, the entire resource is classified as inferred. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

TABLE 14-13: MINERAL RESOURCE ESTIMATE FOR THE NEW BRITANNIA MINE AND BIRCH DEPOSIT ^(1^^)(2)(3)^

Totals may not add up correctly due to rounding.
Mineral resources that are not mineral reserves do not have demonstrated economic viability.
Mineral resources in the above tables do not include mining dilution or recovery factors.

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WSP reviewed all the available data, modeling procedures, block models and technical reports and more specifically, audited the geological model, mineralization wireframes, domain statistical validation, and visual trend slice validation. Available historical data and technical reports were also examined. Based on the independent checks conducted by WSP, the QP is satisfied that the procedures followed by Kinross and Alexis Minerals were adequate to support the declaration of inferred mineral resource estimates.

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15. MINERAL RESERVE ESTIMATES

No mineral reserve estimates are reported for the Pen II, Watts and New Britannia Mine deposits and as a result, this section documents how the joint LOM plan and mineral reserve estimates were developed for the Lalor mine, 1901, WIM and 3 Zone deposits only. The Lalor mine, 1901 deposit and its satellite deposits mineral reserves as of January 1, 2021 are summarized in Table 15-1.

TABLE 15-1: LALOR MINE AND SNOW LAKE DEPOSITS MINERAL RESERVE ESTIMATES (AS OF JANUARY 01, 2021)

Notes:

Totals may not add up correctly due to rounding.
Metal prices of US$1.10 per pound zinc, US$1,500 per ounce gold, US$3.10 per pound copper, and US$18.00 per ounce silver with an exchange rate of 1.30 C$/US$ were used to confirm the economic viability of the mineral reserve estimates. Mineral reserves are estimated using a minimum NSR cut-off of $105 per tonne for waste filled mining areas and a minimum of $116 per tonne for paste filled mining areas.
For Lalor, individual stope gold grades were capped at 10 g/t, as a prudent estimate until reserves to mill reconciliations can establish that the high-grade gold can indeed be entirely recovered. This capping method resulted in the reduction of the global gold reserve grade by approximately 3%.
WIM mineral reserves are estimated assuming processing recoveries of 98% for copper, 88% for gold, and 70% for silver based on processing through New Britannia's flotation and tails leach circuits.
3 Zone mineral reserves are estimated assuming processing recoveries of 85% for gold based on processing through New Britannia's leach circuit.

The Deswik mine design software was used to optimise the mine plan in all cases based on cut-off costs, stope geometry parameters, and mineral resource shapes from the resource model described in Section 14. Appropriate dilution and recovery factors were applied based on mining method with a combination of paste and unconsolidated waste backfill material.

The mineral reserves were sequenced and scheduled into an integrated LOM schedule using Deswik interactive scheduling software and exported to spreadsheets for financial analysis.

The author considers that the mineral reserves as classified and reported comply with all disclosure in accordance with requirements and CIM definitions. The author is not aware of any mining, metallurgical, infrastructure, permitting or other relevant factors that could materially affect the mineral reserve estimate.

There are no legal, political, environmental or other risks that could materially affect the potential developments of the reserves in the report.

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15.1. MINING STRATEGY AND RESERVE ESTIMATIONS PROCESS

Mining Method Selection & Mining Strategy

Lalor Mine has been in production continuously since mid 2012 and has produced 8.8 million tonnes by the end of 2020. Mining methods are the major determinants of dilution, recovery, economics, and production scheduling, which are used to determine mineral reserves. Production at Lalor is planned to be done using predominantly transverse longhole mining with minor longitudinal retreat and cut and fill mining.

Early in the mine life, the complexity of stacked zinc and gold zones within the deposit was realized. In the zinc rich 10 and 11 Lenses, where there is no lens stacking, longhole retreat, transverse longhole and post pillar cut and fill mining were used successfully. See Figure 15-1 for typical mining arrangement in 10 Lens.

Selective post pillar cut, and fill mining was initially used in 20, 21, 25 and 32 lenses as shown in Figure 15-2. Post pillar mining allowed geological evaluation of each round, leading to a better understanding of the ore. This helped to determine: 1) the possibility of mining the 20 and 32 lens zinc ores separately from 21 and 25 lens gold ores and 2) the ability to visually identify and segregate gold ores from surrounding waste. The result was that most zinc ore stopes can be designed and mined separately from the gold, but that the microscopic nature and distribution of the gold in gold zones required additional diamond drilling, sampling and assaying information than required for the zinc stopes to determine production grades.

FIGURE 15-1: 840 LEVEL 10 LENS PLAN VIEW AT LALOR

Updated geological interpretations and the decision to construct a paste backfill plant supported the decision to convert some areas of the mine similar to those in Figure 15-2 from post pillar cut and fill to longhole mining across all lenses. For narrow and steeply dipping stopes typical of 31 lens, longitudinal retreat mining was preferred, while for wide stopes crossing multiple lenses, typical of 20, 32 and 21 lenses shown in Figure 15-2, transverse longhole mining using footwall drawpoints was preferred.

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FIGURE 15-2: 910 LEVEL POST PILLAR C&F AT LALOR

This mining method along with paste backfill also allows the division of stopes into hanging wall and footwall blocks, further reducing the risks and effects of stope failure. This approach was also selected for the upper areas of Lenses 10 and 27 and for the lower part of Lens 27.

The flexibility provided by transverse longhole mining has provided consistent production of 4,500 tonnes per day and will allow concurrent production from both zinc and gold stopes. In addition, commissioning of the paste plant in the third quarter of 2018 has significantly reduced the time to backfill a stope, provides tight competent fill up against the hanging wall and a noticeable improvement to ground conditions. In addition, paste backfill allows greater flexibility in mine design and sequencing options.

The strategy for this update of the mine plan at Lalor was to mine base metals of higher grade zinc production in 2021 while at the same time advancing the development required to start mining the gold and copper gold zones. Following the planned opening of the New Britannia mill in Q1 2022, the strategy changes to targeting a minimum production of 1,500 tonnes per day of gold ore until depletion of the mineral reserves.

Mining Method Selection & Mining Strategy for the Future Operation at the 1901, WIM and 3 Zone Deposits

The mine design and underground infrastructures at 1901 have been based on current practices at Lalor, to maintain commonality with equipment, materials, operating procedures, and workforce skillsets. The primary mining methods will be variations of cut and fill, as based on the local geometry of the resource. Post-pillar cut and fill will be used where the resource has sufficient vertical thickness to mine multiple vertical cuts (overhand), and inclined room and pillar will be used where there is limited vertical thickness and the resource will be mined in a single cut.

1901 will be accessed from the existing Chisel portal/ramp and Lalor ramp as shown in Figure 15-3. There will be a 5 m wide x 5 m high service ramp and a 5 m wide x 5 m high haulage ramp. The ramps also serve as the main ventilation circuit, with fresh air sent down the haulage ramp and return air directed up the service ramp. These two ramps also provide second egress from 1901. The ramps will approach 1901 from the hangingwall side of the deposit, however mine infrastructure will be in the footwall of the deposit.

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FIGURE 15-3: ACCESS TO 1901 DEPOSIT ISOMETRIC VIEW

1901 is expected to produce 1,000 tonnes per day of ore at steady state and has been incorporated into the Lalor production strategy as an extension to the existing mine. Personnel and equipment from 777 and Lalor are to be redeployed to 1901 in future years, to sustain peak production levels at the mine and mills.

At WIM and 3 Zone, longitudinal longhole mining with unconsolidated rock fill is the chosen method for the underground mines. To eliminate the use of cemented back fill, pillars in ore will be left for stability purposes. Each deposit will be mined as a ramp access mine with limited infrastructure, with 3 Zone using the existing portal and ramp development. The existing ramp at 3 Zone will be mucked out and refurbished and new ramp development started once the bottom of the decline has been reached. To keep capital costs low, execution of 3 Zone will begin once WIM beings to ramp down, thus allowing some infrastructure to be moved and installed at 3 Zone (pumps, temporary fans, electrical, main fans, heaters, portable refuge stations and equipment).

The production rates over the LOM are 840 tonnes per day (8 years) for WIM, and 450 tonnes per day (4 years) for 3 Zone, with the ore being processed at the New Britannia Mill once Lalor ore is depleted.

Reserve Methodology

The following steps were followed in developing the reserve estimates:

Calculate two payable (NSR) values for each individual block in the resource model depending on whether processing would occur at the Stall concentrator or at the New Britannia concentrator, using long-term metal prices, concentrator recoveries, metal payability and downstream smelter treatment and refining costs assumptions. Supporting test work, actual performance and forecasts supporting these assumptions are discussed in details in Sections 13, 17, 19 and 21 of this Technical Report.
Design stopes in the Deswik Stope Optimizer, considering depleted mineral resources, existing workings, resource categories and mine and mill operations costs. Dilution and recovery are estimated and applied at this step. Stopes are designed for both the Stall concentrator option and the New Britannia concentrator option
Establish stope economics using a secondary NSR calculation where, along with mine and mill operations costs, mine capital, waste development and offsite administration costs are applied to each stope. Assign whether stopes can be upgraded to mineral reserves based on resource classification.

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Design ore development required for mining the reserves. Deplete development from the stopes. Interrogate grades of designed development for inclusion in mineral reserves.
Sequence and schedule development and stope production for input to a financial Life of Mine (LOM) study to support mineral reserve economics.

The above methodology takes into consideration the three different ore types and the milling options for the future production from Lalor mine.

15.2. DILUTION AND RECOVERY

Planned dilution factors are input in the stope optimizer by mining method assuming those currently experienced at Lalor Mine for transverse longhole open stope, longitudinal retreat longhole and post pillar cut and fill mining.

Dilution

Dilution is included in the conversion of mineral resource to mineral reserves and can be internal (planned) or external. Dilution is reported as the ratio of total waste divided by mineral resource.

Internal dilution is primarily incurred due to the shallow dipping nature of the deposit and stacking of lenses which can result in multiple lenses being grouped together for mining purposes. The space between the lenses that are mined in the course of stope mining is considered as internal dilution. In the Deswik stope optimizer routines multiple lenses may be required to be extracted as a single mining unit based on stope mining parameters as shown in Table 15-2.

Internal dilution included in mineral reserves can also include low grade development required to achieve footwall dip angles in longhole stopes and may include material mined between lenses where multiple lenses are grouped together. Current grade control practices will continue to be used to assign material as ore or waste on a round by round basis to attempt to minimize dilution.

TABLE 15-2: STOPING PARAMETERS BY MINING METHOD

Note: Lalor Cut & Fill length along strike (min and max) values represent a collection of drifts whereas the 1901 values represent the size of each drift.

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At Lalor, external dilution is set at a fixed distance of 1.0m into the hanging wall after the stope geometry shape is finalized. Internal dilution and external dilution are included as part of the optimized mining shape. If dilution occurs within the block model, it is assigned block model grades. Otherwise, it is set at zero grades and a bulk density of 2.8 t/m3. Dilution incurred in converting resources to reserves is summarized by mining method in Table 15-3.

TABLE 15-3: DILUTION INCURRED BY MINING METHOD

Mining Recovery

Mining recovery is defined as the ratio between mineral resource tonnes delivered to the concentrator and in-situ mineral resource tonnes estimated in the model. Mining recovery assumptions used by type of stoping and backfilling method in the Deswik Scheduler are shown in Table 15-4. These assumptions were based on a combination of first principles and actual performance at the mine since the start of operations. Resultant recovery factors by mining method are shown in Table 15-5. Some of the mineral resources not recovered relate to: rib, post and sill pillars required to maintain rock stability, inefficiencies in mining represented by small blocks of ore along ore/waste contacts and underbreak and inefficiencies in mucking reflecting ore unable to be mucked from longhole stopes.

TABLE 15-4: RECOVERY FACTORS BY MINING METHOD

TABLE 15-5: AVERAGE RECOVERY FACTORS BY MINING METHOD

15.3. DESWIK STOPE OPTIMZER

Stopes were built in the Deswik stope optimizer considering measured, indicated, and inferred mineral inventories from the block model. Stope construction considered the NSR block values and appropriate dilution and recovery factors based on resource geometry, mining and backfill methods.

To reduce the possibility of including low grade material in stopes, cut-off costs attempt to replicate the steady state cost of mine and mill operations, and include ore extraction, onsite mine G&A and mill operating costs. This generates stopes that might be mined economically during the normal operation of the mine. Cut-off costs ($/tonne) used to construct stopes in the stope optimizer are shown in Table 15-6.

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TABLE 15-6: CUT-OFF COSTS

Note: with Lalor being an operating mine, a second test is conducted to only retain the stopes that can cover the sustaining capital costs associated with their extraction, which is on average $34.61/t for the Lalor mine.

Iterations were run considering mining methods and preferred milling location. At this stage, the multiple runs of the stope optimizer continued to consider measured, indicated, and inferred resources. The sequence used to build the preliminary Deswik stope inventory is:

1) Attempt to build longhole stopes in the stope optimizer using payability from the flotation only milling process (NSR Stall) assuming longhole mining and Stall milling cutoff costs.

2) Where longhole stopes were unable to be built by the stope optimizer, or where excessive dilution was designed into the stope, attempt to design post pillar cut and fill stopes using the stope optimizer, using higher cut and fill mining costs which incurs less planned dilution and the Stall milling cutoff cost.

3) For gold lenses that are distinctly separated from zinc lenses, attempt to build longhole stopes using payability and costs from the New Britannia concentrator. This assumes higher gold payability and higher mill operating costs.

4) For gold lenses that are distinctly separated from zinc lenses and for which longhole stopes cannot be designed for, or for areas incurring high dilution, attempt to design post pillar cut and fill stopes using payability from New Britannia concentrator and using higher cut and fill mining costs, which incurs less dilution.

At 1901, WIM, and 3 Zone, the in-situ mineral inventory values helped determine which mill the ores will be routed to, and so multiple optimization runs based on either the NSR ZN or NSR AU fields were not necessary for each lens or zone. For these satellite deposits, an elevated cut-off value was applied, with the intention of capturing the expected sustaining capital costs, to ensure that the resulting mineable shapes would pass further economic testing and remain profitable.

15.4. STOPE NET SMELTER RETURN CALCULATION

To determine if stopes could be included in the final mine plan and in the mineral reserve estimates, only those stopes supported by measured and indicated resource estimates were retained. Then, the tonnes and grades contained in the DSO stopes were subjected to a secondary financial test using 2021 budget unit costs. Each stope was assigned a mill circuit to determine stope revenues, and a mining method and backfill method to determine mining costs. Direct mining costs derived from mining and backfill method, onsite G&A costs, offsite administration costs, mine sustaining capital costs and stope specific waste development costs were applied on a stope by stope basis to determine final stope economics:

Mining, milling and onsite G&A costs are based on the 2021 budget and are applied to each stope generated in the DSO as shown in Table 15-6
Capital equipment costs were applied at $8.83/t to all stopes generated in the DSO.

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Capital Major Installations were applied at $4.11/t to all stopes generated in the DSO.
Capital waste development includes ramp and level development metres required to access new mining areas:
Ramp and ancillary development (remucks, sumps, etc.) metres were estimated and applied to mining areas at 2021 budget unit costs.
Haulage drift metres (footwall drifts) were estimated based on the stope arrangement and were applied as required to access additional stopes along strike at 2021 budget unit costs.
Stope specific operating waste development includes longhole stope cross cuts and cut and fill entrance cross cuts. Estimated metres were applied on a stope by stope basis at 2021 budget unit costs.
Offsite administration costs were applied at 2021 budget unit costs and are based on copper and zinc concentrates produced from stopes.

For 1901, WIM, and 3 Zone, pre-feasibility study (PFS) level mine designs, development and production schedules, as well as capital and operational cost estimates were prepared by Stantec (1901) and BBA (WIM and 3 Zone). The key parameters established from these studies were used as input into a consolidated long-range business plan of the Manitoba Business Unit to further confirm the positive economic contribution of each.

15.5. ORE DEVELOPMENT

The final step to establishing mineral reserves was to create ore development shapes required to mine stopes. Planned 3D development shapes were built and removed from stopes in Deswik when creating Interactive Scheduler activities. The development shapes were interrogated and assigned as mineral reserves.

15.6. RECONCILIATION OF RESERVES TO PRODUCTION

Recent reconciliation between reserve estimates and past production at Lalor provide a useful insight into the reliability of the LOM plan to predict future production

The reconciliation results presented in this section encompass 100% of the mining production for the first three quarters of 2020 including long-hole mining, post-pillar cut, and fill and development mined as ore. The mine's production was reconciled to the mill and compared to the reserve estimates for the stopes and cuts entirely mined during this period.

Results presented in Table 15-7 below show that all tonnes and metal grade reconcile within 6%. The 6% difference in gold grade may actually be closer to 5% as 20 thousand tonnes of unaccounted high grade gold production reported to underground stockpiles in preparation for the New Britannia mill start-up in 2021 and is not credited to the 2020 mine's production. Although these results are considered very good overall, they also show that Lalor is still in the process of adapting its past grade control protocols which were previously focused towards maximising base metal production. The mill credited 10% more copper and zinc metal production than the mineral reserve estimates while gold was 2% lower. This is recognised by the mine with on-going initiatives aimed at improving and re-balancing the focus of the underground grade control program to ensure gold mining recovery is maximised going forward including closer drillhole spacing prior to extraction, assaying all blasthole samples to optimize blast design as well as systematic bucket sampling from production. With all these initiatives in place, we intend to maintain a very reliable reserve model and identify further opportunities to maximize gold extraction.

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TABLE 15-7: RESERVE MODEL DEPLETION RECONCILED TO MINE'S CREDITED PRODUCTION BY THE MILL

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16. MINING METHODS

Geology, geotechnical information, orebody geometry, productivity and mining experience are major drivers to select the best suited mining methods including considerations for lateral development, production, backfilling and ore transportation.

The hanging wall and footwall rocks at Lalor are generally of good quality, allowing the use of mechanized drilling and blasting techniques. The mineralized lenses dip at an average of 30°, but locally varying from sub-horizontal to 55°. Mining methods currently in use include cut and fill and longhole open stope. Paste backfill is used to increase recovery and accelerate the mining cycle. Low grade areas will be filled with rock from waste development.

Ore is mucked using Load Haul and Dump (LHD) loaders which are operated remotely in inaccessible areas. Ore is loaded into underground haul trucks or ore passes and transported to the ore handling system at the production shaft for hoisting to surface. Ore delivered to the production shaft is sized to less than 0.55m by one of two rock breakers. Ore is hoisted from the mine by two 16 tonne capacity bottom dump skips in balance. On surface, ore is truck hauled to a primary crusher at the Chisel North mine site, crushed to less than 0.15m, and then trucked to either the Stall or New Britannia concentrator for processing.

For 1901, two mining methods will be used. The primary method will be Post-Pillar Mechanized Cut and Fill (PPMCAF) used where the resource has sufficient vertical thickness to mine multiple consecutive cuts. Where there is limited vertical thickness and the resource can be mined in a single cut, Inclined Room and Pillar will be used. The PPMCAF stopes will be backfilled with unconsolidated waste rock.

At WIM and 3 Zone, longhole retreat mining methods were selected to excavate the orebodies. Utilizing this method will require a bottom-up mining sequence with mining fronts progressing in from the orebody extents towards the center of the orebody. The orebodies have been divided into various mining blocks to allow multiple levels to be extracted simultaneously.

16.1. LATERAL DEVELOPMENT

At Lalor, lateral development is horizontal or near horizontal tunnelling, travelable by rubber tired mechanized equipment. Ramps are typically mined at up to +/- 15% grade to provide access between levels. Rounds are 4m long and typical dimensions are 6m wide by 5m high. Standard ground support, consists of resin grouted rebar and welded wire mesh across the back and to within 1.8m of the sill on the walls and face. Mine services, including compressed air, process water and discharge water pipelines, paste backfill pipeline, power cables, leaky feeder communications antenna and ventilation duct, are installed in main levels and stope entrances.

For 1901, access has been started from the existing Lalor ramp and will approach the deposit from the hangingwall side. The main development profile will be 5 m wide x 5 m high, with dimensions established to allow for the required ventilation flow velocity as well as to meet other criteria like minimum equipment clearances. All ramp and lateral excavations will be developed using traditional drill and blast methods and diesel-powered mobile equipment.

At WIM and 3 Zone, all lateral ore and waste development is designed to accommodate Sandvik LH514 and LH517 LHDs and Sandvik TH550 trucks (or equipment of equivalent size). Levels generally consist of sills in the ore with level access off the main ramp to the sill. Sumps, as well as the access drifts for the return air and fresh air raises, will be off the level access.

16.2. VERTICAL DEVELOPMENT

At Lalor, main ventilation raises and ore pass raises are developed by a mining contractor using a raisebore and/or Alimak climber. Ground support and ladder ways are installed as required. Longhole slot raises, transfer ore passes, and auxiliary ventilation raises are limited to approximately 35m long and are developed using longhole conventional drop raise techniques. Drain, paste backfill and electrical cable holes are drilled using longhole or raisebore drills and are reamed to designed diameter.

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For 1901, major vertical development will consist of an existing 3m diameter fresh air raise (FAR) as well as a planned 3.5m diameter return air raise (RAR) that is to be raisebored roughly 500m in length, Additional internal raises will be used for ventilation and egress and they are planned to be developed with an Alimak method.

The requirements for vertical mine development at WIM and 3 Zone include FARs and secondary egresses. The main ventilation raises are sized based on ventilation requirements and will also be used as a second means of egress. It is envisioned that the raises will be developed using an Alimak method and slashed to the appropriate dimensions as the raise is supported and manway installed.

16.3. STOPE MINING

The two mining methods used at Lalor mine, 1901 deposit and the Snow Lake satellite deposits are variations of cut and fill mining or longhole open stope mining. In general, where the dip exceeds 35° and the orebody is of sufficient thickness, longhole open stope mining is preferred, whereas cut and fill mining methods are used in narrower, flatter areas. At Lalor, longhole mining represents 87% of total stope mining, and stope mining accounts for 92% of all mining, with development accounting for the other 8%.

Cut and Fill Mining

Single pass overhand mechanized cut and fill mining is the preferred method where the ore is flatter than 35° and horizontal widths are less than 10m. The ore is accessed from a footwall drift by a crosscut developed at approximately -15%. Ore is mined in 5m high horizontal cuts. Unconsolidated backfill is placed tight to the back and the entrance crosscut is then back slashed to provide access to the next cut.

Where ore widths are greater than 14m wide, mining is conducted by overhand post pillar cut and fill. Post pillars provide ground support to allow for selective mining of wide stopes. Backfill is placed to within 1.8m of the back and the entrance crosscut is back slashed to provide access to the next cut. Drifts and crosscuts in stopes are typically 7m wide x 5m high, with 7m x 7m post pillars. Mining retreats towards the access to reduce risks associated with poor ground conditions. A typical mining sequence is shown in Figure 16-1 as a plan view.

FIGURE 16-1: TYPICAL POST PILLAR CUT AND FILL MINING PLAN VIEW

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At 1901, in areas that undulate and are typically narrower, an inclined room and pillar method will be used. This method does not require backfill, as there is only one cut per level. A schematic demonstrating this method is shown in Figure 16-2, with the following sequence of steps followed for stope production:

Step 1 - develop the Ore Ramp from the Access Point to the extent of the resource. The Ore Ramp will traverse the ore at a gradient appropriate to keep the floor of the ramp at the FW contact.
Step 2 - develop Sill 1 along strike at the stope limit.
Step 3 - develop Sill 2 along strike.
Step 4 - mine the resource between pillars between Sill 1 and Sill 2, retreating from the stope extents to the Ore Ramp. The configuration of the pillars adjacent to the Ore Ramp will be adjusted to the geometry, while maintaining the planned extraction ratio.
Repeat Steps 2-4 for subsequent sills until mining has retreated to the initial access point.

FIGURE 16-2: TYPICAL INCLINED ROM AND PILLAR MINING PLAN VIEW

Longhole Open Stope Mining

Benefits of longhole mining over cut and fill methods include reducing exposure to risk as a non-entry method and higher productivity/lower cost than cut and fill methods.

The minimum design height of a longhole stope is 17m. In shallower dipping areas of the ore the interval is typically 17 to 20m, and in steeper dipping areas the interval is 20 to 25m. Stope widths assume a minimum mining width of 5m. Maximum open stope length and width is determined considering the rock quality and hydraulic radius of the combined exposed hanging wall and back span of the stope. A 5m rib pillar is added when unconsolidated rock is used as backfill.

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FIGURE 16-3: TYPICAL ISOMETRIC VIEW - TRANSVERSE LONGHOLE OPEN STOPPING

FIGURE 16-4: TYPICAL LONG SECTION - LONGITUDINAL RETREAT LH OPEN STOPING

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Teleremote and Autonomous Mucking

Teleremote and autonomous mucking allows stope mucking to and orepass to continue through shift change and blast clearing times. Lalor Mine uses Epiroc's Autonomous Mucking System which includes an operator's station on surface allowing control of up to three LHD's concurrently. The surface operator drives the LHD remotely to record a route and then switches to autonomous mode. The LHD travels the recorded route without operator input. Safety devices are installed on the mucking level to restrict access by mine workers and equipment during the operation. Future plans are to isolate truck routes from the stoping areas allowing teleremote loading of trucks as the mine expands below shaft bottom.

16.4. BACKFILL

All stopes at Lalor mine except uppers require backfill to maintain long term stability and to provide a floor to work from for subsequent mining. Backfill is either unconsolidated waste rock backfill (URF) or consolidated paste backfill. URF is used in stopes where pillar or wall confinement is not required and the value of the adjacent pillars does not warrant the added expenditure of paste backfill. Paste backfill is an engineered product comprised of mill tailings and a binder (3-5% cement by weight), mixed with water to provide a thickened paste that is delivered by borehole and pipes to stopes. Paste backfill presents several advantages over unconsolidated fill such as slurried mill tailings or loose waste rock including increased stability, better ventilation controls, higher mining recovery and lower mining dilution and it also eliminates potential flows of re-liquefied unconsolidated tails.

At WIM and 3 Zone, URF backfill material will come from underground development. The waste produced before stoping begins will be stockpiled on surface and hauled underground as required to replace removed ore. Once stoping begins, waste will be stockpiled underground for use as fill.

16.5. ORE HANDLING

Ore is mucked by LHD, loaded to underground haul trucks and hauled to one of the two rock breaker/grizzlies on 910m level for sizing to less than 0.55m. A 40m bin below the grizzly provides approximately 1,200 tonnes of coarse ore storage. A chute at the bottom of the bin at 955m level feeds ore to a conveyor that loads a measuring flask with approximately 14 to 16 tonnes of ore. Ore is skipped to surface by two 16 tonne capacity bottom dump skips in balance. The ore enters the headframe chute from the skips and is deposited into the surface ore bin or to the exterior concrete bunker via gravity. From the surface bin or bunker, ore is truck hauled to a primary crusher at the Chisel North mine site, crushed to minus150mm, and then trucked to a concentrator to process.

Ore from 1901 will be truck hauled and dumped onto an ore pad on surface near the Chisel North Portal. A trucking simulation was completed using Simio to determine the number of trucks required to meet the 1,000 tonnes per day maximum production rate envisaged. The simulation also includes backhauling 300 tonnes per day of waste rock in the ore trucks for use as backfill. The simulation was run with a 14.0 trucking-hour day with day and night shifts split evenly (7.0 hours of truck seat time per shift). The most favourable truck scenario to achieve 1,000 tonnes per day of ore haulage and backhaul 300 tonnes per day of waste rock is using two MT65 trucks.

At WIM and 3 Zone, all primary production mucking will be completed by LHD, with ore moving to a remuck bay located near the entrance of the level where it will be rehandled and side loaded onto trucks and hauled to surface. The ore will be stockpiled and hauled to New Britannia mill via surface haul trucks operated by a contractor.

16.6. GEOTECHNICAL DESIGN

A geotechnical database has been established from drill core logs and geotechnical core sample laboratory tests. Geotechnical data is continuously collected following standard procedures. This is done by the Lalor Geology department and data is shared with the Ground Control team. Adequate core samples are collected for all the ore zones to ensure we have sufficient geotechnical data from hanging wall, orebody and footwall to further test rock properties. Ground monitoring instruments (dCABLES) are used for ground movement monitoring and more advanced ground control instrumentation will be considered going forward.

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1901 rock mass characterization studies for geomechanical purposes and mine design have been based on basic geotechnical logging of resource holes, together with limited geomechanics laboratory testing, point load testing of full and half core, and evaluation of structural orientations from acoustic televiewer surveys. Rock mass conditions for the 1901 deposit can generally be described as very competent, massive to slightly jointed, with strong (50-100 MPa) to very strong (100-250 MPa) intact rock strengths.For WIM and 3 Zone, typical stope sizes and dimensions from existing Hudbay operations were considered practical for their respective studies. Although diamond drill holes were reviewed and indicate RQD values of High, future work is required to further identify structure and lithology of the mineralized zones and to characterize the rock mass, conduct rock mass behaviour analysis, and confirm ground support requirements.

16.7. GROUND SUPPORT SYSTEMS

Except when using cut and fill mining methods, drifts have arched backs to increase stability and factor of safety. Ground support is classified as either primary support or secondary support.

Primary support refers to reinforcement of the rock mass immediately following excavation to ensure safe working conditions before re-entry. Secondary support is additional support applied after the installation of primary support to provide further support in large spans, long term infrastructure excavations and structurally controlled areas where wedge failures may be a concern.

With 1901 currently in preliminary stages of project development, typical Lalor support for production levels and access were evaluated based on Q ratings (Barton, N.). For PFS level studies, it was confirmed that the Lalor standards for ground support are adequate for 1901.

Primary Support

The standard for Lalor is #6 or #7 2.3m (7'4") resin rebar installed on a 1.2m x 1.2m (4' x4') square pattern with a same size rebar in the center of the pattern. Screen (#6 gauge welded wire mesh) and bolts are installed on the walls, back and face prior to re-entry. At drift intersections, 3.6m (12') resin rebar in a 1.2m x 1.2m pattern with a center #7 7'4" rebar is used for primary back ground support.

16.8. UNDERGROUND DEVELOPMENT

To access the 1901 deposit, the service ramp and haulage ramp (twin ramps) will provide two working faces for a single jumbo, with additional ancillary faces available (connections between ramps, remucks, safety bays, electrical cut-outs, etc.). The initial priority will be to establish the ventilation and infrastructure at the start of the access, before proceeding toward 1901. Once development reaches 1901, the priority will be to establish a ventilation loop and open additional faces for a second jumbo crew to access the first production areas.

At WIM, a new ramp will be developed to access the underground workings. The ramp will extend from surface to the bottom of the mine, roughly 700m vertical meters below surface. The ramp is typically driven 5.3m wide x 5.5m high with an arched back and a 15% maximum gradient to accommodate the fleet.

The existing ramp at 3 Zone will be rehabilitated and used as an access and then extended down to the bottom of the mine. According to the provided wireframes it currently extends down roughly 130 vertical meters below surface, and the new ramp development will begin around 100 vertical meters below surface. The existing ramp portal is currently filled with rock fill, but the amount of fill used is unknown at this time. For the 3 Zone study, the assumption is that before development can begin, roughly 100m of the existing ramp will need to be mucked out and approximately 900m of rehabilitation will be required before development can begin.

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Vertical Development

Raises

Vertical development (or raises) is planned at Lalor mine and the Snow Lake satellite deposits for establishing or extending the ventilation and secondary egress systems. Drop raises used as slots in longhole stopes are included in stope drilling and blasting. Raises less than 35m can be mined as drop raises. Longer raises are evaluated to choose between raise boring and Alimak considering the conditions, end use and cost of both methods.

Paste Backfill, Drain and Service Holes

At Lalor, paste backfill is distributed to stopes via a series of steel cased vertical holes between levels and pipelines installed in footwall drifts. Paste backfill holes are drilled using a raisebore pilot bit. Service holes are holes drilled between levels for electrical cable distribution throughout the mine.

For 1901, WIM, and 3 Zone, allowances have been included in the cost estimates for drain holes and service boreholes.

16.9. DIAMOND DRILLING

Diamond drilling is done to define the extents of mineralization and is required for forecasting grades from stopes. The delineation holes are typically drilled from a drift located in the footwall or hanging wall. Stope definition holes are drilled from stope drawpoint stubs or undercut drifts for grade planning (which could vary sequence in some cases).

16.10. MINE DRAINAGE AND DEWATERING

Lalor mine is a relatively dry underground mine with no significant hydrological concerns. The drainage for Lalor mine and the past producing Photo Lake and Chisel North mines is interconnected. Mine water from all mine sources is pumped to the Chisel Open Pit, where it is pumped to a water treatment plant and then released to the environment. Main collection areas feeding the water treatment plant are: 140m level pump station at Photo Lake, a pumping station at Chisel North 540m level (referred to as the 990 sump) and satellite sumps at Lalor mine.

1901's mine drainage is to be integrated with Lalor systems. Service water from mining processes and inflows from the rock mass will flow in drainage ditches along the side of drifts and will be collected at each level sump, typically located near a ramp where possible. A main dewatering station on 580L will be outfitted with two filter weir sumps to filter slimes from the water and a clean water retention sump, which will pump to the water reservoir at Remuck No. 3 (in Lalor ramp). This main dewatering station will be established early so as the mine expands, all water from future workings will be collected and filtered prior to dewatering. Water will be directed to the main dewatering station via boreholes or staged through piping from levels below using submersible pumps.

Like Lalor and 1901, the WIM and 3 Zone dewatering systems are based on dirty water pumping achieved through a network of interconnected sumps, trenches, drain holes, and slurry pumps. Portable, standardized, skid mounted equipment to maximize ease of installation, flexibility, and reusability will be used. The pumping equipment includes pump stations, booster stations and clarifiers, and will be reused at 3 Zone once no longer required at WIM.

16.11. MINING OPERATIONS

Mining Fleet

Lalor mine is a ramp and shaft accessible mine with production and development done by rubber tired underground mining equipment. The mine equipment fleet required to achieve 4,500 tpd is shown in Table 16-1. Lalor mine production will be approximately 20% jumbo and 80% longhole. Additional contractor equipment not listed in Table 16-1 is planned to be on-site to complete the sustaining capital lateral development metre requirements in the LOM plan. An allowance for replacement of mine equipment has been included in the mine plan.

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The 1901 mobile equipment fleet will include units common with Lalor mine. There will be one jumbo and three scissorlifts available from the 777 mine, as well as one jumbo, two bolters, and two 65 tonne haul trucks available from Lalor. Although some auxiliary equipment may also be shared with Lalor, the remaining units are planned to be purchased by Hudbay.

WIM's underground fleet will be purchased new from suppliers and transferred to 3 Zone once available.

TABLE 16-1: MINE MOBILE EQUIPMENT

Mine Power

Grid electricity is supplied by Manitoba Hydro, the provincial power utility. Manitoba Hydro's 115 kV power line terminates at the Chisel North mine site, approximately 3.5 road km from the Lalor mine site. This supplies power to the Hudbay owned main distribution substation consisting of two (2) 115-25 kV 24 MVA transformers.

At 1901, the underground electrical power distribution system is to be fed from Lalor's 825mL transformer.

The WIM and 3 Zone mine electrical and power distribution originates at the New Britannia Mill site. Two primary feeders have been proposed to supply the load of the underground mine plants, with both primary feeders being egressed out of the mine services substation via an underground concrete encased ductbank (or direct buried), to ensure that the risk of contact with electrical plant is minimized. The overhead distribution will be constructed alongside the existing gravel road and will consist of wood poles and wires.

Production Schedule

The Life of Mine (LOM) production schedule combining Lalor, 1901, WIM and 3 Zone is shown in Table 16-2: Mine Production Schedule to Table 16-5. The Deswik software was used to assist with the LOM sequencing and scheduling to generate the production schedule. Historic mining rates and systems, equipment or crews capabilities were applied and levelled. From this output, adjustments were made to further balance the capabilities to create the final plan. Concentrates produced from Lalor and contained metal in concentrates are also shown in tables bellow.

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TABLE 16-2: MINE PRODUCTION SCHEDULE

TABLE 16-3: MILL PRODUCTION SCHEDULE

Note: 2021 Stall tonnage includes ore processed at the Flin Flon mill.

TABLE 16-4: MILL RECOVERY SCHEDULE

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TABLE 16-5: METAL PRODUCTION SCHEDULE

Note: Production includes metal contained in concentrate and doré. Totals may not add up due to rounding.

Mine Ventilation

Lalor mine is a "push-pull" system primarily driven by pull. The exhaust shaft is equipped with two fans drawing 1.1 million cfm from the mine (Figure 16-5). Fresh air is supplied into the mine from the Chisel and Lalor downcast raises (0.6 million cfm). The Chisel and Lalor downcast raises are equipped with 600HP and 400HP fans respectively. An additional 0.6 million cfm is downcast via the Lalor production shaft, equipped with two 250HP fans. In the upper mine, ventilation raises and an internal ramp between 795mL to 755mL will provide ventilation to expanding stoping fronts. In the lower mine, an extension of the 810 to 1025mL raise is planned to connect 1025mL to 1090mL. An internal exhaust raise is planned from the 1000mL to the bottom of the exhaust shaft at 840mL. This will improve blast clearing times for the lower part of the mine. Individual mining faces are ventilated using 60 HP to 250 HP fans and 1.37 m ventilation duct. Three propane heaters heat air in the mine during the winter from Chisel, the access ramp and the production shaft.

Ventilation planning utilizes Ventsim modelling to create intermediate stages from the LOM Deswik schedule. The Ventsim model is continually updated as the mine continues to expand both upward and downward in multiple lenses.

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FIGURE 16-5: CURRENT LALOR MINE VENTILATION LONG SECTION

1901's ventilation will be a "push-pull" system, with fresh air introduced underground via a new fan and heater plant installed on surface at the existing 10 ft diameter fresh air borehole, and exhaust air returned to surface through a new 13 ft diameter RAR with main fans underground. The fresh air will be heated as required with a direct fired propane heater.

The WIM and 3 Zone deposits will use an exhaust-air decline approach. This system will be a "push" type system. The main driver of the system will be two-fan intake installation. This installation will be located at the top of the FAR which will also serve as the secondary egress. The seasonally required, mine-air heater system will also be located at the top of the FAR.

16.12. WORKFORCE

The majority of operations, maintenance and technical personnel at Lalor work 11.5 hour shifts on a 5-5-4 day cycle or a 7-7 day cycle or 8 hour day shifts, 40 hours per week. The mine is operated under Collective Bargaining Agreements between Hudbay management and local unions. The mine operations workforce is comprised of Hudbay hourly operations and maintenance personnel as well as salaried supervision, mine administration and technical staff, plus contractor personnel for specialized work and workforce shortages. Personnel count will vary year to year. Steady state personnel requirements are shown in Table 16-6 .

At 1901, development, production, construction, and maintenance work will be completed by Hudbay personnel. A mining contractor will complete specialized work that includes: raises (raisebore and Alimak), diamond drilling, major construction, grubbing and site preparations for the RAR, and main ventilation plant installation.

At WIM and 3 Zone, all development is planned to be completed by contractors, with Hudbay staff carrying out production activities.

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TABLE 16-6: MINE OPERATIONS WORKFORCE

16.13. MINE SAFETY AND HEALTH

All personnel are required to work under the applicable laws of the Province of Manitoba, Canada. All contractors working on site are required to have an approved health and safety program in place and have on site representation. Hudbay Plant Safety Rules and Regulations are used at Lalor mine including safety and health monitoring programs, dust, water and environmental monitoring, personal protective equipment and task analysis and job procedures completed by employees daily.

Refuge Stations

Refuge stations are required at Lalor mine as per mine regulations and Hudbay standards and are incorporated into the mine design. Hudbay's standard refuge station is excavated from rock and requires two ventilation bulkheads, and compressed air. Depending on the size of the refuge and the people working in the area a backup oxygen generator may be installed. Potable water, stretcher kit, first aid supplies and supplies to seal off the bulkheads are installed. In new development areas portable refuge stations replace excavated refuge station.

Three refuge stations will be constructed at 1901. There will be a refuge station in the access ramps and two at the 1901 workings on 580mL and 625mL. The refuge stations will be the same as at Lalor and will also serve as lunchrooms and can be used as a line-up room.

At WIM and 3 Zone, portable refuge stations and sanitary facilities will be strategically placed throughout the mine in active working areas.

Personal Evacuation Infrastructures

The primary route in and out of the Lalor mine is the production shaft equipped with a service cage. The shaft is equipped with a small auxiliary hoist and six person cage. In case of power failure, the auxiliary hoist can be operated by an emergency diesel generator to evacuate personnel from the mine. If the production shaft is not usable, the second egress from the mine is the main ramp to surface at Chisel North mine.

Secondary means of egress are incorporated into the designs of the Snow Lake satellite deposits. At 1901, the twin ramps and internal raises (with escapeways installed) provide alternate egresses back to the Lalor ramp, which is considered the primary entry and exit route. At WIM and 3 Zone, the main ventilation raises (FARs) will be supported and equipped with a manway so that they can be used as second egresses.

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17. RECOVERY METHODS

17.1. INTRODUCTION

The Snow Lake operations consists of two facilities to process ore from the Lalor Mine: The Stall and New Britannia mills. The Stall mill employs conventional crushing, grinding and flotation operations. The nominal throughput rate is 3,800 tonnes per day and the mill operates 24 hours per day, 365 days per year, with scheduled downtime for maintenance as required. The mill produces a copper concentrate with gold and silver credits and a zinc concentrate. The New Britannia mill is currently being refurbished with an expected completion by the end of 2021. Conventional crushing, grinding, flotation and cyanide leach and recovery methods will be used to process ore from the Lalor mine. The nominal throughput rate will be 1,500 tonnes per day, operating 24 hours per day, 365 days per year, with scheduled downtime for maintenance.

17.2. STALL MILL

In 1979, the Snow Lake mill was constructed to process ore from six operating mines within a seven-mile radius of Snow Lake, which were previously being treated at the Flin Flon mill. The new mill had two independent flotation circuits, known as the Stall circuit and the Chisel circuit. Each circuit treated the output of three mines.

The Chisel circuit processed a blend of ore from the Chisel Lake, Lost Lake and Ghost Lake mines. The circuit's capacity was approximately 1,000 tonnes per day, on a five-day basis, producing separate copper, lead and zinc concentrates.

The Stall circuit treated a blend of ores from Anderson Lake, Osborne Lake and Stall Lake mines, at a mill throughput of roughly 2,800 tonnes per day, producing approximately 410 tonnes of copper concentrate and 162 tonnes of zinc concentrate per day. Ores were truck hauled to the new mill, and concentrates were rail-hauled to the Flin Flon metallurgical complex.

The mill operated continuously until shutdown in 1993, because of depletion of the Chisel Lake Open Pit and Stall Lake Mines. The mill was reopened in 1994 to process ore from the Photo Lake Mine and continued to process ore from the Chisel North Mine until 2009, when both the mine and mill were placed on care and maintenance due to depressed metal prices. The mine and mill were re-started in 2010, following recovery of metal prices.

The mill, now referred to as the Stall mill, processes ore from the Lalor Mine, which began commercial operations in 2012. Initial capacity was 2,800 tonnes per day and through modifications and improvements, milling performance has increased to an average rate of 3,800 tonnes per day, although it has been demonstrated to reach peak throughput rates of up 4,500 tonnes per day. This mill is also envisaged to process the ore produced from the 1901 deposit.

Flowsheet

The Stall mill consists of the following unit operations:

Crushing:
- Hewitt Robins 762mm x 1220mm Jaw Crusher
- Tyroc 2.4m x 6.1m double deck screen
- Symons Standard cone crusher
- Ancillary conveying, dust collection, and other systems.
Grinding
- Three fine ore bins and associated belt feeders.
- Two rod mills in primary grinding duty:

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3.2m x 4.9m 600kW Stall Rod Mill
2.1m x 3.0m 150kW Chisel Rod Mill
Two ball mills, in secondary grinding duty:
- 3.8m x 5.5m 1200kW Stall Ball Mill
- 3.2m x 4.0m 600kW Chisel Ball Mill
Ancillary pumps, a trash screen, hydrocyclones, and other systems.
Talc flotation
- Rougher flotation using Wemco 120 self-aspirated cells.
- Two stages of cleaner flotation using Denver DR24 cells as the first cleaner and a Jameson Cell as the second cleaner.
Copper flotation:
- Rougher flotation using a Jameson Cell.
- Scavenger flotation using Wemco 120 self-aspirated cells.
- Regrinding of rougher and scavenger concentrate with a closed circuit 2.4m x 3.6m 400hp ball mill.
- Gravity recovery of gold on the regrind hydrocyclone underflow using a Knelson mill.
- Cleaner and cleaner scavenger flotation using Denver DR24 cells.
- Recleaner flotation using a Jameson Cell.
Zinc flotation:
- Rougher flotation using a Jameson Cell.
- Scavenger flotation using Wemco 120 self-aspirated cells.
- Cleaner flotation using a Jameson Cell.
- Cleaner scavenger flotation using Denver DR24 cells.
Concentrate filtration and concentrate load-out:
- Multistage centrifugal pump set for transfer of copper concentrate from the Stall mill to the New Britannia mill via a 125mm diameter pipeline.
- Filtration of the zinc concentrate using two vacuum disc filters
- Ancillary conveying, storage, and concentrate load-out systems.

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FIGURE 17-1: STALL MILL FLOWSHEET

Crushing

Coarse ore from the mine is transported to a primary crushing plant located at the Chisel mine site and is operated by a third-party contractor. The product from the crushing plant is then trucked to the Stall mill and placed in coarse ore bins or a stockpile. Material from the bins is reclaimed via vibrating feeders and conveyed into a Hewitt Robins 762mm x 1220mm jaw crusher. The discharge is combined with that of the secondary Symons Standard cone crusher to feed a Tyroc 2.4m x 6.1m double deck vibrating screen. The undersize from the bottom deck is conveyed to the fine ore bins (FOB) while the oversize of the two decks is combined to feed the Symons cone crusher and then recirculated to the screen deck in a closed loop.

The described crushing circuit is in the process of being decommissioned and replaced with an external crushing circuit to service both the Stall and New Britannia mills. The new external crushing circuit start-up is aligned with the New Britannia mill commissioning schedule.

Grinding

From the fine ore bins, crushed ore is conveyed to the two independent Stall and Chisel circuits. The Stall circuit, which processes approximately two thirds of the mill feed, includes a 3.2m diameter x 4.9m rod mill and a 3.8m diameter x 5.5m ball mill. The Chisel circuit has a 2.1m diameter x 3.0m rod mill and a 3.2m diameter x 4.0m ball mill.

Each rod mill discharges into a dedicated hydrocyclone feed pump box where it is combined with the ball mill discharge and with dilution water. Each hydrocyclone underflow returns to their respective ball mill. The hydrocyclone overflow from each circuit combines and discharges into the hydrocyclone overflow pump box. The product reports to a flotation feed pump box through a trash screen. Lime is added to each rod mill to control the pH in the flotation feed.

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Flotation

The flotation circuit outlined below is based on the Stall Mill Recovery Project Feasibility Study completed by Ausenco in Q3 2020. The recovery project envisions upgrades to the existing flotation circuit through the addition of flotation cells and instrumentation to improve the metal recovery and product quality. Detailed engineering is progressing through 2021 with completed and startup targeted for Q1 2023.

The upgraded flotation circuit consists of three sequential flotation circuits to enable the production of a copper-gold-silver and a zinc concentrate. These sequential flotation circuits are in order of process:

Talc Flotation
Copper Flotation
Zinc Flotation

Talc Flotation

The talc flotation circuit concentrates and removes hydrophobic gangue. The trash screen undersize is pumped to a bank of five Wemco 8.5m3 cells for rougher flotation. The rougher concentrate undergoes two stages of cleaning via a bank of nine Denver DR24 cells, as the first cleaning stage, and a single Z1200/1 Jameson Cell, as the final cleaning stage. The talc concentrate is discarded to the final tailings. The final cleaning stage tails reports to the rougher concentrate pump box while the rougher and first cleaner tailings become the feed to the copper circuit.

MIBC is the only reagent added to the talc preflotation flotation circuit; although, there is infrastructure to add specialty sulfide depressants to the first and final stage of cleaning as required to minimize valuable mineral loss.

Copper Flotation

The talc flotation circuit tailings are pumped to the copper rougher. The copper rougher consists of a single E3432/8 Jameson cell. The rougher tailings will undergo scavenging in ten Wemco 8.5m3 cells, with the scavenger concentrate combined with the talc flotation circuit tailings and pumped to the rougher Jameson cell. The scavenger tailings is pumped to the zinc flotation circuit. The rougher concentrate is pumped to a 2.44m x 3.66m regrind ball mill with a 300kW motor. The regrind mill is operated in closed circuit with a target P80 of 30 to 40 µm. The option is available to direct the scavenger concentrate to the regrind mill when required.

The reground rougher concentrate is pumped to a bank of twelve Denver DR24 cells for the first stage of cleaning. The first cleaner concentrate undergoes a second stage of cleaner in a single E2514/3 Jameson cell. The final copper concentrate discharges via gravity to the copper concentrate holding tank prior to being pumped to the New Britannia mill. The second cleaner tailings are pumped to the head of the first cleaner. The tailings from the first cleaner are pumped to a bank of ten Denver DR24 cells for cleaner scavenging. The cleaner scavenger concentrate is combined with the second cleaner tailings for retreatment in the first cleaner. The cleaner scavenger report to the rougher Jameson cell, together with the talc circuit tailings and the scavenger concentrate.

The copper flotation circuit utilizes lime to maintain a pH across the rougher and cleaner stages of 10.0 to 10.8. The circuit employs Aerophine® 3418A as the sulfide mineral collectors, and MIBC as the frother.

Zinc Flotation

The copper circuit tailings are pumped to the zinc rougher, which consists of a single E3432/8 Jameson cell. The rougher tailings are scavenged in seven Wemco 8.5m3 cells with the rougher scavenger concentrate being pumped to the cleaner scavenger. The rougher scavenger tailings are the flotation final tailings and are pumped to the final tails pump box. The rougher concentrate can report to two designations; when processing high grade zinc feed, the rougher concentrate is of sufficient grade to report directly to the zinc concentrate thickener; otherwise, the rougher concentrate is pumped to the zinc cleaner.

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The zinc cleaner consists of a single E3432/8 Jameson Cell. The concentrate of the cleaner is at final concentrate grade and is pumping to the zinc concentrate thickener. The cleaner tail is combined with the scavenger concentrate and is pumped to a bank of twelve Denver DR24 cells, which act as a cleaner scavenger. The cleaner scavenger concentrate is pumped back to the cleaner Jameson cell. The cleaner scavenger tailings combine with the rougher tailings and are pumped to the scavengers.

The copper flotation circuit utilizes lime to maintain a pH across the rougher and cleaner stages of 11.0 to 11.5. The circuit employs Aero® 7279 as the sulfide mineral collectors and MIBC as the frother.

Concentrate Dewatering

Flocculated copper concentrate is pumped to a 3.4m diameter thickener. Underflow, at a target density of 65% solids, is pumped to an agitated stock tank. Currently, thickened copper concentrate is further dewatered to approximately 19% moisture on a disc filter, but after start-up of the New Britannia mill, the copper concentrate will be pumped via pipeline to New Britannia for dewatering and the existing copper dewatering section in the Stall mill will be decommissioned.

The flocculated zinc concentrate follows the same process as for copper but is further dewatered to approximately 11 % moisture using two disc filters.

Tailings

Plant tailings, which are composed of zinc rougher scavenger tailings, talc concentrate and certain sumps and flows from the dewatering circuits, are pumped out of the plant to the Anderson Tailings Impoundment Area (TIA) for final disposal. Tails required for paste are diverted to the Anderson booster pump station. The capacity of the pumping station ranges from 110 to 130tph of solids. Two pipelines have been installed between the Anderson booster pump station and the paste plant, located at Lalor mine site. Paste is delivered from a discharge hopper to underground via one of two nominal 8-inch diameter, hardened steel cased boreholes from surface to the 780 m level of the mine, the main distribution level to direct the paste to other levels above and below. Only one borehole is required during normal operation. The second borehole is available as a spare in the event of a plug or excessive wear on the primary hole. A network of underground lateral piping and level to level boreholes transfers the paste from the base of boreholes to the required underground locations.

Water

The Stall mill utilizes two sources of water, freshwater and reclaim water (recycled from the tailings facility).

Freshwater accounts for approximately 25% of the water usage within the mill and is withdrawn from Snow Lake via a 7km freshwater pipeline that feeds a holding tank at the Stall mill. It is then distributed to various areas of the process where clean, high-quality water is required. This includes, worker washroom and shower facilities, gland water, reagent mixing and instrument cleaning.

Reclaim water accounts for the remaining water usage within the mill and is withdrawn from the Anderson TIA. This water is deposited into the Anderson TIA as final process tailings. The water is then reclaimed for re-use within the mill via a 5km pipeline and is the primary source of water for the grinding and flotation processes.

Stall Mill Recovery Improvement Project

Based on the historical performance, the circuit flowsheet configuration and the metallurgical testwork the forecast metal recovery range for the Stall Concentrator treating Lalor Base Metal Ore is outlined Table 13-8. If changes are not made to the current flowsheet, e.g., installation of additional flotation capacity, the recovery of recovery of copper, gold and zinc will be within the historical performance range. Hudbay plans to implement the upgrades to the Stall flowsheet based on the successes of the test program described in section 13 of this report. The upgraded flowsheet design calls for the installation of:

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Increase copper rougher and cleaner capacity with the addition of Jameson cells.
Addition of talc preflotation circuit with the re-configuration of existing flotation cells and addition of a Jameson cell cleaner stage.
Increase zinc circuit cleaning capacity and add froth washing through the addition of Jameson cells.
Increase recovery of free gold through the addition of a Knelson gravity concentrator on the copper regrind cyclone underflow.

The comparison of the current and forecasted recovery and product grades increase as per the upgraded flowsheet are summarised in Table 17-1.

TABLE 17-1: FORECAST METAL RECOVERY FOR THE CURRENT FLOWSHEET AND AN UPGRADED FLOWSHEET

The Stall Recovery Improvement Project feasibility study was completed in 2020 by Ausenco with a Class 3 capital estimation of CAD$24.5M excluding tax. The breakdown of direct, indirect and owners' costs inclusive of contingency is 57%, 33% and 10% respectively. The cost estimation benefited from benchmarking of labor and bulk rates with the New Britannia project.

TABLE 17-2: STALL RECOVERY PROJECT CLASS 3 CAPEX

The project is scheduled to progress through detailed engineering in 2021 and construction in 2022 following the completion of the New Britannia Project. Completion and operation of the Stall Recovery Improvement Project is schedule for Q1 2023.

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Future Stall Projects

As the Lalor mine matures and additional satellite ore sources are identified to supplement the feed blend to the Stall mill, metallurgical projects are actively advanced to match the configuration of the mill to the requirement of the ore. The following are projects that are being advanced through metallurgical testwork, early stage engineering or both.

The current life of mine plan for Lalor shows an increasing Pb:Cu ratio from 0.3 to >1.0. In addition, the 1901 deposit has a Pb:Cu ratio of >1.5 indicating the need for a lead mitigation strategy. Preliminary results indicate that a traditional sequential Cu-Pb-Zn flotation circuit will mitigate the increasing lead in feed and provide additional lead concentrate revenue stream by capturing a significant portion of the copper, gold and silver that is currently lost to the zinc concentrate.
Stall Tails Leach - The flotation recovery of gold is limited by the mineralogy and association of gold with iron sulfides. Due to the mineralogical differences between the gold zone and base metal zone ores, the flotation tailings from the base metal ores are not compatible with the same leaching conditions as the gold zone ore flotation tailings.
Stall Fine Grinding - Along side the Tailings Leach project, a fine grinding project is being advanced to determine the improvements in flotation recoveries when the flotation feed size is reduced from the current p80 of 120µm to 60µm.

17.3. NEW BRITANNIA MILL

In March 1949, the Nor-Acme mine and gold mill was opened. Production continued until 1958, with the Nor-Acme mill recovering approximately 610,000 oz of gold during this period. Between 1958 and the early 1990s, various companies worked on the property with the intention to bring the mine back into production. These efforts were unsuccessful.

Flowsheet

The New Britannia mill consists of the following unit operations:

Ore Receiving
150 tonne capacity dump pocket including a grizzly screen cover.
Associated belt feeder and conveyor systems
Single 2700 tonne capacity fine ore bin
Grinding
3.5m x 4.5m 750kW Rod Mill
4.2m x 6.1m 1500kW Ball Mill
Ancillary pumps, a trash screen, hydrocyclones, and other systems
Talc flotation
Rougher flotation using Direct Flotation Reactors
Copper flotation:

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Rougher and cleaner flotation using Jameson Cells.
Leaching, CIP, Elution and Refining
Four leach tanks
Six carbon-in-pulp tanks
Loaded carbon, carbon dewatering and regenerated carbon screens.
Fine carbon filter
Carbon regeneration kiln
Acid wash vessel
Pressure strip vessel
Two electrowinning cells
EW sludge filter
Induction furnace
Concentrate filtration and concentrate load-out:
8m diameter high-rate thickener
Horizontal plant pressure filter
Tailings
Two cyanide destruction tanks
Multistage centrifugal pump sets for transfer of final tailings from the New Britannia mill to the Stall mill via a 200mm diameter pipeline.

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FIGURE 17-2: NEW BRITANNIA MILL FLOWSHEET

Ore Receiving

Coarse gold and copper-gold ore from the mine is transported to the external crushing plant at the Stall Mill. The product from the crushing plant is delivered to the New Britannia site using 40 tonne trucks, where the material will be loaded into a dump pocket and conveyed to the fine ore bin (FOB).

Grinding

The crushed material is reclaimed from the FOB via conveyors into the rod mill and then discharged into a hydrocyclone feed pump box, where it is combined with the ball mill discharge and dilution water. A variable speed hydrocyclone feed pump feeds a the hydrocyclones cluster. The hydrocyclone underflow returns to the ball mill while the overflow flows via gravity to a trash screen and then is pumped to the flotation circuit.

Flotation

From the hydrocyclone overflow pump box, the slurry is pumped to the talc flotation circuit. Talc flotation consists of two 1.7m diameter Direct Flotation Reactors (DFR), operating as either a hydrophobic gangue removal stage ahead of the Cu flotation circuit to improve final concentrate quality, or as an additional Cu scalping stage in the event of above-design Cu feed grades.

The talc flotation tailings are sent to the flotation feed pump box and from here it is pumped to the first of three E1732/4 Jameson cells. The concentrate from the first Jameson cell, the Cu scalper, is at final product grade and reports to the Cu concentrate thickener. The tailings from the Cu scalper are pumped to the second Jameson cell, the Cu rougher, with the concentrate from this cell reporting to a single stage of cleaning in a Z1200/1 Jameson cell. The tailings from the Cu rougher is pumped to the third Jameson cell, the Cu Scavenger, with the concentrate reporting back to the Cu rougher along with the Cu cleaner tailings. The concentrate from the Cu cleaner reports along with the Cu scalper concentrate to the Cu concentrate thickener. Note that the combination of two 1.7m DFRs in talc flotation duty, combined with three same-sized Jameson cell units in the copper flotation circuit allows for maximum circuit flexibility to changes in head grades or feed mineralogy.

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Concentrate Dewatering

The combined concentrate from the Cu scalper and Cu Cleaner is mix with the copper concentrate from Stall Mill in the 8m diameter copper thickener. The concentrate is thickened, pressure-filtered, and stored in a concentrate shed, before being loaded onto trucks and transported offsite.

The Cu scavenger tails, depleted of copper, reports to the 12 m diameter pre-leach thickener for thickening prior to cyanidation. The underflow product of the thickener is approximately 50% solids.

Thickener overflow from the copper concentrate and pre-leach thickeners is returned to the process water tank.

Leaching

The thickened cleaner scavenger tailings and rougher tailings is fed to four leach tanks in series with aeration. The total retention time is at least 50 hours. Hydrated lime is added to ensure a pH above 11.5 and sodium cyanide solution is added to leach the gold. The slurry stream discharge from the final leach tail gravitates to six CIP leach tanks in series with a 1.2 hr retention time in each. Carbon is added to the sixth tank and is transferred counter-currently, adsorbing gold and silver as a cyanide-gold complex as it progresses upstream to the first CIP tank.

At the first tank, loaded carbon is transferred to a fine vibrating screen where it is separates the carbon from the slurry. The loaded carbon gravitates to a holding tank and is batch transferred periodically and washed with a mild acid solution to remove impurities. The washed carbon is then stripped using a hot caustic cyanide solution (pregnant solution). The pregnant solution proceeds to electrowinning. The sludge collected is dried and treated with fluxes and melted in the induction furnace to obtain the Doré. The waste slag from is collected, crushed, and added to the rod mill.

The stripped carbon, after elution, will retain a significant quantity of organic and inorganic contaminants (foulants) present in the cyanide slurry. Prior to returning to the process, the carbon will be regenerated by removing these contaminants in a kiln.

The final tailings from the sixth CIP tank gravitates to the carbon safety screen where fine carbon is collected and sent to the fine carbon tank. The screen undersize is followed by the cyanide destruction stage. The cyanide destruction is achieved via the INCO SO2/Air process utilising sodium metabisulphite as the SO2 source. The carbon safety screen undersize is pumped to a splitter box which evenly distributes the tailings slurry between two air-sparged cyanide destruction tanks operating in parallel each with 1hr retention time. The tailings slurry post cyanide destruction gravitates to the final tailings pump box.

Tailings

The final tailings from the New Britannia mill is pumped to the Stall mill via a 6.8km pipeline. This same pipeline corridor will house the copper concentrate pipe and the reclaim water (both with flow from Stall to New Britannia). Once the tailings line reaches the Stall mill tailings pump box, the paste plant has the option of directing the flow to the paste plant or to the Anderson TIA, depending on the needs for paste backfill.

Water

The New Britannia mill utilizes three sources of water, fresh water, process water and reclaim water (recycled from the tailings facility).

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The fresh water is utilized in areas of the process where clean high-quality water is required, including the worker washroom and shower facilities, gland water, reagent mixing and instrument cleaning. The freshwater is sourced from Snow Lake to the New Britannia mill. It is fed into a holding tank within the mill building and distributed to the process as required.

Process water is used in the grinding and flotation sections. It is sourced from the concentrate and pre-leach thickener overflow streams.

Reclaim water accounts for the remaining water usage within the mill and is withdrawn from the Anderson TIA. This water is deposited into the Anderson TIA as part of final tailings. The water is then reclaimed for use within the New Britannia mill via a pipeline along the same corridor as the tailings and concentrate pipelines. The amount of reclaim water required by the New Britannia mill is proportional to the quantity of water deposited with the final tailings.

Future New Britannia Projects

As the Lalor mine matures and additional satellite ore sources are identified to supplement the feed blend to the New Britannia mill, metallurgical projects are actively advanced to match the configuration of the mill to the requirement of the ore. The following are projects that are being advanced through metallurgical testwork, early stage engineering or both.

Throughput Increase - The previous operation of the New Britannia mill saw throughput rates of up to 2,000tpd based on the availability of ore from the underground operation. With the increase in the mineral reserves of the Lalor gold zones alongside additional the satellite deposits opportunity exists to increase the throughput rate. The study to date has identified the tailings pipeline, pre-leach thickener and various minor pumping systems require upgrade to accommodate the increased tonnage rate. All the other major systems within the New Britannia mill have sufficient capacity to accommodate the increased tonnage. In addition to the processing equipment upgrades the study includes the upgrade of the pipeline corridor road to accommodate ore haulage from Stall to New Britannia to remove traffic from the public access road.
1901 Gold Zones - The 1901 deposit contains two gold-copper lenses which are suitable feed for the New Britannia mill and would supplement the Lalor gold zone ore. This project is advancing through metallurgical testwork.

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18. PROJECT INFRASTRUCTURE

18.1. AREA SURFACE INFRASTRUCTURE

Lalor Mine is located 16 km by road from the Town of Snow Lake, Manitoba. General area infrastructure includes provincial roads, 115 kV Manitoba Hydro grid power within four km of Lalor and Manitoba Telecom (Bell) land line and cellular phone service.

The Town of Snow Lake is a full-service community with available housing, hospital, police, fire department, potable water system, restaurants and stores. The community is serviced by a 914m gravel airstrip to provide emergency medical evacuation. Hudbay owns a contractor operated 260 person camp in Snow Lake and the New Britannia mine, mill and office.

Figure 18 1 provides an overview of the various infrastructures and of the material flow from the mine to the tailings disposal facility in Snow Lake.

FIGURE 18-1: MAP OF SNOW LAKE MINING, PROCESSING AND TAILINGS FACILITIES

Source: Google Earth, 2021 image

Chisel North Mine Site

The Chisel North mine site is located 3.5 km from Lalor Mine. Infrastructure includes:

Building complex with offices, changehouse and shop,

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Lalor Mine electrical substation with two (2) 115-25 kV 24 MVA transformers,
Water pumping station (booster pump station) with holding tanks and pumps, and
Contractor owned and operated ore crusher. Stockpile capacity at Chisel North is 15,000 tonnes.

South of the Chisel North mine site, along PR 395, there are:

Two mine fresh air raises equipped with downcast fans, mine heaters and each with a 30,000 US gallon propane tank.

Chisel Lake Mine Site

The Chisel Lake mine site is a decommissioned mine located approximately 7.5 km from Lalor. Infrastructure includes:

Mined out open pit used for potentially acid generating (PAG) waste rock disposal and storing mine discharge water and water treatment plant sludge,
High density sludge process acidic water treatment plant that can treat up to 2,500 gpm, and
Fresh (process) water pumps for Lalor Mine.

Anderson Tailings Impoundment Area Site

The permitted Hudbay Anderson TIA is located approximately 12 km from Lalor. Infrastructure includes:

Slurried tails pump station,
13 km of 14" buried pipeline from Anderson TIA to Lalor to deliver slurried tailings to the Lalor paste backfill plant, and
13 km of 10" buried pipeline to return water from the paste plant to Anderson TIA.

Mine Access Corridor

The mine access corridor is the roadway between the PR395 at the Chisel North mine site to Lalor mine site. Infrastructure in the corridor includes:

Four km of 25 kV overhead power lines,
Buried and heat traced process water and mine discharge pipelines, and
Buried slurried tailings and return water pipelines.

WIM and 3 Zone Project Sites

The WIM deposit is located approximately 15 km by road north of the New Britannia mill, 315 m above sea level. Access to WIM is currently via a winter road.

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3 Zone is accessible via road and located approximately 3 km by road northwest of New Britannia Mill. From Lalor, it is located approximately 8 km northeast (12 km by road).

Required infrastructure at both sites includes:

A gravel access road to site from the New Britannia site. This road will require year-round maintenance.
A trailer with washroom facilities and a small conference area.
Powerlines along access road to site.

18.2. MINE SITE SURFACE INFRASTRUCTURE

A site drawing of the Lalor access road and services is shown in Figure 18-2. The Lalor mine surface infrastructure includes the following:

Mine site services: This includes pole and buried electrical power lines, transformers and switchgear, and buried freshwater, mine discharge water and tailings pipelines.
Hoist house: Contains electrical distribution for the site, hoist and communication control room, production hoist (Davy Markham, double drum, 4,828 kW), service hoist (Davy Markham, double drum, 2,414 kW), three GA 250 screw compressors (1,477 cfm @120 psi each).
Headframe: Contains utility hoist (Davy Markham, single drum, 314 kW), bin house (approximate capacity 1,000 tonnes of ore), external bunker (capacity 1,200 tonnes of ore), two 250 hp downcast fans (313,000 cfm each) and mine air heater.
Two exhaust fans (2,500 hp and 575,000 cfm each) located approximately 1.0 km from the production shaft.
Main pump station: Includes holding tanks for discharge water, process water and potable water, PAL water system and pumps for discharge, potable, process and fire water.
Paste backfill plant with tailings receiving and storage tanks, tailings thickener, tailings filters and cement storage silos.
341 person change house complex that also houses staff offices and 100 person and 50 person change houses adjacent to the main complex.
Two 30,000 US gallon propane tanks.
Office complex for health & safety, training and mine rescue.
Fuel tanks and pumps for diesel and gas.
Warehouse and shop.
Multiple storage containers (sea containers) and contractors' office trailers.

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FIGURE 18-2: LALOR MINE SITE GENERAL ARRANGEMENT

Site drawings of the WIM and 3 Zone access roads and services are shown in Figure 18-3 and Figure 18-4. The mines surface infrastructure includes the following:

Maintenance facilities: the primary maintenance facility for surface and underground equipment constructed to service Lalor Mine is too distant to utilize and a small two bay facility has been costed at the site for the duration of operations.
Warehouse: all purchasing is handled by the on-site staff, with regular freight movement between sites. The on-site warehouse space accommodates spares for mining as well as milling operations.
Fuel and lube distribution: the fuel system includes a surface fuel farm and storage tanks with fuel distribution to the underground operations via a fuel and lube truck. This service will be provided by a contractor.
Main administration offices: Management, supervisory and technical teams will be centralized in an office complex, located at either New Britannia mill or Hudbay's Lalor Mine site.
Mine dry & conference space: A trailer will be placed at the WIM and 3 Zone locations to provide facilities for staff. The trailers will have areas for meetings and a lunchroom. The new dry facilities at New Britannia site be utilised and workers bused to the site.
Electrical and power distribution: the above ground electrical distribution originates at the New Britannia Mill site and terminates at the WIM and 3 Zone portals. From there, the underground electrical distribution system will be supplied by two primary feeders from the mine services power distribution switchgear on the surface.

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FIGURE 18-3: WIM SITE LAYOUT

FIGURE 18-4: 3 ZONE SITE LAYOUT

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18.3. MINE UNDERGROUND INFRASTRUCTURE

Mining infrastructure includes excavations and equipment normally used below the shaft collar. It includes for Lalor:

Main Production shaft; 6.9m diameter concrete lined with five compartments equipped with:
Two - 16 tonne skips
One double deck cage (50 people per deck)
Counter weight
Utility conveyance (6 person cage)
Three main shaft stations at 835m, 910m and 955m levels
Lateral development, consisting of 6m x 5m ramps and level development connecting all mine workings.
Ramp from surface to 810m level. Secondary egress from the mine. Total distance is approximately 6.0 km.
Power Distribution consisting of:
2 x 25 kV power cables down the shaft
7.5 MVA 25 kV to 13.8 kV transformers at the 835m level and 910m level shaft stations
Primary distribution throughout the mine is 13.8 kV with transformers to 600 V for local distribution
Compressed air, process water and discharge water pipelines throughout the mine.
Leaky Feeder underground wireless radio communication system.
Fiber-optic backbone for data and video.
Ore handling system consisting of two rock breakers and bins on 910m level feeding chutes and conveyor system on 955m level.
Mobile Maintenance shop located at the Chisel North underground workings. Lalor maintenance shop located near the 910m level shaft station is under construction.
Water Discharge system consisting of a series of drain holes and sumps with submersible pumps. Two settling cones on 910 m level. Clean water sumps on 955m level with 1,250 hp 10 stage Mather Platt pump (740 gpm).
Internal ore passes to transfer ore from upper levels of the mine to 910m level.

Electrical Distribution and Equipment

Grid electricity is supplied by Manitoba Hydro, the provincial power utility. Manitoba Hydro's 115 kV power line terminates at the Chisel North Mine site, approximately 3.5 km by road from the Lalor mine site. This feeds power to the Hudbay owned main distribution substation consisting of two 115-25 kV 24 MVA transformers. The substation is completely equipped with an E-house complete with four GE Powervac circuit breakers and Tie breaker. From there, power is routed as follows:

1) Breaker 52-C1 provides power to:

(a) The Chisel North mine site

(b) The Chisel Lake mine site

(d) Lalor mine ramp downcast fan site

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2) Breaker 52-L1 - Lalor CCT1 provides power to the following areas:

(a) Lalor mine site pump house and water treatment plant

(b) Lalor mine exhaust fans

(d) Hoist house electrical room Bus 1 which distributes power to:

Mine underground Feeder A
VFD drives for production hoist
MCC for production hoist
Hoist house PDC (compressors)
MCC for Service hoist
Auxiliary hoist

3) Breaker 52-L2 - Lalor CCT2 provides power to the following areas:

Mine UG Feeder B which feeds
Headframe Complex
VFD Drives for Service hoist.
Intake fans
Lalor mine Office/Dry

Mine Underground Power Supply

Lalor mine underground mine electrical distribution to the mine workings consist 13.8 kV that is further stepped down to 600V. This power is used to power:

(a) Auxiliary ventilation fans,

(b) Auxiliary pumps,

(d) Underground fiber communication network.

Lalor mine underground mine power distribution equipment currently consists of:

(a) 27 X 1 MVA 13.8-0.6 kV mine power centres (aka. portable sleds),

(b) 16 X 0.75 MVA 13.8-0.6 kV mine power centres, and

The mine power centers are located in constructed underground substations.

Emergency Generators

Emergency generators are required to maintain water flows and to evacuate the mine in the case of power failure. Emergency generators are located at:

Lalor mine pump house and treatment plant - Cummins 350 kW
Booster pump station (Cummins 350 kW)
Lalor auxiliary hoist (Cummins 1000 KVA)
Paste backfill plant (Cummins 1000KVA)

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Anderson pump station (Cummins 1000 KVA)

All generators are equipped with automatic transfer switch that will in the event of a power failure transfer to generator power and vice versa.

1901 Underground Infrastructure

Mining infrastructure at 1901 is based on current practices at Lalor to maintain commonality with equipment, materials, operating procedures, and workforce skillsets. Mine services (compressed air, power, process water, drainage, etc.) are to be integrated with Lalor systems. Key elements include:

Ventilation: push-pull ventilation system - 305 kcfm airflow requirement
Fresh air down existing (unused) surface borehole, through new haul ramp and exhaust to surface via new service ramp and RAR
Fresh air heated as required with a direct fired propane heater
3.5mØ RAR to be raisebored (495m length)
2 x 200 hp main fresh air fans and 2 x 250 hp return air fans
Electrical: feed from Lalor 825mL transformer
9 x electrical sub-stations
Service water (drilling, dust suppression) and drainage:
Mining activity requires 350 L/min during shift average, with potential peak of 1,140 L/min
Ditches, borehole collection sumps, level sumps, main pump station required
Supply from / discharge to water reservoir in Remuck #3 in Lalor Ramp
Other infrastructure such as 3 x refuge stations, SatStat fueling station, explosives & materials storage, escapeways also required

WIM and 3 Zone Underground Infrastructure

A new portal and ramp will be constructed at WIM, while the existing ramp and decline at 3 Zone will be refurbished for use during planned operations. Key elements of the underground infrastructure include:

Dewatering equipment based on a system capacity of 500 GMP (114m³/h). All components are sized to handle that flow rate regardless of the elevation, using the following components:
Pump stations: each portable pump station comprises of a tank with 5 minutes of retention time, two slurry pumps (one operating and one standby), and a power and control panel.
Booster stations: each portable booster station comprises two slurry pumps (one operating and one standby), and a power and control panel.
Clarifiers: each clarifier unit can treat up to 70 m³/h and uses lamella plates to separate solids
Piping: all piping is to be 8" HDPE DR11 with Victaulic couplings.Mine communication and control systems: through leaky feeder and phone line systems installed in the trailer on site. The leaky feeder system is used through the primary development and level accesses to provide radio communication.
Other infrastructure: portable refuge stations and sanitary facilities, SatStat fueling station, explosives & materials storage, and escapeways

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18.4. PASTE PLANT AND ANDERSON PUMP STATION

Construction and commissioning of the paste plant and distribution system was completed in the third quarter of 2018. Paste is critical for the sustainability of the mine production plan. The paste plant is designed to fill voids left by mining of approximately 4,500 tpd. Taking into account waste generated from development in the LOM and the plan to not hoist waste from underground, the combined paste/waste backfilling capacity is approximately 6,000 tpd.

The paste plant is located northeast of the existing headframe complex at the Lalor mine site. Paste delivery capacity is 165 tph solids (tails) or 93 m3/hr paste. The paste plant is capable of varying the binder content in the paste to provide flexibility in the cure rate of the paste where higher and early strength may be required depending on mining method.

Tails are pumped from the Stall concentrator to the Anderson TIA. Tails required for paste are diverted to the Anderson booster pump station. The capacity of the pumping station ranges from 110 to 130 tph to allow for variation in the output of tailings from the concentrator. Tailings are directed into the Anderson TIA when not required for the paste plant.

Two pipelines have been installed between the Anderson booster pump station and the paste plant located at Lalor mine site, approximately a 13 km distance. The tails slurry pipeline is a nominal 14 inch diameter and the return water pipeline is a nominal 10 inch diameter. The main route of this pipeline is on top of the existing abandoned rail bed (property owned by Hudbay), then along the west side of Provincial Road #395 and finally along the south side of the Lalor mine access road to the paste plant.

Paste is delivered from a discharge hopper to underground via one of two nominal 8 inch diameter, hardened steel cased boreholes from surface to the 780m level of the Lalor mine. Only one borehole is required during normal operation. The second borehole is available as a spare in the event of a plug or excessive wear on the primary hole. The boreholes were drilled and cased in 2016.

A network of underground lateral piping and level to level boreholes transfers the paste from the base of boreholes to the required underground locations. The 780m level is the main distribution level to direct the paste to other levels above and below. Underground development connects the paste fill distribution system to mine workings via existing and planned development into stopes.

Operational improvement projects have been implemented to enhance the reliability of the tailings delivery system on surface. This includes the establishment of additional water recirculation, improved environmental containment, flexibility for water input due to tailings level expansion at Anderson tails area, and winterizing improvements at the Lalor Paste Plant. While some of these improvements are already in place, remaining construction on these improvements is slated for completion in the second quarter of 2019.

The paste plant utilizes the Lalor mine compressors. Paste plant and Anderson booster station infrastructure includes:

Four slurry booster pumps
Thickener - 14m x 5.5m
Two filter feed tanks - 14m x 14.5m
Disc filter - 9.4m x 4.8m x 4.5m
Vacuum pump
Paste hopper
Continuous mixer
Bag house dust collector and dust collector blower
Two cement silos - 4.1m x 15.2m

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Two screw conveyors
Tow rotary valves and loss in Wt feeder
One flocculent system
Two MCC rooms - ABS and Paste Plant
One crane
Electric heating installations

18.5. STALL CONCENTRATOR AND ANDERSON TIA

Hudbay operates the Stall concentrator approximately 16 km from Lalor. The mill infrastructure consists mainly of two buildings. The crusher building houses the jaw and cone crushers, the screen deck and conveyors. The mill building hosts the grinding mills, flotation circuits, dewatering processes and concentrate shed areas. Three trailers are on site to provide office space. As of January 2019, the current Stall concentrator infrastructure includes:

Two 350 tonnes coarse ore bins and 3 750 tonnes fine ore bins
One Hewitt Robins 762mm x 1219.2mm jaw crusher
One Symons Standard 7 ft cone crusher
One Tyroc double deck vibrating screen - 2.4m by 6.1m screen frame, 50.8mm on top deck, 19.05mm bottom deck
Two Rod mills - AC - (2.1m x 3.0m) and (3.2mx4.9m)
Two Ball mills - AC - (3.2m x 4.0m) and (3.8m x 5.5m)
One Regrind mill - AC - (2.4m x 3.7m)
17 - 8.5 m3 Wemco cells
22 Denver cells
Two thickeners - 3.4m and a thickener/clarifier - 2.1m
Two Zn Filters - 2.7m and 1.8m x 6 discs agitdisc - and one Cu Filter - 1.8m x 7 discs
Eight cranes
One lime silo
Two MCC Rooms
Three compressors and 3 blowers
Utility pipeline - crusher and mill
Two transformers on site to 4.16 kV and 6 Power Distribution Centers (PDCs)
Two fuel tanks and pumps for diesel and gas respectively, 2 main propane tanks and auxiliary propane tanks for office trailers
One backup generator (1 MW).
Tree x water pump stations including holding tanks, MCCs and generators
Quonset and shed for part storage and muster point and 3 trailers for office space
Dry space for approximately 90 personnel in total

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The Anderson Tailings impoundment area (TIA) is located between the Stall concentrator and Lalor mine. The Anderson TIA has been in use since 1979, when a control dam was built at the east end of Anderson Lake across Anderson Creek. Seasonal discharge of water out of the Anderson TIA occurs during the open-water season, usually from May to October. Water quality at the final Anderson TIA discharge point has at all times been in compliance with applicable regulatory requirements. Tailings are deposited subaqueously into the TIA and no treatment, other than retention in the TIA, has ever been required.

Hudbay has submitted a Notice of Alteration to Manitoba Sustainable Development to expand the TIA within the existing limits to accommodate the future tailings produced through the entire Lalor mine operations. The commissioning of this expansion is anticipated to be completed in the fall of 2019.

18.6. NEW BRITANNIA MILL

The New Britannia site is a brownfield site previous operated from 1995 to 2004 and placed into care and maintenance in 2005. The infrastructure that will be required for processing operations including existing infrastructure will include (Figure 18-5):

Administration building
Dry facilities and warehouse
Maintenance Workshops
Ore receiving dump pocket
Fine ore bin
Mill building (includes mills, reagents mixing, CIP, elution, gold room, control room)
Pre-leach thickener and leach tanks
Flotation and concentrate loadout building
Pipeline corridor (tailings to Stall, reclaim water and copper concentrate to Stall)
Pipeline service road

FIGURE 18-5: NEW BRITANNIA SITE LAYOUT

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The administration building will house nonproduction personnel which includes HR, accounting, safety and environmental. The building is pre-existing from the previous operation and has been refurbished. The mill dry facilities house the plant change rooms, maintenance offices and warehouse. The building is pre-existing from the previous operation and has been refurbished. The maintenance workshops have been relocated from the mill building to redundant preexisting buildings onsite includes the hoist building.

The mill building houses the grinding circuit, CIP, elution, reagent mixing, DFRs, gold room, process/reclaim water systems, lime slaking, cyanide destruction, plant control, assay laboratory and plant production offices. This building is pre-existing from the previous operation and is being extensively refurbished with the complete replacement of several subsystems including but not limited to the carbon regeneration kiln, acid wash vessel and carbon screens.

To the east and outside of the mill building is the pre-leach thickener, fine ore bin and leach tanks. These buildings are pre-existing from the previous operation and is being refurbished. Included in the refurbishment is the expansion and lining of the containment area around the leach tanks.

The flotation building is a new structure immediately to the north of the mill building. It contains the copper flotation circuit, concentrate dewatering system and concentrate loadout. The concentrate pipeline from Stall mill discharges in this building into the concentrate thickener.

The pipeline corridor between New Britannia and Stall mill consists of a 8"' tailings pipeline carrying the New Britannia leach tailings after cyanide destruction to the Stall mill, a 5" concentrate pipeline carrying Stall copper concentrate to the dewatering circuit at New Britannia and a 6" reclaim water pipeline from Stall to New Britannia. The pipeline corridor construction included a light vehicle service road, leak containment system and a bridge to cross Snow Creek.

TABLE 18-1: NEW BRITANNIA MILL REFURBISHMENT - PROJECT DEVELOPMENT SCHEDULE

18.7. SNOW LAKE CAMP

The camp is located in the town of Snow Lake and services Hudbay employees and contractors for the mine and mill operations. The camp began operations in January 2011 and has an area of 13,124m2. As of February 2021, the camp has seven dorms and three bunkers for a total of 260 rooms. There are also three dedicated trailers for non-dormitory purposes: a 640m2 trailer provides a dining room for employees and kitchen for camp employees and the two remaining trailers are a recreation room and a gym.

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Utilities are provided by the Town of Snow Lake, including potable water, sewage and electricity. Heating is electrical and propane. The camp also provides parking spots for employees. Snow Lake Camp

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19. MARKET STUDIES AND CONTRACTS

19.1. METAL MARKETING

Lalor produces the majority of its zinc concentrate and copper concentrate with gold and silver credits from the Stall Mill. During 2021, a portion of Lalor's production will be delivered to and processed in the Flin Flon Mill.

The refurbished New Britannia mill will process run of mine from the Lalor gold zone to produce a copper concentrate and gold doré bar.

Hudbay does not use intermediate agencies to sell its products from Lalor. The production from Lalor is free of encumbrances and from any streaming agreement and there are no existing contracts that would materially impact the planned revenues in the demonstration of economic viability of the mineral reserve estimates.

Zinc Concentrates and Zinc Metal

Lalor zinc concentrate produced from its Stall Mill is delivered by truck to Hudbay's operations in Flin Flon. It is processed in the Flin Flon zinc plant into refined zinc metal and then sold and delivered to customers in Canada and the United States of America. Production of refined zinc metal will occur until the closure of the Flin Flon zinc plant which is expected during the first half of 2022.

Upon the closure of the Flin Flon zinc plant, Lalor's zinc concentrate will either be sold direct to third party zinc smelters located throughout the world or to merchants for delivery to zinc smelters located throughout the world. To date, no sales contracts have been established for the sale of Lalor zinc concentrate but it is anticipated that all sales contracts will be at standard industry terms commensurate with the quality of Lalor zinc concentrate.

Copper Concentrates

Lalor copper concentrate produced from its Stall Mill and the Flin Flon Mill is sold directly to a copper smelter in North America under a long term frame agreement. This agreement references the annual benchmark agreements between major concentrate producers and non-integrated smelters for the purpose of fixing the annual treatment and refining charge. The Lalor copper concentrate is delivered by rail.

Lalor copper concentrate expected to be produced from the refurbished New Britannia Mill have not been committed. It is anticipated that all sales contracts for the sale of copper concentrate produced from the refurbished New Britannia mill will be at standard industry terms commensurate with the quality of New Britannia copper concentrate.

Gold Doré

The refurbished New Britannia mill will commence production of gold doré in 2021. It is forecast to contain approximately 70% silver, 25% gold and 5% other elements. The doré bar will be delivered and sold to refineries using specialized third-party transportation at standard industry terms.

19.2. CONTRACTS

Engineering, supply and construction contracts are initiated, managed and administrated by Hudbay's Manitoba Business Unit. Hudbay follows a standard contracting-out process that specifies contractors' requirements to be eligible to be considered for work. Contractor selection criteria include ability to complete the work within the required time, safety record and programs, price, and proposed alternatives. The Lalor contracts that are in place have rates and charges that are within industry norms.

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Hudbay has a marketing division that is responsible for establishing and maintaining all marketing and sales administrations of concentrates and metals. As well, Hudbay conducts ongoing research of metal prices and sales terms as part of normal business and long range planning process to achieve market terms. Contract terms used in the Lalor financial evaluation are based on this research and the author has reviewed these results and they support the assumptions made in this Technical Report.

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20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

The permits required for the current Lalor operation, including the Lalor mine, Stall concentrator and Anderson tailings facility have all been issued and remain valid.

The 1901 deposit would leverage all existing surface and underground development near Lalor operations. Significant environmental baseline work has recently been conducted by AECOM and in conjunction with the significant amount past studies will be used to gain approvals for this development should it prove viable.

The re-commissioning of the New Britannia concentrator involves the placement of the Lalor gold ore tailings in the fully licensed Anderson tailings facility or the Lalor Paste fill plant via a pipeline connecting Stall concentrator to Lalor Mine. All environmental permits are in place for the refurbishment of the New Britannia mill and associated infrastructure.

The New Britannia site includes the Birch Lake Tailings Management Facility (BLTMF). The BLTMF is currently on care and maintenance since the closure of the mine in 2005. At this stage, and given Hudbay's history in the area, there is no indication that the approvals will not be obtained within the project schedule. Initial environmental reviews and assessments will need to be completed at Watts. Permitting efforts will be completed prior to the project advancement. Hudbay plans to engage qualified consulting firms to carry out baseline studies when there is a better understanding on when the Watts potential mine production would best fit into the integrated business plan of the Snow Lake operations of Hudbay.

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20.1. ENVIRONMENTAL STUDIES AND PLANNING

Lalor Mine Stall Mill and Paste Plant

Prior to commencing advanced exploration work at Lalor, environmental impact assessments including a baseline investigation were initiated in 2007 by AECOM, a consulting group that specializes in environmental service. This investigation included all the required terrestrial and aquatic field studies. The Lalor Advanced Exploration Project plan (AEP) was submitted in the first quarter of 2010 and approved in the second quarter of 2010. The Lalor mine Environment Act Licence (EAL) application was submitted during the second quarter of 2012 and was approved in the first quarter of 2014.

Baseline work and studies were also performed by AECOM for the Anderson Tailings Impoundment Area (TIA) expansion. The notice of alteration was submitted in the third quarter of 2016 and approved in the second quarter of 2018. Baseline work from other Lalor related projects was used for the Lalor paste plant Notice of Alteration (NoA) which was submitted in the fourth quarter of 2016 and approved in the first quarter of 2017. At this time, there are no known environmental issues which could materially impact the ability to extract mineral resources or mineral reserves.

New Britannia Mill

Baseline work and environmental studies were conducted by AECOM and included in the (NoA) that was submitted in March 2019. The NoA was approved by Manitoba regulators in February of 2020. This approval allows Hudbay to restart the New Britannia concentrator and associated pipeline development linking the New Britannia and the Stall concentrators. The water withdrawal permit was updated to accommodate the refurbished concentrator in June 2020.

Snow Lake Satellite deposits

Due to the extensive work completed by AECOM and other existing studies completed as part of environmental effects monitoring programs at the various operations in the Snow Lake area, only limited additional baseline studies may be required for potential future modification of Hudbay's current properties. There is no present indication that future approvals will not be obtained to meet potential future construction schedules.

It is envisaged that AECOM will be utilized to complete the necessary baseline work at Watts. Based on our previous experience in the region, it is anticipated that regular processes familiar to our previous long term permitting efforts that all necessary permits will be obtained to meet the project schedule.

20.2. WASTE AND TAILINGS DISPOSAL AND WATER MANAGEMENT

Lalor Mine, Stall Mill and Paste Plant

There are no known environmental concerns which could adversely affect Hudbay's ability to operate Lalor mine and the 1901, WIM and 3 zone deposit as currently envisaged. Since the mine site and its associated satellite deposits are near existing facilities in the Snow Lake area, Hudbay will be able to utilize infrastructure, services, and previously disturbed land associated with permitted, pre-existing and current mining operations in the Snow Lake area.

Water management initiatives for the Lalor mine and associated projects are designed to minimize the potential impact on the surrounding environment by keeping the footprint of the operations as small as possible and by using existing licensed facilities for the withdrawal of water and disposal of waste. For instance, the NoA for the expansion of the Anderson TIA was prepared by AECOM and utilized the geotechnical tailings dam designs from Hudbay's ongoing Engineer of Record (performed by BGC Engineering Inc.). The permitted design allowed for the entire volume of tailings from Lalor LOM to be stored in Anderson TIA. This design did not discount the volume of tails that are and will be used in the future for paste backfill at the Lalor mine.

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The initial stage of expansion is complete and subsequent raises are being contemplated based on actual production rates and forecasted life of mine plans. The actual tailings production will be measured against remaining volumes based on ongoing bathymetry surveys, as per the existing licence requirements.

The initial stage of expansion was completed in 2020. The next raise is will be initiated in 2021 and is anticipated to be complete in 2022. The planned tailings facility raises are based upon Lalor's long range plan. All base metal and gold ore tailings are incorporated into the long term storage capacity of Anderson TIA.

New Britannia Mill

The re-commissioning of the New Britannia mill will involve the placement of the Lalor gold ore tailings in the fully licensed Anderson TIA via a pipeline that will connect to the Stall mill. Trade-off studies and environmental assessments are ongoing to ensure that the use of Anderson TIA for all Lalor tailings is the most effective option. These studies include the study of potential water quality impacts of potential increases in New Britannia operating capacity. A NoA for the refurbishment and upgrade of the New Britannia concentrator was submitted in the first quarter of 2019 and was approved by Manitoba regulators in the first quarter of 2020.

Satellite deposits

Future production from the Pen II Zone would be processed in the Stall mill and the tailings would be stored at the Anderson TIA.

20.3. PERMITTING REQUIREMENTS

Lalor Mine, Stall Mill and Paste Plant

The existing Lalor mine EAL was obtained in the first quarter of 2014 and covers all facilities on the Lalor site. These include sewage, water treatment facilities and the pipelines which carry freshwater and wastewater. The Lalor mine was permitted such that it uses to the greatest extent possible, facilities that were already licensed. The sources of freshwater and discharge of wastewater at Hudbay's existing operating sites are licensed provincially. Discharge of effluent to the environment is regulated federally under the Metal and Diamond Mining Effluent Regulations (MDMER).

The permits required for the Lalor operation, including the Lalor mine, Stall concentrator, New Britannia concentrator, and Anderson TIA are current and valid licenses are held. The Anderson TIA expansion NoA was submitted for approval in the third quarter of 2016 and were approved in the second quarter of 2018. No additional tailings impoundment area will be required. Federal permits were not required for this expansion.

Lalor paste plant NoA was submitted in the fourth quarter of 2016 and was approved in the first quarter of 2017.

Environmental baseline studies have been completed by AECOM in the area of the 1901 deposit. This will augment existing environmental studies in the region and will be utilized in permitting the development and infrastructure upgrades needed for this project. The existing infrastructure in the Lalor project area will be utilized to the fullest extent to minimize the need for future construction and minimize disturbances.

New Britannia Mine Mill

The New Britannia site includes the Birch Lake Tailings Disposal Facility (BLTDF). The BLTDF is currently in care and maintenance since the closure of the mine in 2005. The seasonal discharge from BLTDF, which has an EAL, is regulated under the MDMER. Hudbay obtained a new water withdrawal permit for this site in June 2020 to accommodate the anticipated freshwater requirement to process Lalor ore. In order to initiate upgrades needed to process material from the Lalor mine; a NoA was submitted for approval in the first quarter of 2019 and was approved in the first quarter of 2020.

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Satellite deposits

Hudbay will initiate baseline studies and applications will be sent to the appropriate regulatory body for approval sufficiently in advance of when these satellite deposits are anticipated to be constructed and put into production. At this stage, and given Hudbay's permitting experience and continued operation in the area, there is no indication that approvals for development of satellite deposits will not be obtained within the anticipated project schedules.

20.4. COMMUNITY SUPPORT

The main settlement in the region of the Lalor mine is the town of Snow Lake, which is an important mining and service centre for the surrounding area. Snow Lake had a population of 899 according to the 2016 Statistics Canada census. This number is not expected to have significantly changed. The majority of Snow Lake residents are directly employed or support the Lalor mine through contracting services.

Through the years, Hudbay and AECOM have carried out public consultation, including meetings to inform local communities on the progress and development of the Lalor mine, expansion of Anderson TIA, New Britannia concentrator, and the environmental effects of these projects. Manitoba Sustainable Development has taken these meetings into account in the environmental licensing process. Since the economy of Snow Lake is based on mining, opposition to the projects is seen as unlikely.

20.5. FIRST NATIONS

Based on Hudbay's long-term (more than 50 years) mining experience in the Snow Lake region, there is no known current First Nation or Aboriginal hunting, fishing, trapping or other traditional use in the zone of potential influence for the Lalor mine and potential future projects. There is no First Nation Registered Trapline District or Reserve in the area that will be affected by the Lalor operation. Although development on the Mine Site involved a loss of vegetation and habitat for wildlife, the vegetation and habitat type is common throughout the region.

The Mathias Colomb Cree Nation ("MCCN"), located 125 km northwest of Snow Lake at the community of Pukatawagan, has asserted a right to be consulted in connection with recent Hudbay projects in the Snow Lake area since the development of the Lalor Mine. During the permitting process of the New Britannia site, the Manitoba Métis Federation ("MMF") asserted a right for consultation regarding this and future projects in the area. Hudbay continues to keep First Nations and Aboriginal communities informed on potential future projects and will provide Manitoba regulators with all information necessary to support Crown Consultation decisions. Regarding First Nations and Crown Consultation it is anticipated that there will be no impact to current operations or delays in project schedule.

20.6. HERITAGE RESOURCES

Operation of the Lalor mine and construction of potential future upgrades will not affect any known site of potential historical, archaeological or cultural significance. The only known site, Trampling Lake, is located approximately 20 km south of the Lalor operation. This site is of one of Manitoba's largest known concentrations of aboriginal pictographs. These paintings are thought to have been created 1,500 to 3,000 years ago by the Algonkian-speaking ancestors of the Cree and Ojibway First Nations. Activities associated with the Lalor operation will not have any impact on this historical site.

20.7. MINE CLOSURE

The Manitoba Mines and Minerals Act require a closure plan and financial assurance for any advanced exploration or mining project. Manitoba accepted Hudbay's Closure Plans prepared by SRK in 2005 and financial assurance to cover the cost of closure for all existing infrastructure that will continue to be used during operation of the Lalor mine. Existing facilities which support the Lalor mine include the Chisel North mine, which is connected by an underground ramp to Lalor, Stall Concentrator and Anderson TIA piping systems associated with milling and tailings deposition, the Chisel Open Pit and the Chisel North water treatment plant.

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Prior to commercial production at the Lalor mine, Manitoba approved the Closure Plan for the Lalor AEP. The Lalor AEP Closure Plan was prepared in 2010 and approved as part of the AEP application process. As a requirement of the Lalor mine EAL, an updated closure plan was prepared by SRK and was submitted in the third quarter of 2014. It is anticipated that the site of the Lalor mine will be substantially returned to its natural state in about five to ten years post closure, after which no monitoring or other measures will be required.

The expansion of Anderson TIA, and upgrades to the New Britannia site also will require the submission of updated closure plans and financial assurance. Allowances for these applications have been included in the cash flow model supporting reserve declaration.

Post closure, all water quality and earthen structures will be monitored and inspected in order to ensure the sites' conditions meet the applicable regulatory requirements.

Expansions and upgrades of the existing operating facilities and contemplated future developments will further extend the useful life of Hudbay's assets in Northern Manitoba. Additional development of ore bodies will extend operations and deferred the closure and reclamation while these facilities are in continued use.

Satellite Deposits

The Wim and 3 Zone closure costs were estimated by Hudbay and are based on costs from Reed Mine, as the mining method and surface infrastructure requirements are similar. The major tasks for closure are assumed to be: revegetation of disturbed areas, removal of contaminated soils (in such areas where fuel tanks, maintenance shops, etc. were constructed), removal of concrete slabs used for building foundations, removal of electrical infrastructure, and capping of mine accesses (decline and raise openings). Hudbay is familiar with closure plan development for AEPs and active mine sites given its activities in Northern Manitoba in recent years. A standard practice is to leverage existing facilities already permitted and covered under existing closure plans. This reduces the capital requirements for new developments, minimizes potential impacts to new land & waterways, and reduces closure costs. It is estimated that the cost of reclamation of the satellite deposits will be in the order of $0.9M.

Timing of Closure

The site closure of the Flin Flon Metallurgical Complex (FFMC) is tied to Hudbay's LOM of Snow Lake area operations. The FFMC will be kept on care & maintenance until all satellite ore bodies are depleted in the region, potentially allowing for the concentrator and tailings facility to be utilized. Although this is dependent on the production schedule of potential new deposits, it allows for further use of the FFMC asset and reduces the risk of permitting delays in the future. An added benefit of extending the care & maintenance period of the FFMC, is the closure costs are then deferred, allowing for additional time to improve the closure plan activities and reduce financial risk associated with uncertainties tied to future closure activities.

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21. CAPITAL AND OPERATING COSTS

Unless otherwise stated, capital and operating costs are estimated in real 2021 Canadian dollars.

21.1. CAPITAL COSTS

The capital expenditures required to execute the life of mine plan at Lalor, 1901, WIM and 3 Zone can be separated between growth projects represented by the New Britannia mill refurbishment, the construction of a pipeline corridor to support the planned increase in gold production and Stall mill enhancement initiatives and sustaining capital which includes capitalized mine development, the replacement/acquisition of mining equipment and other capitalised expenditures related to mining, milling and environmental activities.

New Britannia Mill Refurbishment and Pipeline Corridor

For the New Britannia Refurbishment Project, the plan includes refurbishment costs for the existing mill building, inclusion of an outdoor crushed ore loading system, pre-aeration and leach tanks, and leach thickener. In addition, a new flotation building is being constructed to house the flotation, thickening, dewatering, and concentrate shed components required to produce copper concentrate. New major equipment purchases include outdoor feeder/hopper and conveyor, carbon screens, cyclones, kiln, acid vessel, strip vessel, EW cell, refinery furnace, DFR flotation circuit, Jameson flotation circuit, thickener, filter press, reagent packages and lime silo, and electrical equipment such as transformers MCCs and an emergency generator.

The pipeline includes three individual pipes to facilitate movement of material between the New Britannia mill and the Stall mill (with adjacent Anderson tailings facility). An 8" pipe will carry tailings from New Britannia to Stall and a 6" pipe will carry copper concentrate from Stall to New Britannia. The copper concentrate produced at Stall will be piped to New Britannia for dewatering purposes. A reclaim water line 4'' will transport water from Stall to New Britannia. This arrangement facilitates environmental monitoring, maintenance, operation and efficiencies improving paste plant quality, and provides for lower concentrate moistures for the copper concentrate.

This project includes industry standard equipment and proven processing technology in a brownfield environment. Construction activities started in Q1 2020 and are 76% completed as of February 2021. Plant commissioning for the early gold mill projected for Q1 2021 and Q3 2021 for the flotation plant, ramp up in Q2 2021 for the early mill and Q4 2021 for the flotation plant. The capital expenditure summarised on Table 21-1

TABLE 21-1: NEW BRITANNIA MILL REFURBISHMENT CAPITAL COSTS

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Since the initial project approval in 2019, the project budget has been increased to accommodate additional scope associated with Stall pipeline and tie-ins, fire and security systems upgrades to meet 2021 regulations and guidance ($8M), the acceleration of the mill completion to produce early gold ($4M) and disruption to construction associated with the impact the global pandemic ($8M).

Sustaining Capital and Capitalized Development

The total sustaining capital cost (Table 21-2) at the Lalor mine, 1901 deposit and satellite deposits is $858M, comprised primarily of capitalized development and mine equipment costs, totaling $517M and $283M, respectively. In addition, milling and other sustaining capital costs total $58M.

Sustaining capital costs for the new mine plan are formulated assuming a focus on base metal ore extraction to preserve gold ore with higher precious metals content for extraction following commissioning of the New Britannia mill. Overall, this approach results in a superior economic value for the Snow Lake Operations due to the significantly higher precious metals recoveries at the New Britannia mill.

The largest component of capitalized development is mechanical drift costs arising from approximately 50 kilometres of development required to access stopes over the life of mines (26 km at Lalor, 6 km at 1901, 9 km at WIM, and 10 km at 3 Zone). These costs are incurred primarily in the earlier years of the mine life. Similarly, various types of raises are required for venting, material handling (ore passes), and egresses with approximately 5,400 metres required over the life of mines (3,600 m at Lalor, 600 m at 1901, 700 m at WIM, and 500 at 3 Zone).

TABLE 21-2: SNOW LAKE CAPITAL EXPENDITURES

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Equipment (new purchase, replacement, and/or rebuild) as well as infrastructure (surface and underground) costs also account for a large component of sustaining capital requirements with life of mine costs of $283M. Replacement and rebuild costs are based on the scheduled useful life of individual equipment as well as to replace contractor provided equipment with Hudbay owned equipment as a measure to reduce operating costs.

Stall mill sustaining capital costs of $6M over the life of mine includes scheduled replacement of key equipment, and building upkeep required to maintain production at up to 3,800 tpd. Similarly, New Britannia mill sustaining capital costs of $16M over the life of mine is scheduled to maintain desired throughput levels. Additionally, environmental capital costs of $33M relate to the scheduled Anderson tailings facility raises and stabilization work expected to be completed over the life of mine, to accommodate tailings from both the Stall and New Britannia mills as a result of regular operations.

21.2. OPERATING COSTS

Operating cost estimates at Lalor are based on the 2021 budget figures and were developed by Hudbay with a bottom-up approach using quotes from local suppliers, Manitoba operations experience, labour costs within the region, and actual 2020 costs. The operating cost estimates are presented below, separately for mining, milling, refining and general and administrative activities.

The operational costs at 1901 are expected to be lower than the costs experienced at Lalor. The costs associated with mining ore (direct mining, ore removal, and backfill) are like Lalor's PPCF method, with the exception that hoisting capacity (and associated costs) seen at Lalor will not be incurred. The overall PPCF mining unit operating costs are assumed to be approximately $12 per tonne lower at 1901, and this is largely due to the difference in general mine expense (GME) costs. Personnel requirements are less, and the indirect costs tied to Lalor infrastructure that is not applicable to 1901 has been excluded in cost estimates.

The operating costs for WIM and 3 Zone reflect an owner-operator scenario for production, and contractor pricing for development similar to the way Hudbay developed the Reed mine until 2018. The longhole mining unit-costs (direct mining, ore removal, backfill) are expected to be higher than experienced at Lalor, due to lower production rates, added haulage distance, and use of smaller equipment to mine the narrow orebodies. Due to proximity to Snow Lake, the mines are operated on a "local mine" basis in which most of the employees work standard business hours and shift work. As such, GME costs are based on Reed's historical costs, with adjustment and escalation (to account for inflation) factors applied.

Mining Operating Costs

Table 21-3 presents the details of the Lalor mine and satellite deposits' operating costs. Labour costs constitute the largest component of the mining operating costs for the life of the operations, and typically include both employee and contractor costs arising from the various mining and maintenance activities. At Lalor, the staffing plan includes approximately 600 positions, comprised primarily of hourly paid employees.

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TABLE 21-3: MINE OPERATING COSTS

GME includes a variety of costs such as power, mine administration, heating, ventilation, level maintenance, engineering, geology, and Snow Lake area costs that are allocated between capitalized development and mining operating costs based on accounting rules. For Lalor and the Snow Lake satellite deposits, this cost represents 30-40% of the total mining operating costs over the life of mines.

In terms of mining activities, ore extraction (direct mining, ore removal, and backfill) represents the largest cost driver accounting for 40-50% of mining operating costs over the life of mines, while mine development activities account for the remaining 20%.

Milling Operating Costs

Table 21-4 presents the details of the Stall and New Britannia operating costs. As with mining, labour costs are the largest component of milling operating costs between the mills used to process Lalor ore. The staffing plan includes a total of approximately 80 positions at Stall mill and 50 at New Britannia mill, comprised primarily of hourly paid employees. Concentrates produced at Stall and New Britannia mills are transported to Flin Flon by truck where they are loaded onto railcars and shipped to off-takers.

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TABLE 21-4: MILLING OPERATING COSTS

Refining Costs

Until the closure of the zinc plant in Q2 2022, the Lalor zinc concentrate will be refined into finished zinc to maximize the realizable value from zinc produced. The operating costs presented in Table 21-5 summarize the costs allocated to Snow Lake Operations for the refining of zinc concentrates produced from Lalor mine's production (either through the Stall or Flin Flon mills).

TABLE 21-5: REFINING OPERATING COSTS

General and Administration Costs

General and administration costs (Table 21-6) are comprised of an allocation of shared services, administration, and other office costs to Snow Lake Operations based on relative usage versus Flin Flon Operations. Significant costs in the shared services line item include health & safety, environmental control, logistics, and analytical services. Significant costs in the administration line item include insurance and property taxes, information technology, human resources, and finance.

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TABLE 21-6: GENERAL AND ADMINISTRATION COSTS

Note: decommissioning, pension, and severance costs are considered as corporate legacy costs and therefore not included in the economic valuation of the business plan of Snow Lake Operations.

21.3. UNIT CASH AND SUSTAINING CASH COSTS

With the benefit of the New Britannia mill, the net revenue at Lalor will shift from primarily zinc to primarily gold post 2021, positioning Lalor as a gold mine with significant zinc, copper, and silver by-products. The life of mine gross revenue from Lalor and the Snow Lake satellite deposits is approximately 57% precious metals, 24% zinc and 19% copper. Significant zinc and copper revenue provides a diversified commodity exposure with a significant upside potential from the mineral resource estimates in the Snow Lake that have not yet been converted to mineral reserve estimates both at Lalor and at other Snow Lake gold properties.

Table 21-7 illustrates that Lalor, the 1901 deposit and the Snow Lake satellite deposit's significant by-product credits will reduce its cash operating costs and sustaining cash costs both on a zinc and gold basis. During the first six years of operation with New Britannia (2022-2027), the Snow Lake operations will produce approximately 182,000 ounces of gold annually at a cash cost, net of by-product credits, of US$411/oz. This positions Lalor to be one of the lowest cost gold mines in Canada.

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TABLE 21-7: CASH AND SUSTAINING UNIT CASH COSTS

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22. ECONOMIC ANALYSIS

Hudbay is a producing issuer and has excluded the information required by Item 22 of Form 43-101F1 as the updated mine plan does not represent a material increase in current production.

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23. ADJACENT PROPERTIES

RockCliff owns the claims located immediately to the North of the PEN II deposit. RockCliff has not reported any mineral resource estimates and no significant exploration results on these claims that would have an impact on the mineral resource estimates reported by Hudbay in this Technical Report for the PEN II property.

There are no other relevant adjacent properties in relation to the other properties discussed in this Technical Report.

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24. OTHER RELEVANT DATA AND INFORMATION

There is no other relevant data or information that we are aware of that has not been presented in the other sections of this Technical Report which is necessary to make this Technical Report understandable and not misleading.

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25. INTERPRETATIONS AND CONCLUSIONS

Lalor was discovered in 2007 on 100% owned land using Hudbay's innovative exploration techniques and it is the largest VMS deposit found in the Snow Lake region to-date. In 2009, the company commenced construction of Lalor with initial ramp access from the Chisel North mine. Construction of the main production shaft was approved in 2010 and was completed on time and on budget, achieving commercial production in 2014, only seven years from initial discovery.

Hudbay acquired the New Britannia gold mill in 2015 for approximately $10 million as a potential long-term processing option for the Lalor gold and copper-gold zones. New Britannia was placed on care and maintenance in 2005 by its previous owner after producing 1.6 million ounces of gold. Since acquiring New Britannia, Hudbay has conducted significant technical work to assess the grade, tonnage, mineability and metallurgy of the gold and copper-gold zones at Lalor to support and de-risk the investment required to refurbish New Britannia and maximize the net present value of Lalor.

During the last two years, Hudbay has also dedicated considerable efforts to drilling and test mining in the gold-rich lenses at Lalor as well as advancing engineering studies on the Snow Lake satellite deposits. In addition, the recent discovery of the 1901 deposit in February 2019 has been quickly followed by exploration and infill drilling to define the mineral resource and mineral reserve estimates presented in this Technical Report. The updated estimate of mineral reserves of the integrated mining and milling complex of Snow Lake has increased in-situ contained gold by 12%, copper by 14%, zinc by 34% and silver by 30%, relative to the previous estimate of mineral reserves in Hudbay's 2020 AIF, after adjustment for 2020 mining production.

The New Britannia refurbishment project includes industry standard equipment and proven processing technology in a brownfield environment. The project is well advance and at overall completion of 76% as of February 2021 with construction completed at 70%, engineering and procurement at 99%, commissioning at 12% and with first gold planned for Q3 2021. The completion of the flotation circuit to process the higher copper grade ore is planned for Q4 2021.

In 2019 and 2020, Hudbay has conducted metallurgical testing work to demonstrate the technical and economic viability of several modifications to the Stall mill flowsheet including an increase in copper rougher and cleaner capacity with the addition of Jameson cells, the addition of talc preflotation circuit with the re-configuration of existing flotation cells and addition of a Jameson cell cleaner stage, an increase the zinc circuit cleaning capacity and froth washing through the addition of Jameson cells, and an increase in recovery of free gold through the addition of a Knelson gravity concentrator on the copper regrind cyclone underflow. Implementation of these modifications will be completed by end 2022 for a total investment estimated at $25m with the resulting benefits planned to start in 2023.

The revised mine plan for the Lalor mineral reserves optimizes net present value by maximizing gold-rich ore for processing at the New Britannia mill and zinc-rich ore for the Stall mill, which is expected to result in significantly higher gold and copper recoveries. This revised mine plan for the Snow Lake operations supports a 17 year mine life, based solely on proven and probable reserves, and utilizes the existing mining capacity of 5,300 tonnes per day at Lalor and 1901 until 2027 and 4.650 tonnes per day for 2028 and 2029 before ramping down towards an average of 1,350 tonnes per day for the rest of the mine life. The technical work completed supports 5,300 tonnes per day as the optimal mining rate to maximize net present value, although the Lalor production shaft has the potential to hoist at higher throughput rates.

With the inclusion of the New Britannia mill, net revenue in Snow Lake will shift from primarily zinc to primarily gold positioning the integrated operations as a primary gold production center with significant zinc, copper and silver by-products. Lalor and 1901 significant by-product credits reduce its cash operating costs and sustaining cash costs on both a zinc and gold basis. During the first five years of operation with New Britannia (2022 to 2026), Recovered annual gold production will average approximately 180,000 ounces of gold at a cash cost, net of by-product credits, of $386/oz, positioning Lalor as one of the lowest cost gold mines in Canada. In addition, significant zinc and copper revenue provides diversified commodity exposure.

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Refurbishing New Britannia is expected to also enable new gold and copper-gold exploration opportunities in the Snow Lake region by having an operating processing facility with substantially higher gold and copper recoveries. It also demonstrates the opportunity to create additional value through owning multiple processing facilities in the Flin Flon and Snow Lake regions.

Since 2019, Hudbay has also implemented a more stringent approach to resource reporting for underground deposits. With this approach, the potential for economic extraction of the mineral resource estimates at Lalor and all the satellite deposits with the exception of the former New Britannia mine, are reported within the constraint of a 'stope optimization envelope' process similar in concept to a Lerchs-Grossman pit shell for an open pit deposit. This excludes from the resource estimate small individual resource blocks that may meet economic cut-off criteria on an individual basis but could not be aggregated into mineable shapes. It is anticipated that this approach will result in higher resource to reserve conversion ratios than those experienced in the past.

The updated resource models at Lalor and 1901 includes 6.5 million tonnes of inferred mineral resources, which has the potential to extend the period of the mine life when both Stall and New Britannia mills are fed at their current combined demonstrated throughput capacity of 5,300 tonnes per day. In addition, the mineral resources at Hudbay's satellite deposits in the Snow Lake region, including the 1901, Watts and Pen deposits and the former gold producing New Britannia mine could provide feed for the Stall and New Britannia processing facilities and further extend the mine life and/or maximise the utilization of the 5,300 tonnes per day processing capacity over a longer period.

In 2020, Hudbay has continued to invest in significant exploration efforts in Manitoba including major airborne and ground geophysical surveys and surface exploration drilling in the Flin Flon and Snow Lake areas. This work has identified several base metal and gold targets to be tested in 2021. In parallel, in-mine exploration at Lalor continues with the intent to convert inferred resources to indicated mineral resources and add additional inferred mineral resources in base metal and gold-rich mineralization. Hudbay will continue to advance engineering studies on Lalor and its satellite deposits in an attempt to continue to increase the tonnage of the mineral reserve estimates and the estimated operating life of the Snow Lake processing facilities at or near full capacity.

The Hudbay validation process has confirmed that the resource models are constructed using industry accepted modelling techniques and classified in accordance with the 2014 CIM Definition Standards - For Mineral Resources and Mineral Reserves. The author considers that the mineral resources and reserves as classified and reported comply with all disclosure in accordance with requirements and CIM Definitions. The author is not aware of any mining, metallurgical, infrastructure, permitting or other relevant factors that could materially affect the mineral reserve estimate.

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26. RECOMMENDATIONS

Significant value has been unlocked by leveraging Hudbay's exploration, technical, and operational expertise, its existing portfolio of processing infrastructures and several inexpensive acquisitions. These efforts have highlighted the potential to further increase the tonnage of mineral reserve estimates and the estimated operating life of the Snow Lake processing facilities at or near full capacity through continued exploration success, a disciplined and methodic approach to engineering studies while maintaining excellence in tailings management and environmental stewardship.

Hudbay has developed a well focused surface exploration program aimed at confirming near surface exploration targets between Lalor and the past producing Chisel mine and around the past producing Snow Lake gold properties. The author supports this program which has the potential to define additional inferred resources in the Snow Lake camp over the next one to two years.

Hudbay has in-mine and regional annual exploration budget of approximately $10M to cover the cost of these drilling programs proposed for Lalor and the satellite deposits.

As the Lalor mine matures and additional satellite ore sources are identified to supplement the feed blend to the Stall and new Britannia mills, metallurgical projects are actively advanced to match the configuration of the mill to the requirement of the ore. The following are projects that are being advanced through metallurgical testwork, early stage engineering or both.

The current life of mine plan for Lalor shows an increasing Pb:Cu ratio from 0.3 to >1.0. In addition, the 1901 deposit has a Pb:Cu ratio of >1.5 indicating the need for a lead mitigation strategy. Preliminary results indicate that a traditional sequential Cu-Pb-Zn flotation circuit at Stall will mitigate the increasing lead in feed and provide an additional lead concentrate revenue stream that captures a significant portion of the copper, gold and silver that is lost into the zinc concentrate into a payable stream.

The flotation recovery of gold at Stall is limited by the mineralogy and association of gold with iron sulfides. Due to the mineralogical differences between the gold zone and base metal zone ores, the flotation tailings from the base metal ores are not compatible with the same leaching conditions as the gold zone ore flotation tailings. Metallurgical testwork is on going for this project.

Alongside the tailings Leach project, a fine grinding project is being advanced to determine the improvements in flotation recoveries when the flotation feed size is reduced from the current p80 of 120µm to 60µm.

Historical production at the New Britannia mill saw throughput rates of up to 2,000tpd based on the availability of ore from the underground operation. With the increase in the mineral reserves of the Lalor gold zones alongside additional satellite deposits, an opportunity exists to increase the throughput rate. The study to date has identified the tailings pipeline, pre-leach thickener and various minor pumping systems require upgrade to accommodate the increased tonnage rate. All the other major systems within the New Britannia mill have sufficient capacity to accommodate the increased tonnage. In addition to the processing equipment upgrades the study includes the upgrade of the pipeline corridor road to accommodate ore haulage from Stall to New Britannia to remove traffic from the public access road.

26-180

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

The 1901 deposit contains two gold-copper lenses which are suitable feed for the New Britannia mill and would supplement the Lalor gold zone ore. A scoping study has been initiated to advance this project through metallurgical testwork on composites representative of the two gold lenses for which inferred mineral resource estimates have been reported for the 1901 deposit.

26-181

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

27. REFERENCES

AECOM, Lalor Mine Environmental Act Proposal Report, Weiss, A., Samoiloff, C., Biswanger, S., 2012.
AECOM, Internal Report to Hudbay: Environmental Assessment of Proposed Expansion of the Anderson Tailings Impoundment Area, Sadiq, S., Ryman, J., Kjartanson, S., 2016.
AECOM, Environmental Assessment of Proposed Expansion of the Lalor Mine to add the Lalor Paste Plant, Davis, I., Kjartanson, S., Samoiloff, C., Sadiq, S., 2016.
Agoratek International Consultants inc., Lalor mine review of sampling issues, D. François-Bongraçon, 2019.
AECOM, Notice of Alteration: New Britannia Mill. Cusitar, K., Davies, I., Gilfford, M., 2019.
Bailes, A.H., and Galley, A.G., 1999. Evolution of the Paleoproterozoic Snow Lake arc assemblage and geodynamic setting for associated volcanic-hosted massive sulphide deposits, Flin Flon Belt, Manitoba, Canada. Canadian Journal of Earth Sciences, 36, pp. 1789-1805.
Bailes, A.H., and Galley, A.G., 2007. Geology of the Chisel-Anderson Lakes Area, Snow Lake, Manitoba (NTS areas 63K16SW and west half of 63J13SE). Manitoba Science, Technology, Energy and Mines, Manitoba Geological Survey, Geoscientific Map MAP2007-1, scale 1: 20,000 plus notes.
Beilhartz, D., 2006.Report on the Squall Lake Property, The Pas Mining Division, Snow Lake, Manitoba for Garson Resources Ltd.
BGC Engineering Inc. Anderson Tailings Impoundment Area: Pre-Feasibility Study. Report 0077-018-01, 2011.
BGC Engineering Inc. Anderson Tailings Impoundment Area: Feasibility Study. Report 0077-022-02, 2012.
BGC Engineering Inc. Anderson Tailings Impoundment Area: Deposition Area. Report 0077-027-02, 2013.
Boge & Boge Ltd, Internal Report to Hudbay: Debottlenecking Concept Study - Stall Concentrator, March 2017.
CIM Definition Standards on Mineral Resources and Mineral Reserves, May 10, 2014, website web.cim.org/standards
David, J., Bailes, A.H., and Machado, N., 1996: Evolution of the Snow Lake portion of the Paleoproterozoic Flin Flon and Kisseynew belts, Trans-Hudson Orogen, Manitoba, Canada. Precambrian Research, 80, pp. 107-124.
Franklin, J.M., Gibson, H.L., Jonasson, I.R., and Galley, A.G., 2005. Volcanogenic Massive Sulfide Deposits, in Hedenquist, J. W., Thompson, J. F. H., Goldfarb, R. J., and Richards, J. P., eds., Economic Geology 100th Anniversary Volume: Littleton, CO, Society of Economic Geologists, pp. 523-560.
Froese, E., and Moore, J.M., 1980. Metamorphism in the Snow Lake area, Manitoba; Geological Survey of Canada, Paper 78-27, p. 16.
Galley, A.G., Bailes, A.H., Syme, E.C., Bleeker, W., Macek, J.J. and Gordon, T.M. 1991: Geology and mineral deposits of the Flin Flon and Thompson belts, Manitoba (Field Trip 10); Geological Survey of Canada, Open File 2165, 8th IAGOD Symposium Field Trip Guidebook, 136 p.
Genivar, Feasibility Study, Snow Lake Mine Re-activation Project, Technical Report NI 43-101. A. Roy, J. Lavine, D. West, I.R. Ward, M.R. Parfitt, M.J. Bednarz, 2010.
Golder Associates, 2015. Technical Report on the Wim Copper-Gold deposit, Snow Lake, Manitoba, Canada. Effective Date: May 14, 2015, Hudbay pp. 1-83, filed on sedar.com

27-182

Lalor and Snow Lake Operations, Manitoba, Canada<br>Form 43-101F1 Technical Report

Golder Associates, Internal Report to Hudbay: Lalor Mine Paste Backfill System Snow Lake, Manitoba, March 2017.
Goldfarb, Richard, et al. "Distribution, character and genesis of gold deposits in metamorphic terranes." Society of Economic Geologists, 2005 (pp. 407-450).
Groves, David I., et al. "Orogenic gold deposits: a proposed classification in the context of their crustal distribution and relationship to other gold deposit types." Ore geology reviews 13.1-5, 1998 (pp. 7-27).
Groves, David I., et al. "Late-kinematic timing of orogenic gold deposits and significance for computer-based exploration techniques with emphasis on the Yilgarn Block, Western Australia." Ore Geology Reviews 17.1-2, 2000 (pp.1-38).
Hudbay Minerals, 2012: Pre-Feasibility Study Technical Report on the Lalor Deposit, Snow Lake, Manitoba, Canada Effective Date: March 29, 2012, Hudbay pp. 1-292, filed on sedar.com
Hudbay Minerals: Technical Report on the Lalor Mine, Snow Lake, Manitoba, Canada Effective Date: March 30, 2017, Hudbay pp. 1-270, filed on sedar.com
Lavoie. S., 2003. Géologie de la Mine East-Sullivan, Abitibi-Est, Val-d'Or, Québec. Université du Québec à Chicoutimi
Lindgren, W., 1933, Mineral deposits (fourth edition): New York, McGraw-Hill Book Co., 930 p.
Mercier-Langevin, P., 2009. Field Notes and Observations Lalor deposit, Snow Lake April 22-25th, 2009.
Parrish, I.S., 1997. Mining Engineering Journal, Geologist's Gordon Knot: to cut or not to cut (pp. 45-49)
Rubingh, K.E., Lafrance, B. and Gibson, H.L. 2013: Structural analysis of the McLeod Road-Birch Lake thrust panel, Snow Lake, west-central Manitoba (parts of NTS 63K16, 63J13); in Report of Activities 2013, Manitoba Mineral Resources, Manitoba Geological Survey, p. 106-113.
Syme, E.C., Lucas, S.B., Zwanzig, H.V., Bailes, A.H., Ashton, K.E., and Haidl, F.M., 1998: Geology, NATMAP Shield Margin Project area, Flin Flon Belt, Manitoba-Saskatchewan; Manitoba Energy and Mines, Geological Services, Map A-98-2, scale 1:100 000 and 1:350 000, with accompanying notes, 54 p.
WSP, Hudbay Minerals New Britannia Resource Estimate, report 181-15173-00, T. McCracken, 2019.
Alford Mining Systems, Stope Shape Optimiser Version 4.0.0, Reference Manual, December 2019, p. 1.

27-183

Hudbay Minerals Inc.: Exhibit 99.2 - Filed by newsfilecorp.com

OLIVIER TAVCHANDJIAN

CERTIFICATE OF QUALIFICATION

Re: NI 43-101 Technical Report, Lalor and Snow Lake Operations,

Manitoba, Canada March 29^th^ 2021

I, Olivier Tavchandjian, P. Geo., of Toronto, Ontario, do hereby certify that:

I am currently employed as Vice President, Exploration and Geology, with Hudbay Minerals Inc., 25 York Street, Suite 800, Toronto, Ontario, Canada, M5J 2V5.
I graduated from the Université du Québec à Chicoutimi with a Ph.D. in Mineral Resources in 1996.
I am a member in good standing with the Association of Professional Geoscientists of Ontario, member 0714.
I have practiced my profession continuously for over 30 years and have been involved in mineral exploration, project evaluation, resource and reserve evaluation, and technical support to mine operations in underground and open pit mines for base metals, precious metals and bulk commodities in the Americas, Europe, Asia and Africa.
I have read the definition of "qualified person" set out in National Instrument 43-101 ("NI 43-101") and certify that by reason of my education and affiliation with a professional association and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purpose of NI 43-101.
I have reviewed and approved the Summary of the Technical Report and I am responsible for the preparation of this Technical Report titled "NI 43-101 Technical Report, Lalor and Snow Lake Operations, Manitoba, Canada" (the "Technical Report"), dated March 29^th^ 2021 and effective as of January 1^st^ 2021.
I last visited the Lalor Mine on January 15^th^ 2020, for a duration of two (2) days. I also visited the mine and the other Snow Lake operations several times prior to that date.
As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
I am not independent of the Issuer. Since I am an employee of the Issuer, a producing issuer, I fall under subsection 5.3(3) of NI 43-101 where "a technical report required to be filed by a producing issuer is not required to be prepared by or under supervision of an independent qualified person".
I have been involved with the Lalor Mine and Snow Lake property, which is the subject of the Technical Report, continuously since September 2017.
I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with the instrument and form.
I consent to the public filing of the Technical Report with any stock exchange, securities commission or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated as of March 29^th^ 2021.

Original signed by:

/s/ Olivier Tavchandjian

Olivier Tavchandjian, P. Geo.

Hudbay Minerals Inc.: Exhibit 99.3 - Filed by newsfilecorp.com

CONSENT OF QUALIFIED PERSON

March 29, 2021

British Columbia Securities Commission Alberta Securities Commission Financial and Consumer Affairs Authority of Saskatchewan The Manitoba Securities Commission Ontario Securities Commission Autorité des marchés financiers New Brunswick Financial and Consumer Services Commission Nova Scotia Securities Commission Office of the Superintendent of Securities, Prince Edward Island Office of the Superintendent of Securities, Service Newfoundland and Labrador Office of the Superintendent of Securities, Northwest Territories Office of the Yukon Superintendent of Securities Nunavut Securities Office

Dear Sir/Madam:

Re: Technical Report titled "NI 43-101 Technical Report, Lalor and Snow Lake Operations, Manitoba, Canada" dated March 29, 2021, with an effective date of January 1, 2021

I, Olivier Tavchandjian, consent to the public filing of the technical report titled "NI 43-101 Technical Report, Lalor and Snow Lake Operations, Manitoba, Canada" dated March 29, 2021, with an effective date of January 1, 2021 (the "Technical Report") by Hudbay Minerals Inc. ("Hudbay").

I also consent to the use of any extracts from, or a summary of, the Technical Report in Hudbay's news release dated March 29, 2021 and Annual Information Form dated March 29, 2021 (the "Disclosure Documents").

I certify that I have read the Disclosure Documents and that they each fairly and accurately represent the information in the Technical Report for which I am responsible.

/s/ Olivier Tavchandjian

Olivier Tavchandjian P. Geo.

Vice President, Exploration and Geology, Hudbay

Hudbay Minerals Inc.: Exhibit 99.4 - Filed by newsfilecorp.com

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

CAUTIONARY NOTE REGARDING FORWARD LOOKING INFORMATION

This National Instrument ("NI") 43-101 Technical Report (this "Technical Report ") contains "forward-looking statements" and "forward-looking information" (collectively, "forward-looking information") within the meaning of applicable Canadian and United States securities legislation. All information contained in this Technical Report, other than statements of current and historical fact, is forward-looking information. Often, but not always, forward-looking information can be identified by the use of words such as "plans", "expects", "budget", "guidance", "scheduled", "estimates", "forecasts", "strategy", "target", "intends", "objective", "goal", "understands", "anticipates" and "believes" (and variations of these or similar words) and statements that certain actions, events or results "may", "could", "would", "should", "might" "occur" or "be achieved" or "will be taken" (and variations of these or similar expressions). All of the forward-looking information in this Technical Report is qualified by this cautionary note.

Forward-looking information includes, but is not limited to, our objectives, strategies, intentions and expectations, production, cost, capital and exploration expenditure guidance and potential revisions to such guidance, anticipated production at our mines and processing facilities, expectations regarding the impact of the COVID-19 pandemic on our operations, financial condition and prospects, and our ability to effectively engage with local communities in Peru and other stakeholders, expectations regarding the timing of mining activities at the Pampacancha deposit and any additional delivery obligations under the Constancia stream agreement, the potential and our anticipated plans for advancing our mining properties surrounding Constancia and elsewhere in Peru, anticipated production at our Constancia mine and processing facilities and events that may affect Hudbay's operations, anticipated improvements to metallurgical recoveries, anticipated cash flows from operations and related liquidity requirements, the anticipated effect of external factors on revenue, such as commodity prices, estimation of mineral reserves and resources, mine life projections, reclamation costs, economic outlook, government regulation of mining operations, and expectations regarding labour and community relations. Forward-looking information is not, and cannot be, a guarantee of future results or events. Forward-looking information is based on, among other things, opinions, assumptions, estimates and analyses that, while considered reasonable by us at the date the forward-looking information is provided, inherently are subject to significant risks, uncertainties, contingencies and other factors that may cause actual results and events to be materially different from those expressed or implied by the forward-looking information.

our ability to continue to operate safely and at full capacity during the COVID-19 pandemic;
the availability, global supply and effectiveness of COVID-19 vaccines, the effective distribution of such vaccines in the countries in which we operate, the lessening of restrictions related to COVID-19, and the anticipated rate and timing for each of the foregoing;
our ability to achieve production and unit cost guidance;
no significant interruptions to our operations or significant delays to our development projects in Manitoba and Peru due to the COVID-19 pandemic;
the availability of spending reductions and liquidity options;
the timing of development and production activities on the Pampacancha deposit;
the timing for reaching additional agreements with individual community members and no significant unanticipated delays to the development of Pampacancha
the successful renegotiation of collective agreements with the labour unions that represent certain of our employees in Manitoba and Peru;
the success of mining, processing, exploration and development activities;
the scheduled maintenance and availability of our processing facilities;
the accuracy of geological, mining and metallurgical estimates;
anticipated metals prices and the costs capital projects and production;

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

the supply and demand for metals we produce;
the supply and availability of concentrate for our processing facilities;
the supply and availability of third party processing facilities for our concentrate;
the supply and availability of all forms of energy and fuels at reasonable prices;
the availability of transportation services at reasonable prices;
no significant unanticipated operational or technical difficulties;
the execution of our business and growth strategies, including the success of our strategic investments and initiatives;
the availability of additional financing, if needed;
the availability of personnel for our exploration, development and operational projects and ongoing employee relations;
maintaining good relations with the labour unions that represent certain of our employees in Peru and no labour disruptions;
maintaining good relations with the communities surrounding Constancia, including the neighbouring Indigenous communities and local governments;
no significant unanticipated challenges with stakeholders at our various projects;
no significant unanticipated events or changes relating to regulatory, environmental, health and safety matters;
no contests over title to our properties, including as a result of rights or claimed rights of Indigenous peoples or challenges to the validity of our unpatented mining claims;
no significant unanticipated litigation;
certain tax matters, including, but not limited to current tax laws and regulations and the refund of certain value added taxes from the Peruvian government; and
no significant and continuing adverse changes in general economic conditions or conditions in the financial markets (including commodity prices and foreign exchange rates).

The risks, uncertainties, contingencies and other factors that may cause actual results to differ materially from those expressed or implied by the forward-looking information may include, but are not limited to, the political situation in Peru, risks generally associated with the mining industry, such as economic factors (including future commodity prices, currency fluctuations, energy prices and general cost escalation), risks related to the new Constancia mine plan, risks related to the schedule for mining the Pampacancha deposit (including risks associated with COVID-19, with reaching additional agreements with individual community members and the impact of any schedule delays), dependence on key personnel and employee and union relations, risks related to political or social unrest or change, risks in respect of Indigenous and community relations, rights and title claims, operational risks and hazards, including unanticipated environmental, industrial and geological events and developments and the inability to insure against all risks, failure of plant, equipment, processes, transportation and other infrastructure to operate as anticipated, planned infrastructure improvements in Peru not being completed on schedule or as planned, compliance with government and environmental regulations, including permitting requirements and anti-bribery legislation, depletion of Hudbay's reserves, volatile financial markets that may affect our ability to obtain additional financing on acceptable terms, the failure to obtain required approvals or clearances from government authorities on a timely basis, uncertainties related to the geology, continuity, grade and estimates of mineral reserves and resources, and the potential for variations in grade and recovery rates, uncertain costs of reclamation activities, Hudbay's ability to comply with its pension and other post-retirement obligations, our ability to abide by the covenants in our debt instruments and other material contracts, tax refunds, hedging transactions, as well as the risks discussed under the heading "Risk Factors" in our most recent annual information form dated March 29, 2021.

Should one or more risk, uncertainty, contingency or other factor materialize or should any factor or assumption prove incorrect, actual results could vary materially from those expressed or implied in the forward-looking information. Accordingly, you should not place undue reliance on forward-looking information. We do not assume any obligation to update or revise any forward-looking information after the date of this Technical Report or to explain any material difference between subsequent actual events and any forward-looking information, except as required by applicable law.

iii

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

SIGNATURE PAGE

This Technical Report titled "NI 43-101 Technical Report, Constancia Mine, Cuzco, Peru", dated March 29^th^ 2021 and effective as of January 1^st^ 2021, was prepared under the supervision and signed by the following author:

Dated this 29^th^ day of March 2021.

/s/ Olivier Tavchandjian Signature of Qualified Person

Olivier Tavchandjian, P. Geo.

Vice President, Exploration and Geology

Hudbay Minerals Inc.

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

TABLE OF CONTENTS

1. Summary	1-1
1.1. Property Description and Location	1-1
1.2. Geological Setting and Mineralization	1-2
Constancia	1-2
Pampacancha	1-3
1.3. Exploration	1-3
1.4. Drilling	1-4
1.5. Sample Preparation, Analyses, and Data Validation	1-4
1.6. Mineral Processing and Metallurgical Testing	1-5
1.7. Mineral Resource Estimates	1-6
Constancia	1-6
Pampacancha	1-8
1.8. Mineral Reserve Estimates	1-9
1.9. Mining Methods	1-10
Mine Phases	1-11
Mine Production Schedule	1-12
Waste Rock Facility (WRF)	1-13
Ore Stockpiles	1-13
Mine Equipment	1-15
Mine Engineering	1-15
1.10. Recovery Methods	1-15
1.11. Project Infrastructure	1-15
1.12. Market Studies and Contracts	1-17
1.13. Environmental Studies, Permitting and Social or Community Impact	1-18
1.14. Capital and Operating Cost	1-19
1.15. Economic Analysis	1-20
1.16. Conclusions	1-21
1.17. Recommendations	1-21
2. Introduction	2-22
2.1. Qualified Person	2-22
2.2. Source of Information	2-22
2.3. Unit of Measure Abbreviations Used in Report	2-24
2.4. Acronyms and Abbreviations Used in report	2-25
3. Reliance on Other Experts	3-26
4. Property Description and Location	4-27
4.1. The Property Location	4-27

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
4.2. Mineral Rights	4-27
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4.2.1. Peruvian Mining Law	4-27
4.2.2. Mining Concessions	4-29
Recently Acquired Mining Concessions	4-32
4.2.3. Surface Legal Right Access	4-34
4.3. The Terms of any Royalty or Payments	4-36
4.4. Environmental Property Liabilities	4-36
5. Accessibility, Climate, Local Resources, Infrastructure and Physiography	5-37
5.1. Accessibility	5-37
5.2. Climate	5-38
5.3. Local Resource	5-39
5.4. Infrastructure	5-39
5.5. Physiography	5-39
6. History	6-40
6.1. Early History	6-40
6.2. Prior Ownership	6-40
6.3. Hudbay Ownership and Project Development	6-40
6.4. Production History	6-41
7. Geological Setting and Mineralization	7-42
7.1. District Geology	7-42
7.2. Property Geology	7-42
7.2.1. Constancia	7-42
Stratigraphy	7-42
Intrusions	7-44
Structural Geology	7-45
7.2.2. Pampacancha	7-45
Stratigraphy	7-45
Intrusions	7-45
Structural Geology	7-46
7.3. Alteration	7-46
7.3.1. Constancia	7-46
Potassic Alteration	7- 46
Propylitic Alteration	7- 46
Phyllic Alteration	7- 46
7.3.2. Pampacancha	7-46
Skarn Alteration	7- 46
Potassic Alteration	7- 46
7.4. Mineralization	7-47
7.4.1. Constancia	7-47
7.4.2. Pampacancha	7-47

vii

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
8. Deposit Types	8-49
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9. Exploration	9-50
Caballito	9-52
Maria Reyna	9-52
Quehuincha Norte	9-52
Kusiorcco	9-52
10. Drilling	10-42
10.1. Overview	10-54
10.2. Rio Tinto (2003-2004)	10-55
10.3. Norsemont Mining (2005-2010)	10-55
10.4. Hudbay (2011-2020)	10-56
11. Sample Preparation, Analyses and Security	11-57
11.1. Core Logging	11-57
11.2. Sample Selection	11-57
11.3. Core Photographs	11-58
11.4. Core Splitting	11-58
11.5. Sample Dispatching	11-58
11.6. Sample Preparation	11-58
11.6.1. Historical	11-58
11.6.2. Hudbay	11-59
11.7. Bulk Density	11-60
11.7.1. Historical	11-60
11.8. Assay Methodology	11-61
11.8.1. Rio Tinto	11-61
11.8.2. Norsemont	11-61
11.8.3. Hudbay	11-61
Constancia	11-61
Pampacancha	11-61
11.9. Data Validation	11-62
11.9.1. Drill Hole Database	11-62
Drill Hole Collars	11-62
Downhole Surveys	11- 62
Data Security	11- 62
11.10. Quality Assurance and Quality Control Programs	11-62
11.10.1. Historical	11-62
11.10.2. Norsemont	11-63
11.10.3. Hudbay	11-64
Constancia	11-64
Pampacancha	11-65
11.11. Site Visits	11-66

viii

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
12. Data Validation	12-67
---	---
12.1. Constancia	12-67
12.1.1. Norsemont Quality Control Protocol and Results	12-67
12.1.2. Hudbay Quality Control Protocol and Results	12-69
12.1.3. 2017 Drill Program	12-70
12.1.4. 2019 Drill Program	12-71
12.1.5. 2020 Drill Program	12-71
12.1.6. Core Recovery	12-72
12.1.7. Drilling Data Base	12-72
12.2. Pampacancha	12-73
12.2.1. Quality Assurance / Quality Control	12-73
13. Mineral Processing and Metallurgical Testing	13-75
13.1. Plant Treatment Capacity	13-75
13.2. Plant Recovery	13-77
13.2.1. Ore Types	13-77
13.2.2. Empirical Recovery Models	13-78
13.2.2.1. Hypogene Ores	13-78
13.2.2.2. Mixed and Supergene Ores	13-78
13.2.2.3. Skarn and Hi Zinc Ores	13-78
13.2.3. Flotation Circuit Simulation	13-79
13.3. Test Work	13-79
13.3.1. Constancia Ore Body	13-79
13.3.1.1. DFS Test Work	13-79
13.3.1.2. Post-Production Test Work	13-79
13.3.2. Pampacancha Ore Body	13-82
13.3.2.1. Ore Types	13-82
13.3.2.2. Metallurgical Tests	13-83
13.4. Concentrate Quality	13-88
14. Mineral Resources Estimates	14-89
14.1. Constancia Mine	14-89
Modeling Database	14-89
Modeling of the Mineralized Domains	14-89
Correction factors	14-91
Specific Gravity	14-92
Compositing	14-92
Exploratory Data Analysis	14-93
Grade Capping	14-93
Variography	14-93
Grade Estimation and Interpolation Methods	14-94
Grade Estimation Validation	14-94

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
Visual Inspection	14-95
---	---
Global Bias Checks	14-95
Smoothing Assessment	14-96
Smoothing Correction	14-96
Reconciliation	14-97
Classification of Mineral Resource	14-98
Reasonable Prospects of Economics Extraction and Mineral Resource Estimates	14-100
14.2. Pampacancha	14-100
Modeling Database	14-100
Modeling of the Mineralized Domains	14- 101
Specific Gravity	14-101
Compositing	14-101
Grade Capping	14-102
Variography	14-102
Grade Estimation and Interpolation Methods	14- 102
Grade Estimation Validation	14- 102
Visual Inspection	14- 102
Global Bias Checks	14- 103
Classification of Mineral Resource	14-103
Reasonable Prospects of Economics Extraction and Mineral Resource Estimates	14-103
14.3. Mineral Resource Statement	14-104
15. Mineral Reserves	15-106
15.1. Pit Optimization	15-106
Resource Models	15-106
Net Smelter Return	15-106
Optimization Results	15-108
Pit Design Criteria	15-108
15.2. Mineral Reserves	15-110
16. Mining Methods	16-112
16.1. Mine Overview	16-112
16.2. Mine Phases	16-113
16.2.1. Design Criteria	16-113
Pit Slope Guidance	16-113
16.2.2. Mine Phases and Ultimate Pit	16-117
Constancia - Mine Phase 1	16-118
Constancia - Mine Phase 2	16-118
Constancia - Mine Phase 3	16-118
Constancia - Mine Phase 4	16-119
Constancia - Mine Phase 5	16-120
Constancia - Mine Phase 6	16-121

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
Constancia - Mine Phase 7	16-122
---	---
Constancia - Mine Phase 8	16-123
Constancia - Mine Phase 9	16-124
Pampacancha - Mine Phase 1	16-125
Pampacancha - Mine Phase 2	16-125
16.3. Mine Production Schedule	16-126
16.3.1. Production Scheduling Criteria	16-126
16.3.2. Mill Feed and Cut-Off Grade Strategy	16-126
16.3.3. Overburden Stripping Requirements	16-128
16.3.4. Mine Plan	16-128
16.4. Mine Facilities	16-137
16.4.1. Waste Rock Facility (WRF)	16-137
16.4.2. Ore Stockpiles	16-137
16.4.3. Tailings Management Facility (TMF)	16-138
16.5. Mine Equipment	16-139
Equipment Operating Parameter	16-139
Mine Equipment Calculation	16-140
16.6. Mine Operations	16-141
16.6.1. Drilling	16-141
Drill Patterns and Parameters	16-142
16.6.2. Blasting	16-142
16.6.3. Loading	16-142
16.6.4. Hauling	16-142
16.6.5. Support Equipment	16-143
16.7. Mine Engineering	16-144
16.7.1. Hydrogeological Parameters	16-144
Constancia Pit	16-144
Pampacancha Pit	16-145
17. Recovery Methods	17-146
17.1. Introduction	17-146
17.1.1. General Layout and Facility Description	17-146
17.1.2. Buildings	17-149
17.2. Processing Plant	17-149
17.2.1. General	17-149
17.2.2. Primary Crushing	17-150
17.2.3. Crushed Ore Stockpile and Reclaim System	17-151
17.2.4. Grinding	17-152
17.2.5. Cu-Mo Bulk Flotation	17-153
17.2.6. Cu-Mo Flotation Arrangement	17-154
17.2.7. Molybdenum Separation	17-154

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
17.2.8. Copper Concentrate Thickening and Filtration	17-156
---	---
17.2.9. Copper Concentrate Storage and Loadout	17-157
17.2.10. Molybdenum Concentrate Thickening, Filtration and Loadout	17-157
17.2.11. Tailings Disposal	17-158
17.2.12. Reagents	17-158
17.3. Typical Operating Parameters	17-159
17.4. Metallurgical Accounting and Control	17-159
17.5. Optimization Projects	17-160
17.5.1. Pebble Crushing	17-160
17.5.2. Grinding	17-161
17.5.3. Bulk Flotation	17-162
17.5.3.1. Copper Recovery	17-162
17.5.3.2. Copper Concentrate Quality	17-163
17.5.4. Copper/Molybdenum Separation	17-163
17.5.5. Other Projects	17-164
18. Project Infrastructure	18-165
18.1. Tailings, Waste and Water Storage	18-165
18.1.1. Tailings Management Facility (TMF)	18-165
18.1.2. Waste Rock Facility (WRF), Topsoil and Ponds	18-166
18.1.3. Water Management	18-166
18.2. On Site Camps	18-167
18.3. Electric Power Supply and Transmission	18-167
18.3.1. Power Supply	18-167
18.3.2. Power Transmission	18-167
18.3.3. Current Status	18-167
18.4. Roads and Port	18-168
18.4.1. Roads	18-168
18.4.2. Port	18-169
19. Market Studies and Contracts	19-171
19.1. Concentrate Marketing	19-171
Copper Concentrates	19-171
Molybdenum Concentrate	19-171
19.2. Contracts	19-171
20. Environmental Studies, Permitting and Social Or Community Impact	20-172
20.1. Environmental	20-172
20.1.1. Legal framework	20-172
20.1.2. Local and Regional Context	20-173
20.1.3. Environmental Setting	20-173
20.1.4. Environmental Certifications	20-174
20.1.5. Environmental Management	20-175

xii

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
20.1.6. Mine Closure Plan	20-176
---	---
20.2. Permitting	20-176
20.3. Community Relations	20-177
Pampacancha	20-178
21. Capital and Operating Costs	21-179
21.1. Capital Costs	21-179
21.2. Operating Costs	21-180
22. Economic Analysis	22-182
23. Adjacent Properties	23-183
24. Other Relevant Data and Information	24-185
25. Conclusions	25-185
26. Recommendations	26-186
27. References	27-187

xiii

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

LIST OF FIGURES

Figure 1-1: Constancia Mine Location	1-2
Figure 1-2: Constancia and Pampacancha Ultimate Pit Design	1-11
Figure 1-3: Constancia Site Layout	1-17
Figure 4-1: Constancia Mine Location	4-27
Figure 4-2: Constancia Mine Concession Boundaries	4-30
Figure 5-1: Location of the Constancia Mine	5-37
Figure 7-1: Simplified Geology of the Andahuaylas Yauri Area	7-43
Figure 7-2: Geological Map of the Constancia Deposit	7-43
Figure 7-3: Pampacancha	7-48
Figure 9-1: Near Mine Constancia Exploration Potential	9-51
Figure 10-1: Map of Drill Hole Location	10-55
Figure 12-1: Cu in Duplicate Samples	12-67
Figure 12-2: Cu in Check Samples - ACME vs ALS	12-69
Figure 12-3: Precision Evaluation CuT - Duplicates Samples	12-70
Figure 12-4: Precision Cu (ppm) - Duplicate Samples	12-74
Figure 13-1: Cumulative Distribution of Sgi (Left) and Bond Wi (Right), Constancia	13-80
Figure 13-2: Distribution Of Sag Grindability Index In Pampacancha	13-87
Figure 13-3: Rougher Recovery of Gold and Copper	13-87
Figure 14-1: 3D View of Constancia Domains	14-90
Figure 14-2: Cross Section of Constancia Domains	14-91
Figure 14-3: Pair Analysis Between Blastholes and Drillholes in Supergene Domain	14-92
Figure 14-4: Visual Validation of the Composites vs the Block Grade	14-95
Figure 14-5: Example of Smoothing Subzones	14-96
Figure 14-6: Months Rolling Average Reconciliation Between the Reserve Model and Mill Credited Production of the Mine Compared at same Operational Cut-off	14-97
Figure 14-7: Resource classification (raw & smoothed)	14-99
Figure 14-8: Simplified Geological Framework for the 2017 Long Term Model	14-101
Figure 15-1: Constancia Ultimate Pit Design	14-109
Figure 15-2: Pampacancha Ultimate Pit Design	14-110
Figure 16-1: Contstancia Site Layout	14-112
Figure 16-2: Constancia and Pampacancha Ultimate Pit Design	14-113
Figure 16-3: Constancia Pit Structural Domains	14-114
Figure 16-4: Constancia Pit Slope Sectors	14-115
Figure 16-5: Pampacancha Pit Slope Sectors	14-116
Figure 16-6: Constancia Phase Design, Section View SN	14-117
Figure 16-7: Pampacancha Phase Design, Section View NS	14-117
Figure 16-8: Constancia Pit Phase 2	14-118
Figure 16-9: Constancia Pit Phase 4	14-119

xiv

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
Figure 16-10: Constancia Pit Phase 5	14-120
---	---
Figure 16-11: Constancia Pit Phase 6	14-121
Figure 16-12: Constancia Pit Phase 7	14-122
Figure 16-13: Constancia Pit Phase 8	14-123
Figure 16-14: Constancia Pit Phase 9	14-124
Figure 16-15: Pampacancha Pit Phase 1	14-125
Figure 16-16: Pampacancha Pit Phase 2	14-125
Figure 16-17: Destination Flowsheet	14-129
Figure 16-18: Mine Plan of Periods 2021 and 2022	14-130
Figure 16-19: Mine Plan of Periods 2023 and 2024	14-130
Figure 16-20: Mine Plan of Periods 2025 and 2026	14-130
Figure 16-21: Mine Plan of Periods 2027 to 2030	16-131
Figure 16-22: Mine Plan of Periods 2031 to 2034	16-131
Figure 16-23: Mine Plan of Periods 2035 and 2037	16-131
Figure 16-24: Total Movement by Destination	16-133
Figure 16-25: Total Movement by Phase	16-133
Figure 16-26: Ore Tonnage Milled by Phase	16-133
Figure 16-27: Total Copper Production (2021 - 2037) in Concentrate	16-134
Figure 16-28: Waste Rock Facility	16-137
Figure 16-29: Location of Ore Stockpiles	16-138
Figure 16-30: TMF - Constancia - Final Configuration	16-139
Figure 16-31: Truck Study Results	16-141
Figure 16-32: Truck Speed by Gradient	16-143
Figure 16-33: Predicted Hydraulic Conductivities in Constancia and Pampacancha Pits	16-144
Figure 17-1: 3D View of the Concentrator	17-146
Figure 17-2: Constancia Process Flowsheet	17-148
Figure 17-3: Processing Plan Layout	17-149
Figure 17-4: Primary Crusher Isometric View	17-151
Figure 17-5: Coarse Ore Stockpile Isometric View	17-151
Figure 17-6: Grinding Circuit Isometric View	17-152
Figure 17-7: Cu-Mo Flotation Circuit Isometric View	17-154
Figure 17-8: Cu-Mo Separation Circuit Isometric View	17-156
Figure 17-9: Cu Concentrate Thickening and Filtration Area Isometric View	17-156
Figure 17-10: LOM Copper Concentrate Production	17-157
Figure 17-11: 3d Render of the Pebble Crusher Installation	17-160
Figure 17-12: Rougher mass recovery versus copper recovery (left) and versus rougher concentrate mass flow (right)	17-162
Figure 18-1: Average Power Consumption	18-168
Figure 18-2: Peru - Constancia Site to Port Access Roads	18-169
Figure 18-3: Matarani Port	18-170

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
Figure 18-4: Receiving System, Warehouse C and Ship Loader (Aerial View)	18-170
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xvi

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

LIST OF TABLES

Table 1-1: Drilling Programs by Year (in metres drilled)	1-4
Table 1-2: Copper Production Per Year	1-6
Table 1-3: Constancia Mineral Resource Estimates Exclusive of Mineral Reserves as of January 1st, 2021	1-7
Table 1-4: Comparison Between the january 1, 2020 and january 1, 2021 Mineral Resource Estimates for Constancia	1-8
Table 1-5: Pampacancha Mineral Resource Estimates Exclusive of Mineral Reserves as of January 1, 2021	1-9
Table 1-6: Constancia and Pampacancha Mineral Reserves as at January 1st, 2021	1-9
Table 1-7: Comparison of the Mineral Reserve Estimates at Constancia between january 1^st^, 2020 and january 1^st^, 2021	1-10
Table 1-8: Constancia and Pampacancha Breakdown by Phases as of January 1^st^, 2021	1-12
Table 1-9: Mine Production Schedule Criteria	1-12
Table 1-10 Ore Processing Plan 2021 - 2037 (Total)	1-14
Table 1-11: Life of Mine Capital Expenditures (Sustaining and Growth)	1-20
Table 1-12: Operating Costs	1-20
Table 2-1: Technical Report Participants	2-23
Table 4-1: Hudbay Concessions	4-31
Table 4-2: UEA Pampacancha	4-32
Table 4-3: Kusiorcco Concessions	4-33
Table 4-4: Cabaillito and Maria Reyna Concessions	4-33
Table 4-5: Surface Legal Access Rights Table	4-35
Table 5-1: Approximate Distances and Driving Times to Constancia	5-38
Table 6-1: Production performance	6-41
Table 9-1: Historic Drill Results from the Maria Reyna Property	9-53
Table 10-1: Drilling Programs by Year (In Metres Drilled)	10-54
Table 10-2: Drilling Programs by Type	10-54
Table 11-1: Summary of the Sample Preparation Method by Companies	11-59
Table 11-2: Standard Reference Material	11-63
Table 11-3: Quality Control by Drilling Campaigns	11-64
Table 11-4: Ore Research and Exploration Standard Reference Materials	11-65
Table 11-5: Matrix Match Standard Reference MAterials	11-65
Table 12-1: Comparative Summary of the 2019 and 2020 Drillhole Databases (bias results from truncation during importing process between the Datashed and Minesight softwares)	12-73
Table 13-1: Copper and Molybdenum Recovery	13-75
Table 13-2: Summary of Sag Mill Surveys and Model Calibration	13-76
Table 13-3: Summary of Ball Mill Surveys and Model Calibration	13-77
Table 13-4: Head Grades Of Variability Samples Tested For Rougher Flotation Kinetics	13-81
Table 13-5: Rougher Recovery of Variability Samples	13-82
Table 13-6: Phase 1: Pampacancha Comminution Results	13-83
Table 13-7: Phase 1: Pampacancha Locked Cycle Test Results - Revised Conditions	13-83

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
Table 13-8: Phase 1: Rougher Variability Testing Results	13-84
---	---
Table 13-9: Phase 1: Mineralogical analysis (Mineral Mass %)	13-85
Table 13-10: Pampacancha Locked Cycle Test Conditions	13-85
Table 13-11 Pampacancha Phase 2: Locked Cycle Test Results	13-86
Table 13-12: Phase 2: Rougher Variability Testing Results	13-86
Table 13-13: Cleaner Recovery as a Function Of % Solids, Ph, And P80	13-88
Table 14-1: Drilling Data by Year	14-89
Table 14-2: Oxidation Zone Formulation	14-90
Table 14-3: Variogram Parameters (Spherical Models)	14-94
Table 14-4: Search Ellipse Parameters	14-94
Table 14-5: Summary of the Applied Smoothing Corrections for Copper	14-97
Table 14-6: Constancia Mineral Resource Estimates EXclusive of mineral reserves	14-100
Table 14-7: Cost Inputs for Pit Shell Construction	14-104
Table 14-8: Recovery vs Material Assumed for Pit Shell Construction	14-104
Table 14-9: Pampacancha Mineral Resource Estimates Exclusive of mineral Reserves	14-104
Table 15-1: Operating Cost	14-107
Table 15-2: Metal Recoveries	14-107
Table 15-3: Pit Design Parameters	14-109
Table 15-4: Constancia and Pampacancha Mineral Reserve estimates as at January 1, 2021	14-110
Table 15-5: Comparison of the Mineral Reserve Estimates at Constancia as at December 31, 2019 and as at January 1, 2021	14-111
Table 16-1: Recommended Pit Slope Angles by Structural Domains	14-114
Table 16-2: Constancia, Recommended Pit Slope Angles	14-115
Table 16-3: Pampacancha, Recommended Pit Slope Angles	14-116
Table 16-4: Constancia and Pampacancha Breakdown by Phases as of December 31st, 2020	14-126
Table 16-5: Mine Production Schedule Criteria	14-126
Table 16-6: Mine Plan Criteria	14-127
Table 16-7: Distances between Phases and Facilities	14-129
Table 16-8: Constancia Operations Mine Plan at January 1, 2021	16-132
Table 16-9: Constancia Ore Processing Plan 2021 - 2037	16-134
Table 16-10: Pampacancha Ore Processing Plan 2021 - 2037	16-135
Table 16-11: Ore Processing Plan 2021 - 2037 (Total)	16-136
Table 16-12: Constancia, Waste Material Balance	16-137
Table 16-13: Tailing Management Facility (TMF) Stacking Plan	16-139
Table 16-14: Material Characteristics	16-140
Table 16-15: Mine Equipment 2021 - 2029	16-140
Table 16-16: Mine Equipment 2030 - 2037	16-141
Table 16-17: Parameters for Drilling Production	16-142
Table 16-18: Truck Fixed Time	16-143
Table 17-1: Key Process Design Criteria	17-147

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report
Table 17-2: Operating Parameters	17-159
---	---
Table 17-3: Capital Cost Summary For Pebble Crushing	17-161
Table 17-4: Capital Budget For Recovery Projects	17-163
Table 21-1: LIFE OF MINE CAPITAL EXPENDITURES (SUSTAINING AND GROWTH)	21-179
Table 21-2: On Site Operating Costs	21-180
Table 21-3: Operating Costs - OPEX - Mining/Activity	21-180
Table 21-4: Operating Costs - OPEX - Mining/Components	21-180
Table 21-5: Operating Costs - OPEX - Milling and G&A	21-181
Table 21-6: Capitalized Stripping (Total Estimate and Unit Costs)	21-181
Table 21-7: Operating Costs and Sustaining COsts per Pound	21-181

xix

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

1. SUMMARY

This Technical Report has been prepared for Hudbay Minerals Inc. ("Hudbay") to support the public disclosure of mineral resources and mineral reserves at Constancia mine as of January 1st, 2021. This Technical Report conforms to NI 43-101 Standards of Disclosure for Mineral Projects.

Hudbay is an integrated Canadian mining company with assets in North and South America, principally focused on the discovery, production, and marketing of base metals and precious metals. Hudbay's objective is to maximize shareholder value through efficient operations, organic growth and accretive acquisitions, while maintaining its financial strength.

Hudbay's operations at Constancia include the Constancia and Pampacancha pits, an ore processing plant, a waste rock facility, a tailings management facility and other ancillary facilities that support the operations.

As of the date of this Technical Report, the Constancia mine is in steady state production. The pre-stripping started in March 2014 and the concentrator ramp-up started in December 2014. Commercial production was achieved on April 30, 2015 and the operations continue in alignment with management's expectations in terms of metal production and cost.

The Qualified Person (the "QP") who supervised the preparation of this Technical Report is Olivier Tavchandjian, P. Geo. Mr. Tavchandjian is Vice-President Exploration and Geology of Hudbay.

1.1. PROPERTY DESCRIPTION AND LOCATION

The Constancia mine is located approximately 600 km southeast of Lima in the south-eastern Andes of Peru, in the Chamaca, Livitaca and Velille districts, province of Chumbivilcas, department of Cusco, at approximately longitude 71°47' and latitude 14°27' south.

Hudbay owns an indirect 100% interest in the property, which consists of 36 mining concessions, all of them duly granted by the Peruvian State and recorded in the name of HudBay Peru S.A.C. ("Hudbay Peru"), a wholly-owned subsidiary of Hudbay. Figure 1-1, shows the location of Constancia.

The Constancia mine is accessible from Lima by air, via either Arequipa or Cusco, and then by vehicle over paved and gravel roads. The Constancia mine is located approximately 80 km from Yauri by road. Power for the Constancia mine is supplied from the new 220kV transmission line from Tintaya to Constancia that was built for and is owned by Hudbay but operated and maintained by a third party.

1-1

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

FIGURE 1-1: CONSTANCIA MINE LOCATION

1.2. GEOLOGICAL SETTING AND MINERALIZATION

CONSTANCIA

The Constancia deposit is a porphyry copper-molybdenum system which includes copper-bearing skarn mineralization. Multiple phases of monzonites and monzonite porphyry have intruded a sequence of sandstones, mudstones and micritic limestone of Cretaceous age.

The majority of the mineralization is associated with potassic alteration and quartz veining, occurring as chalcopyrite-(bornite)-molybdenite-pyrite mineralization in "A" and "B" type veinlets, and replacing ferromagnesian minerals or filling fractures. Copper grades are highest where fracture-filling style copper mineralization is superimposed on earlier disseminated copper mineralization. The higher-grade hypogene copper mineralization is hosted by a dense A-veinlet stockwork developed in an early porphyry phase. The pyrite/chalcopyrite ratio is typically below 2:1. Molybdenite commonly increases with depth in association to "B" veinlets. Bornite occurs sporadically especially at deeper levels, sometimes associated with some gold values.

Propylitic alteration is transitional to the potassic alteration and extends more than one kilometre from the porphyry intrusive contacts. The propylitic alteration mineral assemblage includes epidote-chlorite-calcite- pyrite-rhodochrosite. Subordinate chalcopyrite is also present, filling fractures or replacing mafic minerals. Sphalerite-galena veinlets and veins are distributed as a halo to the copper-molybdenum mineralization within the propylitic alteration zone up to 3 km away from the porphyry copper system.

Phyllic alteration forms a pervasive carapace surrounding and sometimes overprinting potassic alteration. The phyllic alteration accompanies almost complete destruction of primary rock textures; the mineral assemblage includes sericite-quartz-pyrite, limited amounts of chalcopyrite and associated occasional "D" veins and veinlets.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

At the contact between intrusions and limestones, a magnetite garnet skarn develops, while a pyroxene-diopside (garnet-epidote) association is more common in calcareous sandstones and arkoses of the Chilloroya formation. Skarn mineralization is volumetrically much smaller, but grades are normally higher.

Structural deformation has played a significant role in concentrating the hydrothermal alteration and the copper-molybdenum-silver-gold mineralization, including skarn formation. Major inter and post mineral fracture systems in the deposit area strike northeast and include the Barite fault system. This is represented by a number of nearly parallel vein-faults carrying base metal sulphides and barite which have been exploited by artisanal workings throughout the property. A second important system strikes north-south. It appears to be more recent than the Barite system and controls part of mineralization and most of the silicified breccias (sometimes mineralized) in the system.

PAMPACANCHA

The Pampacancha deposit is a porphyry Cu-Mo-Au related Skarn system. Oligocene unmineralized basement diorite is intruded by the diorite porphyry cited as the source for skarn mineralization. This in turn is cut by intra-mineral monzonite intrusions which provide minor local increases in Cu-Au and also locally replaces skarn Cu-Au mineralization which is most developed at the upper and lower margins of the limestone body. Magnetite-chalcopyrite-pyrite skarn ranges to marginal less well mineralized garnet and pyroxene skarn, locally overprinted by epidote-bearing retrograde skarn.

Epithermal mineralization as low sulphidation quartz-sulphide Au + Cu style accounts for common supergene enriched Au anomalies along with other features such as hydrothermal alteration and veins typical of near porphyry locations.

1.3. EXPLORATION

Copper and gold exploration in the Constancia area dates back to the early 1990s. The San José prospect (now part of the Constancia Mine) was explored by Mitsui during the 1980s with a focus on delineation of high-grade ore amenable to process at the Katanga facility. Exploration consisted of detailed mapping, soil sampling, rock chip sampling, ground magnetic and IP surveys with several drilling campaigns, mainly located in the western and southern sides of the prospect.Rio Tinto expanded the exploration coverage in the Constancia area between 2003 and 2004.

Exploration activities between 2007 and 2011 were carried out by Norsemont Exploration Inc. and included geological mapping rock and stream sediment sampling and ground magnetic and IP surveys. From 2005 to 2010, Norsemont drilled 153,556m for infill, condemnation, metallurgical, geotechnical, hydrogeological and exploration, drilled by core and reverse circulation.

Following its acquisition of the Constancia Project in March 2011, Hudbay's exploration activities focused on delineating the "new" Pampacancha deposit and understanding the district exploration upside. Since acquisition, Hudbay has completed over 45,000 m of drilling for resource definition, condemnation, metallurgical, geotechnical, hydrogeological and exploration drilling. These holes were drilled by core and reverse circulation.

From 2019 to 2020, additional exploration activities were carried out to the northwest of Constancia, between the northwest limit of the pit and the area known as Yanajaja. A succesful drill program was conducted over this area now referred to as Constancia Norte and the results have been incorporated in the updated mineral resource and mineral reserve estimates presented in this Technical Report.

In December 2017, Hudbay completed the acquisition of a large, contiguous block of mineral rights including the Caballito, Maria Reyna and Kusiorcco concessions to explore for mineable deposits within trucking distance of the Constancia processing facility. Hudbay has made significant progress towards permitting, community relations and technical activities required to access and conduct drilling activities on these properties. Drilling started early in 2021 on the Quehuincha North target, a skarn prospect in a similar geological environment than Pampacancha and located approximately 10 km north of the Constancia mill. Exploration agreements with local communities are required before Hudbay can access and carry out exploration on the other satellite properties.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

1.4. DRILLING

Extensive drilling has been conducted at the Constancia mine and the Pampacancha deposit by several successive property owners. The most recent drilling was done by Hudbay, with prior drilling campaigns completed by Rio Tinto and Norsemont.

Table 1-1 summarizes the drill holes used to estimate the current mineral resource estimate, with regional exploration holes excluded.

TABLE 1-1: DRILLING PROGRAMS BY YEAR (IN METRES DRILLED)

1.5. SAMPLE PREPARATION, ANALYSES, AND DATA VALIDATION

Rio Tinto conducted its own internal Quality Control / Quality Assurance (QA/QC) program to independently evaluate the quality of the assays reported by ALS. Standards and blanks were systematically inserted in the sample stream.The core from the Norsemont drilling programs from 2005 to 2010 were transported to SGS and ALS Laboratories in Lima, Peru for preparation and analysis. Drill core samples across the Constancia deposit were also measured for specific gravity at ALS Chemex, Lima, Peru. Norsemont conducted its own internal QA/QC program to independently evaluate the quality of the assays reported by SGS and ALS. Duplicates, standards and blanks were systematically inserted in the sample stream.

During the Hudbay 2011 campaign, the samples were transported to the SGS laboratory in Lima, Peru for preparation and analysis while in 2014, the samples were transported to the SGS Constancia laboratory at the mine site for preparation. Once samples were pulverized, a 250 g subsample pulp was collected and air freighted to the SGS laboratory in Lima, Peru, for analysis. In 2015, the samples were transported to the SGS Constancia laboratory at the mine site for preparation and analysis and in 2017, for the twin hole drilling program, blanks and standards were inserted at site, prior to dispatching the core boxes to the Certimin laboratory in Lima for sample preparation followed by analysis at the Certimin and SGS laboratories in Lima. Since 2018, operation of the Constancia laboratory has been transferred to Bureau Veritas where all samples collected since have been sent for sample preparation. The pulps were then sent to the Bureau Veritas laboratory in Lima for ICP analysis. The SGS laboratory in Lima has been retained for external checks.

As part of Hudbay's QA/QC programs, samples were systematically introduced in the sample stream to assess adequate sub-sampling procedures, potential cross-contamination, precision, and accuracy.

A total of 2,166 samples from 292 drill holes were collected for specific gravity determinations and sent to ALS Chemex and SGS labs in Lima, Peru. Density measurements were done using immersed wax coated core. The drill core samples, 8 to 10 cm long, were taken every 50 m from half-split core.

1-4

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

An internal validation of the drillhole database against the original drill logs and assay certificate information was carried out by Hudbay in 2020. During this exercise, approximately 5% of the total database was entirely validated with the exception of 4 samples representing 0.09% of the total and therefore the database can be considered very reliable. A comparison with previous versions of the resource modelling database used between 2014 and 2019 highlighted that precision had been truncated to the second decimal place resulting in an under-estimation of the gold grade and with no significant impact on the other metals of economic significance. The correction of this truncation in 2020 has resulted in the elimination of a conservative bias in gold content of approximately 10% previously experienced and that had not been detected since at least 2014.

1.6. MINERAL PROCESSING AND METALLURGICAL TESTING

The ore types that are currently being processed in the concentrator were established during the Definitive Feasibility Study (DFS) as the hypogene, supergene, skarn, mixed and high zinc types. The metallurgical responses of these ores are acceptable in terms of treatment rate, recovery and molybdenum and copper concentrate grades. Metallurgical test work has been and continues to be performed at laboratory and plant levels to optimize the reagents, operating conditions, and flowsheet configuration for each ore type.

The process plant tonnage is based primarily on current and historical plant performance. Geometallurgical test work indicates that the Constancia ore hardness is increasing, but the softer Pampacancha ores allow for the deferral of the pebble crushing circuit, now scheduled for installation in 2023 and operation as of 2024. Basic engineering has been completed for the project, and the capital cost is estimated at $22 million.

Metal recovery from Constancia ores is estimated based on plant performance and geometallurgical testing. Current plant recovery is lower than design because of the coarser nominal grind size (150 µm) attributable to the increase in plant treatment capacity from 76,000 TPD to 90,000 TPD. A comprehensive recovery optimization program is scheduled for implementation in 2023, aimed at debottlenecking the regrind and cleaning circuit and increasing the recovery to design levels. The capital cost estimated for this project is $7.85 million.

The Pampacancha ore recovery and throughput assumptions when processed at the Constancia plant are based on prefeasibility level engineering and were updated for this technical report based on a feasibility level laboratory test program Table 1-2 shows the estimated copper and molybdenum recoveries estimated for the years of 2021 to 2037. Pampacancha recoveries are constant at 85%.

1-5

Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

TABLE 1-2: COPPER PRODUCTION PER YEAR

1.7. MINERAL RESOURCE ESTIMATES

CONSTANCIA

The Constancia mineral resource estimate is effective as of January 1st, 2021.

The initial mineral resource and mineral reserve estimate for Constancia was completed by AMEC in 2012, on behalf of Hudbay, and published by Hudbay in a technical report dated October 15, 2012 and titled "National Instrument 43-101 Technical Report, Constancia Project, Province of Chumbivilcas, Department of Cusco, Peru".

Additional work performed by Hudbay between 2012 and 2017 led to mineral resource and mineral reserve estimate updates supporting a technical report filed on the property by Hudbay in March 2018 (the "2018 Technical Report").

After carrying out additional work between 2018 and 2020, Hudbay has completed a new mineral resource and mineral reserve estimate for Constancia, with an effective date of January 1st, 2021 which constitutes the basis for the updated mineral resource estimate presented in this Technical Report. A total of 143,917 m in 492 holes have been used in the construction of the resource model for the Constancia deposit.

Resource modeling at Constancia is based on integrated geological and assay interpretations of information recorded from diamond core logging and assaying and is comprised of the following key steps: Construction of mineralized envelopes, Exploratory Data Analysis, Modelling (Composites, Variography and Interpolation) and Validations including a thorough assessment of the smoothing effect occurring during grade interpolation by domain with consistent drill spacing and statiscal distritbutions and correction for over-smoothing when required. This methodology is validated through reconciliation between mineral reserve estimates and mill credited production.

The Constancia geological model was developed from an initial interpretation of six lithology domains, six alteration types and five oxidation zones. The oxidation model is based on sequential copper formulation and logging codes. This geological framework was then used to model two continuous estimation domains hosting the mineralization grading above 0.1% Cu respectively in the supergene, hypogene and Northern skarn domains. Grade estimation was also conducted separately in post mineralization barren dykes cross-cutting the mineralization as well in the host rocks and in the oxidised/leach cap domain. Mineral resource and mineral reserve estimates are only reported from the supergene, hypogene and skarn domains.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

For the three domains of hypogene, supergene and skarn mineralization, resource classification is based on the kriging slope of regression which is a function of drill spacing, the continuity of the mineralization and mining block geometry. This parameter has been found to provide a reliable resource classification criteria by Hudbay through quarterly and annual reconciliation between reserve estimates and mill credited production at its operating mines. The kriged regression slope for measured resources at Constancia is >80% and is between 60 and 80% for indicated resources. Some local adjustments were made to produce resource category domains that are smoother and more continuous.

The mineral resource estimates exclusive of mineral reserve estimates, which is consistent with the presentation in Hudbay's Annual Information Form, are reported in Table 1-3 at a US$6.14/tonne Net Smelter Return (NSR) cut-off and inside an economic pit shell demonstrating that they have reasonable prospects for economic extraction under the same economic assumptions as used for reserve reporting.

TABLE 1-3: CONSTANCIA MINERAL RESOURCE ESTIMATES EXCLUSIVE OF MINERAL RESERVES AS OF JANUARY 1ST, 2021

Notes:

Totals may not add up correctly due to rounding.
Mineral resources are exclusive of mineral reserves and do not have demonstrated economic viability.
Mineral resource estimates do not include factors for mining recovery or dilution.
Metal prices of $3.10 per pound copper, $11.00 per pound molybdenum, $1,500 per ounce gold, and $18.00 per ounce silver were used to estimate mineral resources.
Constancia mineral resources are estimated using a minimum NSR cut-off of $6.14 per tonne and assuming metallurgical recoveries (applied by ore type) of 85.8% on average for the life of mine.
Mineral resources are based on resource pit designs containing measured, indicated, and inferred mineral resources.

The comparison between the January 1^st^, 2020 and January 1^st^, 2021 mineral resource estimates is presented in Table 1-4. The measured, indicated and inferred resource tonnages exclusive of mineral reserve estimates have been reduced in the 2020 estimates by 33Mt and 36Mt respectively due to resource to reserve conversion and the removal of low grade inferred resources in host rocks. The overall metal contained in mineral resource estimates has been maintained as the reduction in tonnage has been offset by an improvement in average grade for all metals for the most part due to the success of the drilling program at Constancia Norte which has resulted in the addition of high grade skarn mineralization not yet converted to reserves either due to lack of drilling or high strip ratio.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

TABLE 1-4: COMPARISON BETWEEN THE JANUARY 1, 2020 AND JANUARY 1, 2021 MINERAL RESOURCE ESTIMATES FOR CONSTANCIA

Notes:

Totals may not add up correctly due to rounding.
Re-evaluation of economic viability.
Mineral resources are exclusive of mineral reserves and do not have demonstrated economic viability.
Mineral resource estimates do not include factors for mining recovery or dilution.
Metal prices of $3.10 per pound copper, $11.00 per pound molybdenum, $1,500 per ounce gold, and $18.00 per ounce silver were used to estimate mineral resources.
Constancia mineral resources are estimated using a minimum NSR cut-off of $6.14 per tonne and assuming metallurgical recoveries (applied by ore type) of 85.8% on average for the life of mine.
Mineral resources are based on resource pit designs containing measured, indicated, and inferred mineral resources.

PAMPACANCHA

Resource estimation at Pampacancha is based on a total of 140 holes (38,240 metres) with 28 of those being derived from reverse circulation drilling and the remaining 112 from HQ diameter diamond drilling. All holes were drilled from surface by Geotec. Core recovery is near 100% for all holes. There have been no change to the mineral resource model for Pampacancha since 2017.

The skarn unit hosts the vast majority of the copper-gold-silver-molybdenum mineralization.

The Pampacancha geological model was developed from an initial interpretation of a high grade and a low grade skarn domains and using a smooth grade envelope defined at a 0.1% Cu cut-off to limit grade smearing into the host rocks. No hard boundary was used for grade interpolation purposes between the low grade and high grade skarn sub-domains.

There has been no change to the mineral resource estimates for Pampacancha as of January 1, 2021. Pampacancha mineral resource estimates are reported at a US$6.14/tonne Net Smelter Return (NSR) and inside a pit shell demonstrating that they have reasonable prospects for economic extraction.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

TABLE 1-5: PAMPACANCHA MINERAL RESOURCE ESTIMATES EXCLUSIVE OF MINERAL RESERVES AS OF JANUARY 1, 2021

Notes:

Totals may not add up correctly due to rounding.
Mineral resources are exclusive of mineral reserves and do not have demonstrated economic viability.
Mineral resource estimates do not include factors for mining recovery or dilution.
Metal prices of $3.10 per pound copper, $11.00 per pound molybdenum, $1,500 per ounce gold, and $18.00 per ounce silver were used to estimate mineral resources.
Mineral resources are estimated using a minimum NSR cut-off of $6.14 per tonne and assuming metallurgical recoveries (applied by ore type) of 85.8% on average for the life of mine.
Mineral resources are based on resource pit designs containing measured, indicated, and inferred mineral resources.

1.8. MINERAL RESERVE ESTIMATES

Mineral Reserves estimates for the Constancia and Pampacancha deposits, which are presented in this Technical Report, were prepared by Hudbay. The Qualified person who supervised the preparation of this Technical Report is Olivier Tavchandjian, P.Geo. Mr. Tavchandjian is Hudbay's Vice-President Exploration and Geology of Hudbay.

Hudbay has opted to use an NSR optimization model taking into account the Cu, Mo, Ag, and Au grades, mill recoveries, contained metal in concentrate, deductions and payable metal values, metal prices, freight costs, smelting and refining charges and royalty charges.

The 2020 Proven and Probable Reserve estimates at the Constancia mine total 532.5 million tonnes at a copper grade of 0.31% that supports a 17 year mine life. The mine plan is based on the capacity of the process plant, which in turn relies on the grinding circuit throughput. The plant has the capacity to process 31 Mtpy (90ktpd at 94% availability)

TABLE 1-6: CONSTANCIA AND PAMPACANCHA MINERAL RESERVES AS AT JANUARY 1ST, 2021

Notes:

Totals may not add up correctly due to rounding.
Metal prices of $3.10 per pound copper, $11.00 per pound molybdenum, $1,500 per ounce gold, and $18.00 per ounce silver were used to estimate mineral reserves.
Mineral reserves are estimated using a minimum NSR cut-off of $6.14 per tonne and assuming metallurgical recoveries (applied by ore type) of 85.8% on average for the life of mine.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

A comparison of the December 31^st^, 2019 and December 31^st^, 2020 Mineral Reserve estimate is summarised in Table 1-7 for the Constancia pit. The January 1^st^, 2021 Constancia Reserve estimate shows an increase of 12% after substracting the mining depletion in the copper contained in the Mineral Reserves estimates at Constancia. The reserve addition at Constancia results from a successful drilling campaign at Constancia Norte and a revision and improvement of the mine design and mine plan.

TABLE 1-7: COMPARISON OF THE MINERAL RESERVE ESTIMATES AT CONSTANCIA BETWEEN JANUARY 1^ST^, 2020 AND JANUARY 1^ST^, 2021

Notes:

Totals may not add up correctly due to rounding.
Re-evaluation of economic viability.
Mineral resources are exclusive of mineral reserves and do not have demonstrated economic viability.
Metal prices of $3.10 per pound copper, $11.00 per pound molybdenum, $1,500 per ounce gold, and $18.00 per ounce silver were used to estimate mineral reserves and resources.
Constancia mineral reserves and resources are estimated using a minimum NSR cut-off of US$6.14 per tonne.
Metallurgical recoveries are applied by ore type and assumed to be 85.8% on average for the life of mine.

The updated mineral reserve estimates are supported by reserve to mill credit reconciliation work conducted by Hudbay since 2017.

1.9. MINING METHODS

The Constancia mine is a traditional open pit operation using conventional truck and shovel mining. The Constancia ultimate pit design will measure approximately 1.6 km east to west, 1.7 km north to south, and have a maximum depth of around 705 m. The Pampacancha ultimate pit design will measure approximately 0.6 km east to west, 1 km north to south, and have a maximum depth of about 300 m. A primary waste rock facility (WRF), which is located to the south and east of the Constancia pit, is intended to be used for both deposits.

The processing facility is located approximately 1 km west of the Constancia Pit. A non-acid generating (NAG) stockpile for waste material is located at the south side while the tailings management facility (TMF) is located 3.5 km to the southwest of the Constancia pit.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

MINE PHASES

Final pit limit designs have been created for Constancia and Pampacancha based on the selected pit shells from the pit optimization. Nine pit stages were developed for Constancia and two pit stages for Pampacancha Pit as described in Table 1-8 below. The minimum mining width for each phase is 60 metres which allows for operation of a shovel, trucks (in two lines) and a drill. The phased development strategy consists of extracting the highest metal grades along with minimum strip ratios during the initial years to maximize the economic benefits of the ore-body, while enabling smooth transitions in waste stripping throughout the life of the mine to ensure enough ore exposure for mill feed. The following Figure 1-2 shows the footprint of the ultimate pits for Constancia and Pampacancha at the end of the mine life.

Total ore reserves included in the final two pits is estimated to be 532.5 million tonnes and total waste material is estimated to be 569.4 million tonnes. Approximately 6.8 million tonnes of medium and low grade ore had been hauled to the temporary stockpiles as of the end of 2020. This material will be sent to and reclaimed from the stockpiles through the life of mine. At the end of the mine's life, the ore from these stockpiles will be reclaimed and processed.

FIGURE 1-2: CONSTANCIA AND PAMPACANCHA ULTIMATE PIT DESIGN

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

TABLE 1-8: CONSTANCIA AND PAMPACANCHA BREAKDOWN BY PHASES AS OF JANUARY 1^ST^, 2021

MINE PRODUCTION SCHEDULE

The operating and scheduling criteria used to develop the mining sequence plans are summarized in Table 1-9 below.

TABLE 1-9: MINE PRODUCTION SCHEDULE CRITERIA

Pit and mine maintenance operations are being scheduled around the clock. Allowances for down time and weather delays have been included in the mine equipment and manpower estimations.

A variable cut-off grade strategy is implemented to bring forward the higher grade ore from the pit to the early part of the ore production schedule. Delivering higher grade ore to the mill in the early years will improve the net present value and internal rate of return of the project. Priority plant feed consists of higher grade material (NSR > 10$/t). The lower grade material is processed as needed or sent to long term ore stockpiles to be reclaimed at the end of mine life.

The mine production plan includes 569.4 Mt of waste and 532.5 Mt of ore (from both pits and stockpiles), yielding an average stripping ratio (waste/ore) of 1.1. An average yearly mining rate of 77.0 Mtpy, through the first 13 years, with a maximum of 81 Mtpy, is required to provide a nominal ore process feed rate of 31.3 Mtpy based on a variable throughput by ore type (90-94 ktpd and 94% availability). The ore production schedule for the operation presents step down trends with the Cu grade averaging 0.44% Cu from 2022 to 2024 when Pampacancha is at its peak production, then reducing to 0.34 % Cu from 2025 to 2029 with the preferential scheduling of higher grade mineralization from the Contancia pit and finally decreasing to 0.25% Cu to the end of the mine life. The LOM average grade is 0.311% Cu, 0.009% Mo, 0.065 g/t Au and 3.04 g/t Ag for the remaining 17 years of the mine life. The estimated plant feed schedule is summarized in Table 1-10.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

WASTE ROCK FACILITY (WRF)

The final design geometry for the WRF incorporates the same slope profile as the original WRF: 20 m high benches with 1.4H:1V (36°) of bench slopes and 32 m wide benches. The overall slope of the stockpile will be about 3.0H:1V (18°). The current remaining capacity of the WRF is estimated at 413 Mt. and as such there is sufficient capacity to accommodate the waste rock needs for the revised LOM plan.

ORE STOCKPILES

The Constancia mine is planned to have four operational stockpiles, which will segregate the ROM material by ore type and grade range. The lift height of these stockpiles ranges between 12 and 20 meters. The planned maximum stockpiles capacities are:

Stockpile 01 : 600 Kt
Stockpile 02 : 500 Kt
Stockpile 03 : 600 Kt
Stockpile 04 : 5,000 Kt (stock located inside WRF, lift 4240, and used until 2025)
Stockpile 05 : 2,000 Kt (a finger from stock 04 destinated for Pampacancha ore)
Stockpile 06: 70,000 Kt (low-grade stock located at south from WRF)

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

TABLE 1-10 ORE PROCESSING PLAN 2021 - 2037 (TOTAL)

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

MINE EQUIPMENT

There are several different rock types at the Constancia operation but for production estimation purposes, the weighted average of all rock types was used. Major loading and haulage equipment is equipped with electronic monitors which will ensure maximum loading. All production is reported in dry metric tonnes which is consistent with the reserve model. A moisture content ranging between 3.5 and 4.5 percent was measured, and 4.0 percent was used for haulage calculations.

Mine equipment requirements were developed based on the annual tonnage movements projected by the mine production schedule with bench heights of 15 metres, operating two twelve-hour shifts per day, 365 days per year. Productivity is based on current performance and material characteristics specific to the deposit. The fleet requirements also considered:

production drilling;
loading and hauling of sulfide ore to the primary crusher (located on the east side of the pit), and to the WRF and TMF areas;
maintenance of mine haulage and access roads; and
maintenance of the WRF areas, ore stockpiles, TMF and berms, and regrading of slopes and final surfaces.

MINE ENGINEERING

The pit slope angles used for pit optimization of the Constancia and Pampacancha pits are based on engineering studies conducted at pre-feasibility or higher levels of engineering. The pit slope design was developed by Golder, in December 2019 for Constancia and by TWP/Itasca in August 2013 for Pampacancha.

The predicted total dewatering rates for the Pit Expansion Design and the area of influence will be similar to the 180 L/s level approved in the Groundwater Use Licence. The design was partially driven by requirements for mine water supply but the main goal of the dewatering is to depressurize the pit wall to assure the overall pit slope stability.

1.10. RECOVERY METHODS

The process plant was designed, built and commissioned by Ausenco. The recent concentrator production rate is approximately 90,000 tpd with a 94% physical availability. The plant consists of the following areas: crushing, stockpiling, milling, bulk rougher flotation, regrinding, bulk cleaner flotation, concentrate thickening, molybdenum flotation, copper concentrate filtration, and tailings thickening. The final products are copper and molybdenum concentrates, which are shipped by road from site to the port of Matarani for shipment to customers.

Power supply is provided by the south national energy system in the highland sector (New Tintaya sub-station). Water for the process plant is provided by recirculation from the tailings thickeners, TMF, and pit dewatering wells.

For plant automation and process control, a distributed control system (DCS) is used for starting and stopping equipment and control of level, lubrication systems, flows, temperatures, and equipment protection, among others

1.11. PROJECT INFRASTRUCTURE

The infrastructure facilities that support the current operation including the WRF, TMF, water management, electric power supply and transmission, roads and the port.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

The TMF is located on the south side of the Chilloroya River and has been designed to store 600 million tonnes (Mt) of tailings with a final design elevation of 4,190 masl. The TMF has two stages. The first designed by KP (engineering detail) to 4,160 masl and the second stage design by Golder (prefeasibility study) that rises from 4,160 to 4,190 masl without moving the previous footprint. The actual elevation of embankment is 4,097 masl, as approved by the authority for operations until 2019 with with two metres of freeboard. The design of the TMF includes a LLDPE geomembrane liner and underdrains for valleys. According to the LOM plan of the TMF, and as per water balance, construction periods will occur in 2021, 2023, 2025, 2027, 2029, 2031, 2033 and 2035. Optimization of the LOM to ensure sufficient quality rock from the pit for the TMF construction has been completed, such that the mining fleet is utilized for hauling, placement and compaction of structural fill over the downstream and upstream zone of the embankment.

Tailings have been deposited from designated off-take points from a distribution pipeline located along the upstream crest of the embankment at an elevation of 4,097m (spigotting system) and around the perimeter of the facility (discharge diffuser).

The Potentially Acid Generating and Non-Acid Generating Waste Rock Facilities (PAG & NAG WRF) are located in the Cunahuiri Valley east of the Constancia Pit and provide storage for 581 Mt. The facilities receive mine waste material from the operation of the Constancia pit and from the Pampacancha pit. Underdrain systems include groundwater underdrains to collect the water to WRF Retention pond.

The WRF Retention Pond is located downstream from the WRF in the Cunahuiri Valley and provides energy dissipation for surface water. The WRF Containment Pond is located downstream of the WRF Retention Pond to provide storage capacity for surface runoff, direct precipitation and seepage collected from the WRF and its contact diversion channels. Water management structures also include diversion channels, sediment ponds, water management ponds and main sediment pond. Hudbay Peru performed a site-wide process water balance for the final design phase of the Constancia mine based on the updated Mine Plan.

There are two camps on-site: Constancia is a permanent camp with a 2,930 people capacity and Fortunia is a pioneer camp and served as an overflow camp during the construction stages for 627 people. These camps have been designed with reference to the IFC standards for camp construction for mining activities.

Power supply for the Constancia mine comes from the new 220kV transmission line from Tintaya to Constancia that was built for and owned by Hudbay, but is operated and maintained by a third party. This transmission line is connected to the Constancia electric substation that transforms 220kV to 23 kV. The electric control room manages the 23kV system that provides energy to the entire mine. In case of emergency, there are three generators that provide energy in a semi-automated way.

The primary road to site, consists of a 65 km road (National Route PE-3SG) from Coporaque to the Puente Bailey/Constancia. These roads (and bridges) have been upgraded as necessary, to meet the needs of construction and life of mine use. In order to ensure good road conditions, Hudbay performs maintenance as necessary due to heavy truck traffic and the transportation of concentrate and consumables.

Copper concentrate is transported from the Constancia mine to the Matarani port by trucks (485km). These trucks are equipped with a hydraulically operated covered-box hinged at the rear, the front of which can be lifted to allow the concentrate to be deposited in the concentrate shed assigned to Hudbay by TISUR, the port operator. Pier C has been assigned to Hudbay and has a 75 Kt capacity. A chute from the shed feeds a tubular conveyor into a ship loader system with a nominal capacity of 1,500 tons per hour. The same conveyor and ship loading equipment is shared with other copper concentrate exporters. Hudbay is the primary customer for Pier C that was designed to take concentrates by truck and railroad.

Figure 1-3 provides an overall site layout showing the location of the various storage facilities relative to the Constancia and Pampacancha pits and the mill.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

FIGURE 1-3: CONSTANCIA SITE LAYOUT

1.12. MARKET STUDIES AND CONTRACTS

Constancia copper concentrate is a clean, medium grade concentrate containing small gold and silver by-product credits. It is a desirable feedstock for copper smelters in China which is the most geographically appropriate freight destination but is also suitable for processing by smelters in Europe, India and South America.

Constancia copper concentrate is sold directly to a variety of copper smelters in Asia, Europe and India as well as internationally recognized trading companies. Between 85% and 90% of sales are made pursuant to longer-term frame contracts which typically reference annual benchmark agreements between major concentrate producers and smelters for the purposes of fixing key terms such as treatment and refining charges. The balance of projected annual concentrate production has not been committed for sale in order to provide flexibility in the event of potential fluctuations in annual production and will be sold into the spot market each year at then-current market terms.

Production at Constancia is subject to a precious metals streaming agreement with Wheaton Precious Metals (Previously "Silver Wheaton (Caymans) Ltd.") consisting of 50% of payable gold and 100% of payable silver. Hudbay will receive cash payments equal to the lessor of the market price and US$400 per ounce for gold and $5.90 per ounce of silver, subject to a 1% annual escalation starting three years after the completion date in 2016. Under the terms of the stream, gold recovery for the purposes of calculating payable gold will be fixed at 55% for gold mined from Constancia and 70% for gold mined from Pampacancha.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

1.13. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

The Constancia Mine Environmental and Social Impact Assessment (ESIA) was approved by the Ministry of Energy and Mines (MINEM) in November 2010 and the first amendment to the ESIA (MOD I) was approved in August 2013 to increase the processing capacity and to match the Detailed Design Feasibility Study.

On April 2015, the second amendment to the ESIA (MOD II) was approved to increase reserves through the expansion of the Constancia Pit and inclusion of the Pampacancha deposit, and to expand the WRF and TMF, among other things.

In December 2020, the National Environmental Certification for Sustainable Investments Service (SENACE) began with the evaluation of the third amendment to the EISA (EISA MOD III), which includes the mining plan optimization of Constancia and Pampacancha pits, changes in the TMF, NAG waste deposit expansion, the addition of ore piles 3 and 4, as well as other updates to ensure the feasibility of these modifications.

Hudbay has secured all necessary permits and authorizations on time to start construction activities for Pampacancha and operate the mine. A summary of these permits is detailed in Section 20.

Constancia's closure plan was updated in 2015 to incorporate recent studies and technological changes that will reduce costs and financial guarantees that Hudbay must provide annually to the Peruvian government. The current closure plan was approved by the MINEM in June of 2016. This document provides updated measures of closure and post-closure of the mine, according to MOD I and MOD II.

The environmental and social impacts have been assessed and appropriate mitigation measures have been implemented. The ESIA and amendments comply with national regulation and have adopted the Equator Principles and International Finance Corporation's (IFC) Performance Standards. The Towards Sustainable Mining (TSM) standards of the Mining Association of Canada (MAC) have also been implemented.

Hudbay has engaged an Independent Project Review Board (IPRB) to conduct periodic reviews of the major earth structures on the mine with regular visits to site, as well as for monitoring the progress of construction of the TMF. This activity was initiated during the construction stage and now continues during operation. Design reviews and site visits have been carried out at critical junctures throughout the course of construction and mine operations to observe the performance of the structure and development of subsequent raises to the dam and the impoundment.

The hydrogeological report (Golder 2013), which led to a more refined hydrogeological model and plant water balance, provided clarity to three principal areas: (1) probability of hydrological connectivity between the open pit dewatering influence and neighboring distal reservoirs, (2) determined the impact of slope stability on the pit wall design; and (3) confirmed the availability of groundwater supply through the life of the mine.

The Hudbay water management strategy includes the following aspects:

Zero water discharge to the environment from the plant, and PAG waste rock facility.
Discharge from the TMF supernatant during the rainy season.
Prioritization of water sources for the plant.
Water available for compensation purpose for base flow of Chilloroya users and downstream community users.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

The Constancia environmental monitoring program is summarized in the March 2016 Environmental Monitoring Plan (PLA-AMB-04: Plan de Monitoreo Ambiental). This plan covers monitoring of air, noise and vibration, surface water quality and flow, surface lake/reservoir levels, groundwater levels and quality, domestic water consumption and quality, effluent monitoring, topsoil, soil and sediment, biological (flora and fauna) monitoring and hydrobiological and meteorological data collection. During the operation phase, Hudbay is maintaining a joint environmental monitoring committee with the communities that are directly and indirectly influenced by the mine.

From a biodiversity point of view, Hudbay is working on specific management plans for wetlands, flora and fauna. Hudbay has developed a Biodiversity Action Plan (BAP), as generally anticipated and described in the ESIA and as required in the MAC TSM protocol and IFC performance standard N° 6.

Regarding social aspects, between February and April 2012 Hudbay reached agreements with the neighbouring communities of Uchucarco and Chilloroya for land required for the Constancia mine. These agreements were validated with support from two thirds of the community members. Both of these agreements have been recorded in the public land registry. A compensation plan was also prepared for the 36 landholding families from the lchuni area of Chílloroya. As a result of the Resettlement Action Plan (RAP) implementation, prepared in compliance with national laws and international provisions and standards in resettlement matters and in particular the IFC Performance Standards, the resettlement process was successfully completed in 2016.

In February 2020, the community of Chilloroya formally approved a surface rights agreement with Hudbay for the Pampacancha satellite deposit located near Constancia. Throughout the remainder of the year, Hudbay focused on negotiating individual agreements with those members of the Chilloroya community who made use of the Pampacancha lands. As of the date of this Technical Report, there is one remaining land-user family who needs to complete an agreement.

Specific social programs were designed for the mitigation and prevention of the identified impacts of the resetlement program, considering infrastructure, social, economic and strategic development and claims and dispute resolution.

Since 2015 Hudbay has obtained the ESIA MOD II and the mine closure plan, which are the required environmental certifications to support mining activities in Pampacancha.

The Ministry of Energy and Mines authorized the start of the exploitation activities for the Pampacancha pit in December 2020 (Directorial Resolution No. 790-2020-MINEM-DGM) This permit included the Prior Consultation process with the community of Chilloroya.

1.14. CAPITAL AND OPERATING COST

The LOM sustaining capital is estimated to be $898 million (excluding capitalized stripping) for Constancia and Pampacancha while the project capital is estimated to be $51 million. Capitalized stripping represents $350 million over the life of the mine (Table 1-11).

The total capital cost includes expenses required for major mining equipment acquisition, rebuilds, and major repair. The cost also includes site infrastructure expansion (tailings management facility, waste rock facility and others) and process plant infrastructure however they exclude all the cost related to mine closure. Project capex associated includes Pampacancha project development expenditures but excludes additional costs related to the remaining individual land user agreements.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

TABLE 1-11: LIFE OF MINE CAPITAL EXPENDITURES (SUSTAINING AND GROWTH)

The operating costs at Constancia are developed annually as part of the site budget process. The operating costs are divided in three centers of importance: (i) mining, (ii) milling and (iii) G&A. The LOM operating costs are shown in Table 1-12.

TABLE 1-12: OPERATING COSTS

1.15. ECONOMIC ANALYSIS

Pursuant to NI 43-101, producing issuers may exclude the information required for Section 22 Economic Analysis on properties in production, unless the technical report includes a material expansion of current production. As Hudbay is a producing issuer, it has excluded information required by Item 22 of Form 43-101F1 as the Pampacancha expansion does not represent a material expansion of the current production facilities at Constancia.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

1.16. CONCLUSIONS

The Constancia mine has been in continuous operation since declaring commercial production on April 30, 2015. The mine has continually improved its performance indicators to a level supporting the projected throughput and forecasted results as outlined in this report with a processing plant capacity of 90,000 tpd producing both copper and molybdenum saleable concentrates. The resource and reserves stated for the Constancia and Pampacancha deposits are compliant to industry best practices as outlined in the Canadian Institute of Mining, Metallurgy and Petroleum Definition Standards for Mineral Resources and Mineral Reserves (May 2014).

The Pampacancha satellite deposit provides higher copper equivalent mill feed where mining is planned to start in 2021 Q2, following the agreement reached with the local community of Chilloroya in 2020. Additional exploration concessions recently acquired by Hudbay within trucking distance of the processing plant offer potential opportunities to mine high grade skarn deposits in addition to their potential for finding a new large porphyry system.

Thorough reconciliations from mineral reserve estimates to mill credited production will continue to be closely monitored in order to continue to validate the performance of the reserve model. The updated mineral reserve estimates show an increase in total tonnage and in all metal content as result of the integration of successful drilling results from Constantia Norte, a refinement of the mine design, the correction of an under-estimation bias in gold grade in the resource modelling database and other minor adjustments to the modelling methodology. A reduction in the measured, indicated and inferred mineral resource tonnage is compensated by a material increase in grade in all resource categories. The lower tonnage and higher grade resource estimates result from the replacement of low grade maginal mineralization in host rocks mostly classified as inferred by high grade skarn mineralization mostly added from the positive drilling results at Constancia Norte. . The improved grade of the mineral resource estimates increases the potential for economic viability when the mineral reserve estimates are depleted.

Other than the risks described in this Technical Report, the general political and social risks associated with perating in Peru and the other risk factors described in Hudbay's most recent annual information form dated March 29^th^ 2021, there are no known significant risks and uncertainties that could reasonably be expected to materially affect the potential development of the mineral reserve and resource estimates in this Technical Report.

1.17. RECOMMENDATIONS

Since 2017, Hudbay has consolidated a very prospective regional land package within trucking distance of the Constancia mining and milling infrastructures. Considerable progress has been made in concluding exploration agreements with certain of the local communities resulting in two exploration agreements being signed and obtaining a drill permit to test a skarn target at the Quehuincha Norte prospect. Additional community agreements and permits will be required to commence exploration activities on the other prospective lands.

A significant portion of the mineralization delineated during the Constancia Norte drill program is still classified as inferred resources and would be amenable to reserve conversion with infill drilling. In addition, some of the mimeralization not included in the open pitable resource may be amenable to production through a small satellite underground operations that will be further investigated through mining trade-off studies.

Recent metallurgical testing has helped validate the forecast model for mill throughput and metals recovery at Pampacancha and various process optimization projects have come to fruition and are now integrated in the life of mine plan supporting the updated mineral reserve estimates.. Optimization studies will continue with a focus on reducing levels of zinc and lead contanimants and increasing copper concentrate grades.

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2. INTRODUCTION

This Technical Report has been prepared for Hudbay to support the public disclosure of mineral resources and mineral reserves at Constancia mine as at December 31, 2020.

Hudbay is an integrated Canadian mining company with assets in North and South America principally focused on the discovery, production, and marketing of base and precious metals. Hudbay's objective is to maximize shareholder value through efficient operations, organic growth and accretive acquisitions, while maintaining its financial strength.

Hudbay's operations at Constancia include the Constancia and Pampacancha pits, an ore processing plant, a Waste Rock facility, a Tailings Management facility and other ancillary facilities that support the operations.

At the date of this Technical Report, the Constancia mine is in steady production. The pre-stripping started in March 2014 and the concentrator ramp-up started in December 2014. Commercial production was achieved on April 30, 2015 and the operations continue with management's expectations in terms of ore, grade, recovery and cost.

2.1. QUALIFIED PERSON

This Technical Report has been prepared in accordance with NI Form 43-101 F1. The QP who supervised the preparation of this Technical Report is Olivier Tavchandjian, P.Geo., Vice-President Exploration and Geology of Hudbay.

Olivier Tavchandjian is not independent of Hudbay, and this is not an independent technical report, but Hudbay is a "producing issuer" as defined in NI 43-101. As such, this Technical Report is not required to be prepared by or under the supervision of an independent QP.

Mr. Tavchandjian has been directly involved on a regular basis with the exploration, geology, resource-modelling, mine planning as well as with the estimation of operating and capital costs for the Constancia mine and the Pampacancha deposit. Mr. Tavchandjian has visited the Constancia operation on a regular basis since September 2017 as well as the core storage facilities and the relevant internal and external laboratories and has directly overseen the mineral resource and reserve estimation process.

2.2. SOURCE OF INFORMATION

Geology and Mineral Resources sources of information are: drilling log and sample data, blast hole sample data and in-pit geology mapping.

Mineral Reserve sources of information are the Mineral Resources, actual production and monitoring data (since 2014), budget projections.

Metallurgy, processing, and economic sources of information are the actual operating data acquired since copper production commenced in 2014, operating budget estimates.

All other relevant information has been updated with information or reports provided and translated by senior site personnel.

Multiple participants have worked on this Techncail Report. Discussions were held with personnel from Hudbay and Hudbay Peru:

Javier Del Rio, Vice- President South America Business Unit
Javier Toro, Director - Technical Services
Marc-Andre Brulotte, Manager - Resource Evaluation

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Jose Escudero, Director Social and community relations
Carlos Castro, Director - Corporate Affairs and Social Responsibility
Marcial Medina, Manager - Process Plant
Engels Trejo, Manager - Technical Services
Julio Roncal, Manager - Environmental
Carlos Salazar, Manager - Exploration
Richard Rodriguez, Superintendent - Mine Geology

Table 2-1 lists the participants to this Technical Report under Olivier Tavchandjian as QP:

TABLE 2-1: TECHNICAL REPORT PARTICIPANTS

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2.3. UNIT OF MEASURE ABBREVIATIONS USED IN REPORT

(%) - Percentage
(g/t) - Grams per tonne
(Mt) - Million tonnes
μg/m3 - Micrograms per cubic meter
Ha - Hectare
Km - Kilometres
kV - Kilovolts
kWh/m3 - Kilowatt - hours per cubic meter
kWh/t - Kilowatt per tonne
m - Meter
m3 - Cubic meter
m3/day - Cubic meter per day
m3/h - Cubic meter per hour
m3/year - Cubic meter per year
Mt/yr ore - Million tonnes per year of ore
Mtpa - Million tonnes per annum
MW - Megawatt
MW-h/y - Mega watt hours per year
MWhr - Mega watt hour
ppb - Parts per billion
Ppm - Parts per million
sq-km - Square kilometres
t - Tonne
t / m3 - Tonne per cubic meter
t/a - Tonnes per annum
t/BCM - Tonnes per billion cubic metres
t/Hole - Tonnes per hole
t/m3. - Tonnes per cubic meter
Tonne/Hr - Tonnes per hour
Tpd - Tonnes per day
US$/dmt - US dollars per dry metric tonne
US$/oz - US dollars per ounce
US$/wmt - US dollars per wet metric tonne
(AAN) Andesite
(AAS) Atomic Absorption Spectroscopy
(AVRD) Absolute Value of the Relative Difference
(BAP) Biodiversity Action Plan
(CC-CR) Constancia Condemnation Drilling
(CG) Constancia Geotechnical Drilling
(CH) Constancia Hydrogeological Drilling
(CM) Constancia Metallurgical Drilling
(CO) Constancia Infill Drilling
(CR) Reverse Circulation
(CR) Reverse Circulation

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2.4. ACRONYMS AND ABBREVIATIONS USED IN REPORT

(CRMs) Certified Reference Materials
(EIA) Environmental Impact Assessment
(FEED) Front End Engineering and Design
(GWI) Ground Water International Consulting Hydrogeologists
(HA) Hectar
(HG) High Grade Material
(ICP-AES) Inductively Coupled Plasma - Atomic Emission Spectrometry
(IP) Induced Polarisation
(IRA) Inter-Ramp Angle
(JV) Joint Venture
(LG) Low Grade Material
(LTEP) Long Term Equilibrium Prices
(LTEP) Long Term Equilibrium Prices
(MASL) Metres Above Sea Level
(MCCs) Motor Control Center
(MG) Medium Grade Material
(MINEM) Peruvian Ministry of Energy and Mines
(MMP)Micro Monzonite Porphyry
(MP1) Monzonite Porphyry 1
(MP2) Monzonite Porphyry 2
(MR) Mining Royalty
(Mt) Million tonnes
(NAG) Non-Acid Generating Material
(NI) National Instrument
(PAG WRF) Potentially Acid Generating Waste Rock Facility
(PAG) Potentially Acid Generating Material
(PFS) Pre-Feasibility Study
(PG) Pampacancha Geotechnical Drilling
(PO, PR) Pampacancha Exploration-Infill Drilling
(QC) The Quality Control
(QMP) Quartz Monzonite Porphyry
(RE) Relative Error
(RMR) Rock Mass Rating
(RQD) Rock Quality Designation
(SD-EIA) Semi-Detailed Environmental Impact Assessment
(SEIN) Peruvian National Interconnected Electric System
(SMT) Special Mining Tax
(SMU) Selective Mining Unit
(SO, SR) Chilloroya South Exploration
(SO, SR) Chilloroya South Exploration
(TMF) Tailings Management Facility
(UCS) Unconfined Compressive Strength
(UEA) Unidad Economica Administrativa
(UO) Uchucarco Exploration
(WRF) Waste Rock Facility

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3. RELIANCE ON OTHER EXPERTS

Hudbay has followed standard professional procedures in preparing the contents of this Technical Report. Data used in this Technical Report has been verified and the author has no reason to believe information has been withheld that would affect the conclusions made herein.

The information, conclusions, opinions, and estimates contained herein are based on:

Information available to Hudbay and the author at the time of preparation of this Technical Report; and
Assumptions, conditions, and qualifications as set forth in this Technical Report.

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4. PROPERTY DESCRIPTION AND LOCATION

4.1. THE PROPERTY LOCATION

The Constancia mine is located in the south-eastern Andes of Peru, in the Chamaca, Livitaca and Velille districts, province of Chumbivilcas, department of Cusco. The property is approximately 600 km southeast of Lima at elevations of 4,000 to 4,500 masl. Figure 4-1 shows the general Constancia mine location. Road access to the property is from Arequipa (7 hours by road) or Cusco (6 hours by road). Geographic coordinates at the centre of the property are longitude 71°47' west and latitude 14°27' south.

FIGURE 4-1: CONSTANCIA MINE LOCATION

4.2. MINERAL RIGHTS

4.2.1. PERUVIAN MINING LAW

The General Mining Law of Peru defines and regulates different categories of mining activities, from sampling and prospecting to commercialization, exploration, exploitation, general labour, processing and mining transportation (concessions granted by the Peruvian state are only required to conduct the latter five activities). Mining concessions (which grant the rights to explore and exploit mineral deposits) are granted using UTM coordinates to define areas generally ranging from 100 ha to 1,000 ha in size.

Mining concessions are irrevocable and perpetual, as long as the titleholder complies with two main mining obligations: (i) pay an annual validity fee ("Derecho de Vigencia"); and, (ii) achieve an annual minimum production level on the concessions.

The "Derecho Vigencia" or validity fee is an annual maintenance fee of $3/ha (for metallic mineral concessions at general regimen) for each concession actually acquired, or for a pending application (petitorio), payable by June 30th of each year. Non-compliance with this obligation for two (2) consecutive years results in the cancellation of the respective mining concession.

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The General Mining Law of Peru sets a time limit within which the holder of the mining concession must comply with obtaining the minimum production level. If the established production is not met by the deadline, then the holder of the mining concession must pay a penalty.

There are currently two systems for calculating the amount to be paid as a penalty. Each of these regimes will be applicable depending on whether the mining concession analyzed was granted until 2008 or later. Each of these regimes is explained below:

(i) Old Regime - Concessions granted until 2008: The minimum production that the holder of the mining concession must reach in these concessions is US$ 100 per hectare per year, for a term of six years from the date of granting the title. If the holder of the mining concession does not comply with said production, he or she must pay a penalty equivalent to US$ 6.00 per year per hectare. This penalty will be increased to US$ 20.00 per year per hectare if the non-compliance persists after year twelve.

Under this regime, if the holder of the mining concession fails to pay the penalty for two consecutive years, this will result in the cancellation of the respective concession. It should also be noted that the holder of the mining concession may be released from payment of the aforementioned penalty if he or she provides proof to the mining authority that he or she has made investments in the concession in question for an amount equivalent to ten times the amount of the applicable penalty.

(ii) New Regime - Concessions granted from 2009: The minimum production to be reached by the holder of the mining concession in these concessions is one Tax Unit (UIT) per hectare per year for metal concessions and 10% of the Tax Unit (UIT) per hectare per year for non-metallic concessions, for which it will have a term of 10 years from the date of granting the title. On January 5, 2017, Legislative Decree No. 1320 that modified articles 40 and 41 of the General Mining Law (modified previously by Legislative Decrees No. 1010 and No. 1054), established new terms and amounts to be paid as penalties for failure to reach the minimum production of the New Regime, applicable from January 1, 2019, as detailed below:

(a) If the minimum production is not met by the tenth year after the mining concession was granted, a penalty of 2% of the value of the UIT per year and per hectare must be paid from the following year until the minimum production or minimum annual investment is reached.

(b) If the minimum production is not met by the fifteenth year after the mining concession was granted, a penalty of 5% of the UIT value per year and per hectare must be paid until the minimum production or minimum annual investment is reached from the following year onwards.

(c) If the minimum production is not met by the twentieth year after the mining concession was granted, a penalty of 10% of the UIT value per year and per hectare actually granted until the minimum production or minimum annual investment is reached from the following year onwards.

(d) If the minimum production is not met within thirty years of the mining concession being granted, the concession expires, with no exceptions.

(e) No penalty is required to be paid in each year if an investment of not less than ten times the applicable penalty amount is made.

As of January 1, 2021, the concessions of the new and old regimes that do not demonstrate compliance with the minimum production of the previous yearmust pay a penalty of 2% of UIT per year and per hectare granted for each concession, continuing in case of non-compliance in subsequent years, with the terms, penalty amounts, and conditions detailed in the previous paragraph.

The holder of a mining concession is entitled to all the protection available to all holders of private property rights under the Peruvian Constitution, the Civil Code, and other applicable laws. A Peruvian mining concession is a property-related right; distinct and independent from the ownership of land on which it is located, even when both belong to the same person. The rights granted by a mining concession are defensible against third parties, are transferable and chargeable, and, in general, may be the subject of any transaction or contract.

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To be enforceable, any and all transactions and contracts pertaining to a mining concession must be entered into a public deed and registered with the Public Mining Registry. The holder of a mining concession must develop and operate his/her concession in a progressive manner, in compliance with applicable safety and environmental regulations and with all necessary steps to avoid third-party damages. The concession holder must permit access to those mining authorities responsible for assessing that the concession holder is meeting all obligations established in the applicable regulation.

4.2.2. MINING CONCESSIONS

Excluding the concessions included in the Transfer Agreement and the Transfer Option and Mining Assignment Agreements, subscribed by Hudbay Peru S.A.C. with Panoro Apurimac S.A. and a private Peruvian company respectively, in December 2017, Hudbay has a total of sixty-six concessions in the province of Chumbivilcas, in the department of Cusco which belong to the Constancia Mine. These sixty-six concessions occupy an area of 43,536.7769 Ha.

Most of the known mineralization of Constancia mine is located in the following concessions: Katanga J, Katanga Q, Katanga K, Peta 7, Peta 6, Constancias 13 and Constancias 8.

Figure 4-2 shows the concessions corresponding to the UEAs Constancia and Pampacancha and the concessions included in the Transfer Agreement and the Transfer Option and Mining Assignment Agreements.

Administrative Economic Units ("Unidad Económica Administrativa" - "UEA")

Fourteen mining concessions out of the sixty-six concessions comprising the Constancia mine have been grouped together in an UEA called "Constancia", while ten concessions have been grouped together in an UEA called Pampacancha.

UEAs are groups of two or more mining concessions of the same class and nature which allow the respective holder to work only in one or some of the concessions, instead of working simultaneously in all of those encompassing the UEA. However, the holder is obliged to obtain the minimum production levels corresponding to all the mining concessions comprising the UEA in those mining concessions which are exploited.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

FIGURE 4-2: CONSTANCIA MINE CONCESSION BOUNDARIES

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

The UEA Constancia (code N° 01-00029-83-U) - previously called Katanga Este - was originally approved by Resolución Directoral No. 148/86-EM-DCM dated 1 September 1986 comprising only the "Katanga J" and "Katanga K" mining concessions. Through Resolución Jefatural N° 5404-2006-INACC/J issued on the 14th of December 2006, eleven mining concessions were added, bringing a total of 13 concessions to the UEA. On December 14, 2015, the name was changed to UEA Constancia. In December 2016, by Resolución de Presidencia No. 2207-2016-INGEMMET-PCD-PM one more concession was added to this UEA. The total of the area of this UEA is 6,146.6569 Ha.

Table 4-1 shows the names, codes, dates and areas of the concessions comprising the UEA Constancia. All the properties are located on the Livitaca (29-S) map sheet. The UEA Pampacancha (code 010002216U) was approved by Resolución de Presidencia No. 1899-2016-INGEMMET-PCD-PM on December 19, 2016. The total area of this UEA is 8,085.9011 Ha

Table 4-2 shows the names, codes, dates and areas of the concessions comprising the UEA Pampacancha.

TABLE 4-1: HUDBAY CONCESSIONS

It is worth noting that all of the mining concessions involving the Constancia and Pampacancha UEAs have been titled and have also been recorded with the Public Registries in the name of Hudbay Peru S.A.C. Likewise, all of these concessions are currently in good standing and concession fees and applicable penalties have been paid for the calendar year 2020. All mining concessions are in good standing (i.e. the mining concessions are in full force and effect) as of January 1, 2021 and there are no foreseeable circumstances that could result in the mining concessions being cancelled by the Peruvian State.

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TABLE 4-2: UEA PAMPACANCHA

RECENTLY ACQUIRED MINING CONCESSIONS

In December 2017, Hudbay Peru S.A.C. executed a Transfer Agreement with Panoro Apurimac S.A. and a Transfer Option and Mining Assignment Agreements with a private Peruvian company. Through these agreements, Hudbay Peru S.A.C. acquired title over 24 mining concessions.

The Transfer Agreement was subscribed by Hudbay Peru S.A.C. and Panoro S.A. in December 28, 2017. This contract was subsequently entered in the Public Registries. The total area of the concessions mentioned in this agreement is 3,961.7457 Ha. Table 4-3 shows the names, codes, dates and areas of the concessions transferred by the aforementioned Agreement:

The Transfer Option and Mining Assignment Agreement was subscribed by Hudbay Peru S.A.C. and a private Peruvian company on December 14, 2017. This contract was subsequently entered in the Public Registries. The total area of the concessions mentioned in this agreement is 5970.3258 Ha. Table 4-4 shows the names, codes, dates and areas of the concessions transferred by the aforementioned Agreement:

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TABLE 4-3: KUSIORCCO CONCESSIONS

TABLE 4-4: CABAILLITO AND MARIA REYNA CONCESSIONS

As of January 1, 2021, these concessions are up to date in the payment of the "Derecho de Vigencia" and penalties.

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4.2.3. SURFACE LEGAL RIGHT ACCESS

The Constancia mine area is located within six lands owned by Hudbay. For rights purposes, Norsemont previously purchased the Fortunia property that covers most of the main resource area.

On March 19, 2012, Hudbay and the Uchucarco Community entered into an agreement granting Hudbay a right to use 256.50 hectares located in the so-called "Sayhualoma" sector through the life of the mine. Likewise, one significant plot of 750 ha in the so-called "Ichuni" sector has been purchased by Hudbay from the Chilloroya Community on April 12th, 2012. Hudbay has also purchased two pieces of land of 1361.87 ha and 7.73 ha, respectively called "Quinsachata Sub Lote 1" and "Quinsachata Sub Lote 2" for relocation purposes. Through these agreements, Hudbay has secured all surface land rights required for the construction and development of the Constancia Mine.

On May 9, 2012, Hudbay entered into an agreement with the Chilloroya Community in order to access and use for exploration purposes 6,148 hectares covering the Pampacancha deposit. This agreement was in force until 2015.

On Sep 25th, 2018, Hudbay entered into an agreement with the Quehuincha community in order to access and use, for exploration purposes, 1,053 hectares covering the Quehuincha North Project.

During the first half of 2019, Hudbay successfully concluded the biennial review of LOM Agreements with both, Chilloroya and Uchucarcco communities, which allowed to the company starts the negotiation talks on Pampacancha and exploration targets of Caballito, Maria Reyna and Kusiorco.

On March 3rd, 2020, Hudbay signed an agreement with the Chilloroya community for the right of use of 474.32 hectares in Pampacancha area. Likewise, negotiations with individual landholders (possessors) are ongoing, having closed 28 of 29 agreements. Table 4-5 provides a summary of the land rights hold by Hudbay.

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TABLE 4-5: SURFACE LEGAL ACCESS RIGHTS TABLE

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4.3. THE TERMS OF ANY ROYALTY OR PAYMENTS

Hudbay's mining concessions are free and clear of recorded liens, encumbrances or agreements except for a 0.5% Net Smelter Return ("NSR Royalty") that has been established over the sale of the mineral products obtained from the following concessions: Katanga Q, Peta 5, Peta 6, Peta 7 and Peta 17. Such NSR Royalty has been granted in favor of Compañia Minera Katanga S.A. ("Katanga") and Minera Livitaca S.A. ("Livitaca") up to a maximum amount of US$ 10,000,000 (ten million and 00/100 Dollars of the United States of America)., this has already been paid and the NSR Royalty has expired.

By virtue of the Mining Concessions Transfer Agreement dated December 31, 2008, and its addendum, dated February 19, 2009, both executed by and between Norsemont and Rio Tinto, a legal mortgage up to the amount of US $500,000 has automatically been established (and is currently registered) over the following mining concessions in order to secure the timely payment of the additional or contingent payment.

The Transfer Agreement subscribed with Panoro S.A. in December 28, 2017 establishes a lien in the form of a 2% Net Smelter Return ("NSR Royalty") over the sale of the mineral products obtained from the following concessions: Aluno Dos 2002, Aluno Tres 2002, Valeria Sesentaicinco 2004, Valeria Sesentaicuatro 2004, Valeria Sesentaidos 2004, Valeria Sesentainueve 2004, Valeria Sesentaiseis 2004, Valeria Sesentaisiete 2004, Valeria Sesentaitres 2004, Valeria Sesentaiocho 2004, PML1.

The Transfer Option and Mining Assignment Agreement subscribed with a private Peruvian company in December 14, 2017 establishes that in case Hudbay Peru S.A.C. exercises the transfer option, it must sign a Royalty Agreement for the following concessions: Doris 7, Doris 8, Katanga, Juana 4, Leviatan, Maria Reyna I, Maria Reyna II, Maria Reyna III, Monte Rojo No4, Monte Rojo No6, Monte Rojo No7 and Peta.

Hudbay has LOM agreements with the two local communities, Chilloroya and Uchucarco. These agreements were part of the land access and purchase agreements and they outline a commitment to sustainable development in the respective communities.

4.4. ENVIRONMENTAL PROPERTY LIABILITIES

During the preparation of the environmental baseline for the ESIA, Knight Piesold Consultores S.A. determined the existence of 21 mining environmental liabilities within the Constancia mine in the areas called Sacsa Orcco, the San José pit, Chilloroya and the Yanaccocha Lake.

Since these environmental liabilities were generated by former operations and titleholders, Hudbay is not obligated to undertake any measure for its reclamation or closure.

Additionally, as a result of the preparation of the environmental baseline of the first amendment of the ESIA, an extra mining environmental liability was identified in the San José area. Hudbay Peru S.A.C. declared it before the competent authority. This area of the San José pit is part of the final closure plan of the Constancia mine.

The Second amendment to the EISA, approved in 2015, is the currently environmental management instrument that incorporates the exploitation of the Pampacancha pit. After concluding with the land permits and the sectorial requirements, we obtained de exploitation start authorization on December 2020. Likewise, in 2019 and 2020 exploratory drilling was developed to the northwest of the Constancia pit.

A Third Amendment to the Eisa, that is Currently Under Final Review for Validation by the Environmental Authority Includes Additional Drilling to the Northeast of the Constancia Pit for 2022.Risk Factors Property

Hudbay has secured all legal rights on the mining concessions comprising the Constancia mine as well as on the lands required for the development of this mine. Considering that such rights have been registered before the Public Registries and are enforceable before the Peruvian state and third parties, Hudbay does not foresee any risk in losing its rights over the Constancia Mine, provided that the legal obligations applicable to keep such properties in full force and effect are duly complied with.

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5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1. ACCESSIBILITY

The Constancia mine is located in Chamaca and Velille districts, Chumbivilcas province, Cusco region, an approximate distance of 634 km southeast from Lima, capital of Peru, with an altitude ranging between 4,000 and 4,500 metres. The mine is accessible from Lima by air, via either Arequipa or Cusco, and then by vehicle over paved and gravel roads. Figure 5-1 below shows the main route from Arequipa and Cusco to the mine.

FIGURE 5-1: LOCATION OF THE CONSTANCIA MINE

The routes with approximate distances and driving times are shown in Table 5-1.

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TABLE 5-1: APPROXIMATE DISTANCES AND DRIVING TIMES TO CONSTANCIA

5.2. CLIMATE

Climate at the Constancia site can be classified as humid and seasonably cool with well-defined, rainy, and drier seasons. The majority of the precipitation typically occurs over the period of October to April and the summer months. Elevation and physiograpy also influence the climate. The length of the operating season is not affected by the climate and mitigation measures are in place to address the challenges presented by the rainy season.

Following is a list of important climatological data:

Annual Precipitation:
Annual average precipitation: 1,000 mm
Annual maximum precipitation: 1,353 mm
Annual minimum precipitation: 590 mm
Average wet season (October to April) precipitation: 932 mm
Average dry season (May to September) precipitation: 72 mm
Air Temperature:
Daily average maximum: between 13 and 16 °C
Daily average minimum: between (-11 and 0°C)
Average Annual Evaporation:
Potential annual evaporation: 961 mm
Annual evaporation from existing ground: 480 mm
Evaporation from dry tailing: 455 mm
Evaporation from wet tailing surface: 865 mm

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5.3. LOCAL RESOURCE

The main economic activity in the area is subsistence agriculture and cattle farming. Food and basic supplies can be obtained in Yauri/Espinar (62,000 people). Cusco (population 367,000 people) and Arequipa (population 864,000 people) are the nearest major centers, 6 and 7 hours drive from the mine, respectively.

Several permanently flowing streams which supply good flows are present in the area. More significant water sources include the Apurimac or Chilloroya rivers.

Near the community of Chilloroya, the local population is engaged in the informal mining of superficial gold. The company and members of the community have raised concerns with Peruvian authorities regarding environmental and labour conditions related to the informal mining activities. For example, levels of mercury found in sediments near a rudimentary gold processing area exceed international standards for the protection of aquatic life.

5.4. INFRASTRUCTURE

Arequipa and Cusco are the nearest major centres, and both are over 300 km from the Mine site by road. A public, but privately managed, landing strip is located in Espinar, about 2 hours from the mine. The air strip is paved and maintained by Glencore, and Hudbay has authorization to use the strip. Small commercial and/or charter flights can use the air strip for daytime operations.

The road to Yauri (in the general vicinity of Glencore's Antapaccay Mine) was upgraded to meet construction and life of mine transportation requirements. From Yauri, the road system is also currently used for supplies to several operating mines in the region. It is routed through the Arequipa area to the port at Matarani which is currently also used to ship copper concentrates from other mines in Southern Peru.

Power supply to Constancia is brought from a new 220kV transmission line 69 Km from Constancia that was built and is owned by Hudbay but operated and maintained by a third party.

Further details regarding the mine infrastructure can be found in Section 18 of this Technical Report.

5.5. PHYSIOGRAPHY

The Constancia region hosts fluvioglacial valleys, flood plains, hills and mountains with either moderate slopes or steep terrain.

Natural hazards in the area were identified through photo interpretation of the World View satellite image taken on June 19, 2012, with a spatial resolution of 0.5 m. The assessment identified six main geodynamic processes: karst geoshapes, rock falls, hillside erosion, river erosion, flood plains and soil saturation. In general, these processes occur locally; on the left and right margins of the Chilloroya river, in the Sacrane, Cunahuiri, Casanuma, Pumacocha, Huayllachane, Telaracaca, Ccatunhuaycco streams, and close to the lakes located northeast of the Mining Unit (Quesoccocha, Pincullune and Sillco).

According to the seismic hazard study conducted for the Constancia mine (Alva J. 2013), major earthquakes recorded intensities from VI to X on the Modified Mercalli scale (MM). Results of the probabilistic seismic analysis indicate that the value for the maximum horizontal acceleration for the OBE (Operating Basis Earthquake) is equal to 0.24 g, corresponding to a PGA (peak ground acceleration) for a return period of 475 years. The value of maximum horizontal acceleration for the MDE (Maximum Design Earthquake) is equal to 0.37 g, corresponding to a PGA for a return period of 2,475 years. The deterministic seismic analysis indicates that the largest acceleration for the Constancia Mining Unit area is 0.44 g, corresponding to the movement of the earth caused by the earthquake of intraplate deep subduction based on the average plus one standard deviation, called MCE (maximum credible earthquake)

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6. HISTORY

6.1. EARLY HISTORY

Copper and gold were exploited at Katanga, located approximately 3km northwest of Constancia outside the Constancia property boundary from early last century to the early 1990s.

The Katanga deposit consists of narrow skarn bodies developed in the contact between marbles and monzonite stocks, with Cu, Ag and Au mineralization in hypogene sulphides.

6.2. PRIOR OWNERSHIP

The San José Prospect (now part of the Constancia Mine) was explored by Mitsui during the 1980s with a focus on locating more high-grade ore to add to the nearby Katanga mine operation. Exploration consisted of detailed mapping, soil sampling (1,949 samples), rock chip sampling (1,138 samples), ground magnetic and induced polarization (IP) surveys with several drilling campaigns. Drilling was mainly located in the western and southern sides of the prospect. Mitsui completed 24 drill holes (4,190.5m) and Minera Katanga completed 24 shallow close-spaced drill holes (1,239.8m) at San José.

In 1995, reconnaissance prospecting identified evidence for porphyry style mineralization exposed over an area 1.4 x 0.7 kilometres open in several directions, with some Cu enrichment below a widespread leach cap developed in both porphyry and skarn. Negotiations to acquire an interest in the property were unsuccessful at this time.

In May 2003, the area was revisited by Rio Tinto and the presence of a leached cap and potential for a significant copper porphyry deposit was confirmed. Negotiations with Mitsui, Minera Livitaca and Minera Katanga resulted in an agreement being signed on October 2003 with the underlying owners. The agreements included a joint venture (JV) option between Rio Tinto and Mitsui and Purchase Option Agreements with Minera Livitaca and Minera Katanga to acquire 100% interests in their property. Rio Tinto commenced exploration in December 2003.

The Rio Tinto exploration activities consisted of geological mapping, soil, and rock chip sampling, surface geophysics (magnetics and IP) and completed 24 diamond drill holes. In late 2004, Rio Tinto sought partners for the Constancia prospect, mostly because the property was thought to be too small.

Eventually, Norsemont entered into negotiations with Rio Tinto and these negotiations led to an agreement in early 2005. The first Norsemont geologists visited the property in June of 2005. By 2009, Norsemont had entered into a transfer agreement with Mitsui (November 2007) for the purchae of their 30% interest in the Constancia Project and had exercised the final option with Rio Tinto to acquire the remaining 19% interest in Constancia during 2008 providing 100% ownership of the Constancia Project.

During the period between 2005 and 2010 Norsemont's exploration activities were mainly focused on delineating the known Constancia and San José zones, although surface exploration of adjacent areas were also carried out, specifically focusing on the two new Cu-Au-Mo targets leading to the discovery of the Pampacancha porphyry-related skarn and Chilloroya South porphyry.

Norsemont got approval from the Environmental and Social Impact Assessment for the Constancia project by the Energy and Mines Ministry of Peru on November 24th, 2010.

6.3. HUDBAY OWNERSHIP AND PROJECT DEVELOPMENT

In March 2011, Hudbay acquired a 100% interest in the Constancia Project through its acquisition of all of the outstanding shares of Norsemont by way of a formal take-over bid.

Following the acquisition of the Constancia project in 2010, Hudbay's exploration activities focused on delineating the Pampacancha deposit.

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From January to November 2011, there was a campaign of infill drilling and exploratory growth in Pampacancha, in order to define measured and indicated resources of Cu-Au-Mo. The most important was the identification at a north-western area of high grade skarn Cu-Au below post mineral rocks, which has meant a substantial increase in the amount of resources and the deposit grade.

On August 8, 2012, Hudbay's board of directors approved a US$1.5 billion investment to fund the development and construction of its Constancia project in Peru. Following substantial completion of the detailed engineering in the third quarter of 2013, the board approved a revised capital cost estimate for the project of US$1.7 billion.

Hudbay developed Environmental and Social Impact Assessment for the project that were approved by the Energy and Mines Ministry of Peru on August 20th, 2013 and April 15th, 2015.

Construction of the Mine-Milling complex of Constancia was completed on time and on budget at the end of 2014 with first ore mined in the fourth quarter of 2014.

6.4. PRODUCTION HISTORY

The start of 2015 brought substantial change to the Peru team with the transition from a construction project to operations. Constancia achieved commercial production in the second quarter of 2015, and subsequently ramped up to full production in the latter half of the year. Over the course of 2015, Constancia mine operations produced 105,897 tonnes of copper concentrate and 47,263 ounces of precious metals in concentrate.

The port expansion at Matarani was completed at the end of June 2016, which improved access to Hudbay's designated pier and reduced port costs.

Since 2015, Constancia mining operations and cost optimization continued as planned. The production performance over the past 5 years is summarised in Table 6-1 below. The reduction in copper, gold and silver production since 2016 is related to the planned transition from supergene high grade mineralization in the early phases of production to the lower grade deeper hypogene mineralization representing an increasing portion of the mill feed. This trend will be reversed and metal production will increase over the next 5 years due to the addition of the high grade mineralization from the Pampacancha deposit. The increase in molybdenum production is related to a significant improvement in both throughput and recoveries of the molybdenum concentrator.

TABLE 6-1: PRODUCTION PERFORMANCE

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7. GEOLOGICAL SETTING AND MINERALIZATION

7.1. DISTRICT GEOLOGY

The oldest rocks in the area (Figure 7-1) correspond to a sequence of white, red, violet or grey medium-grained sandstones with intercalations of reddish mudstones of the Lower Cretaceous Chilloroya Formation (also referred to as the Murco Formation). The Arcurquina Formation discordantly overlies the Chilloroya Formation and correlates with the Upper Cretaceous Ferrobamba Formation. These rocks are exposed in a north-south trend, 15 km long by 5 km wide, comprising a sequence of limestones, calc-arenites and lenses of conglomerates.

These sedimentary formations have been intruded by plutonic rocks belonging to the Andahuaylas-Yauri Batholith of the Oligocene Age. The batholith composition grades in composition varies from dioritic to granodioritic in composition, with plagioclase and orthoclase feldspar, quartz, hornblende, biotite, apatite, zircon and sphene being the main rock-forming minerals. Small seams, veins and lenses of massive magnetite skarn related to the emplacement of the batholiths are common.

Several monzonitic stocks, dykes or laccoliths intrude and cross-cut all the lithologies mentioned above. Where these rocks have intruded limestones, it is common to find mineralized skarns, some of which contain Cu-Au-Ag mineralization such as those at the Katanga mine. Some of the stocks have characteristics typical of porphyry copper deposits such as at Constancia.

7.2. PROPERTY GEOLOGY

7.2.1. CONSTANCIA

The Constancia porphyry copper prospect is located on the eastern margin of the Andahuaylas- Yauri Batholith, approximately 3 km southeast of the old Katanga mine (Figure 7-2).

STRATIGRAPHY

The oldest stratigraphic unit recognized on the prospect comprises clastic sediments corresponding to the Chilloroya Formation, consisting of a sequence of white, red, violet and grey medium-grained sandstones with intercalations of reddish mudstones. Immediately overlying this basal unit are massive, grey micritic limestones with minor intercalations of shales, which outcrop sporadically around the prospect and near the contacts with monzonite, sometimes occurring as roof pendants.

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FIGURE 7-1: SIMPLIFIED GEOLOGY OF THE ANDAHUAYLAS YAURI AREA

FIGURE 7-2: GEOLOGICAL MAP OF THE CONSTANCIA DEPOSIT

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This unit corresponds to the Arcurquina Formation (locally known as Ferrobamba Formation). When in contact with intrusive rocks, these alter to marble or pyroxene diopside-garnet-magnetite-epidote skarn, with or without sulphides. The limestones and skarns dip gently south-east, away from the principal monzonite in the southern part of the Constancia mineralized zone. The overall thickness of the sedimentary package is unknown.

The determination of the base of this limestone unit may be of importance at Constancia, as this potential contact seems to correlate with the favourable skarn horizon at the ex Tintaya Mine and elsewhere in the region. This contact may be found at depth south and east of the presently drilled area, if not intruded by the monzonite.

The underlying clastic sediments of the Murco Formation, and possibly the upper parts of the further underlying Hualhuani Formation (locally known as Soraya Formation), which lithologically consist of sandstones and siltstones, with occasional calcareous and quartzite horizons, are known to host copper mineralization in several copper systems within the belt. At Mitsui's Quechua deposit near Tintaya, and also at the Antillas porphyry copper deposit near Antabamba (north of Constancia), these rocks are the main host for the copper mineralization. The same style of mineralization occurs at Haquira (10 km south of Las Bambas), where Murco sediments are the host for most of the copper oxide resource of the system.

At Constancia, these types of clastic sediments, especially the Murco or Chilloroya Formation which is the host for the copper oxide mineralization at Haquira, have been identified in the southern sector of the property, around the Chilloroya village, where recent surface exploration has identified evidence of porphyry-related copper-gold-molybdenum mineralization associated with the sediments and altered porphyry rocks. Common to these clastic sediments is their iron-stained colouration, which comes from oxidation of former disseminated pyrite.

Glacial moraines cover the northern and eastern margins of the Constancia deposit. To the east these moraines entirely cover potentially important extensions of copper mineralization along broad east-west structural zones.

INTRUSIONS

Multiple phases of monzonite and monzonite porphyry characterize much of the surface area of the prospect, as well as dominating the rock types observed in the drilling to date.

At least four main phases of intrusion are recognized, with the second oldest being associated with the main mineralization event. From oldest to youngest they are:

Diorite: while not part of the intrusive event associated with mineralization, the Andahuaylas- Yauri Batholith forms the 'Intrusive Basement' to the Constancia deposit.

Monzonite Porphyry 1 (MP1): this unit outcrops as a large stock on the Constancia hill, extending west to San José. It hosts most of the porphyry-related mineralization. It is characterized by abundant plagioclase phenocrysts with hornblende in elongated crystals.

Micro Monzonite Porphyry (MMP): characterized by a fine-grained texture with plagioclase crystals, biotite and magnetite in the light gray matrix. This body outcrops as a stock in the south, but is more widespread west of the deposit, including the San José zone.

Quartz Monzonite Porphyry (QMP): this unit occurs mostly as wide, north- south to north-northwest trending dykes with dark, fine-grained chilled boundaries. Abundant plagioclase with tabular, well- formed hornblende crystals occur as phenocrysts in a greenish matrix. These vertical dykes are up to 60 m wide. There is no mineralization associated with the dykes.

Monzonite Porphyry 2 (MP2): this dyke-like monzonite porphyry outcrops between the Constancia and San José zones. It is characterized by abundant plagioclase phenocrysts sub-rounded in a whitish matrix with little magnetite and biotite. The monzonite porphyry occurs as dykes up to 150 m wide, which strike north-south with a steep easterly dip.

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Andesite (AAN): dark-gray, aphanitic rock with plagioclase and hornblende phenocrysts. This occurs mainly as narrow dyke-like bodies, some of them close to the contacts with quartz monzonite porphyries.

STRUCTURAL GEOLOGY

As with most porphyry copper complexes, structural activity at Constancia has played the most significant role in preparing and localizing the hydrothermal alteration and accompanying copper- molybdenum-silver-gold mineralization, including skarn formation.

Major inter-mineral and post-mineral fracture systems in the deposit area strike northeast. One of the best known is the 'Barite Fault', formed by a number of nearly parallel vein-faults carrying base metal sulphides and barite. These veins have been exploited by artisanal workings throughout the property. This system is clearly visible in the ground magnetic maps, controlling major features including topographic and shear fabric changes in the San José zone.

A second important system strikes north-south. It seems to be more recent than the Barite system, controlling part of the San José deposit and most of the silicified breccias (some of them mineralized) in the system. It shares the same direction as the post-mineral dykes, and may have originated as tension gashes to the Barite direction.

Latest in the sequence is the north-northwest to south-southeast oriented fault system, with the Yanak Fault as the main example. These faults generate wide areas of gouge and milled rock, some of which show high hydraulic gradients.

7.2.2. PAMPACANCHA

STRATIGRAPHY

The main stratigraphic unit in the Pampacancha area is a massive, gray micritic limestone part of the Upper Cretaceous Ferrobamba Formation.

INTRUSIONS

The sedimentary sequence described above is intruded by dioritic porphyry, which is the phase that generated the magnetite skarn that hosts the economic Cu-Au-Mo mineralization. Other intrusive rocks present can be assigned to the unmineralized Oligocene basement dioritic batholiths. This diorite is also intruded by the mineralizing dioritic porphyry (described above). The younger intrusive phases include intra-mineral monzonite intrusions which provide minor local increases in Cu-Au and also locally cut the skarn Cu-Au mineralization.

Diorite (DIO): this unit is characterized by a faneritic texture and in composition shows abundant plagioclase phenocrysts with also hornblende and biotite diorite and limestone constitute the host rock of deposit.

Diorite Porphyry (PDI): this rock is characterized by a porphyritic texture with hornblende phenocrysts and sulfides (pyrite-chalcopyrite-molybdenite) disseminated and veinlets with quartz inside an aphanitic light green matrix.

Monzonite Porphyry 2 (MP2): This rock in Pampacancha occurs as lopolith stock in the middle west of deposits cutting all rock; it is characterized by K-feldspar, plagioclase phenocrysts, hornblende, biotite phenocrysts and magnetite inside a medium gray aphanitic matrix.

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STRUCTURAL GEOLOGY

The structural geology of Pampacancha is a dominant factor for the emplacement of Cu-Mo porphyry-skarn and Au Veins mineralization. The occurrence of porphyry is associated to a NE-SW fault and shear zone system that contains three mineralized structural lineaments components NE-SW, NW-SE and NNE-SSW.

7.3. ALTERATION

7.3.1. CONSTANCIA

POTASSIC ALTERATION

The potassic alteration assemblage is characterized by secondary potassium-feldspar, and by variable amounts of hydrothermal biotite replacing earlier ferromagnesian minerals and rock matrix. Quartz veining is common; especially "A" and "B" type veinlets. Intensity of alteration is variable, ranging from weak to strong. Hydrothermal magnetite is also present as disseminations and associated with type veinlets in deeper sections. Anhydrite veinlets are also common. Within the potassic alteration zone, chalcopyrite-(bornite)-molybdenite-pyrite mineralization is present in A and B veinlets, and also replacing ferromagnesian minerals or filling fractures.

The high-grade (hypogene) copper mineralization is hosted by a dense A-veinlet stockwork developed in an early porphyry phase. Pyrite/chalcopyrite ratio is typically low and in the order of 1:1 to 2:1. Molybdenite also commonly increases with depth, related to "B" veinlets. Bornite occurs sporadically, especially at deeper levels, sometimes associated with gold values.

PROPYLITIC ALTERATION

Propylitic alteration is transitional to the potassic alteration and extends more than one kilometre from the porphyry intrusive contacts. The propylitic alteration mineral assemblage includes epidote-chlorite-calcite-pyrite-rhodochrosite. Subordinate chalcopyrite is also present, filling fractures or replacing mafic minerals. Sphalerite-galena veinlets and veins are distributed as a halo to the copper-molybdenum mineralization within the propylitic alteration halo, occurring at distances of up to 3 km away from the porphyry copper system.

PHYLLIC ALTERATION

Phyllic alteration forms a pervasive carapace surrounding and sometimes overprinting potassic alteration. This alteration accompanies almost complete destruction of primary rock textures. The phyllic alteration mineral assemblage includes sericite-quartz-pyrite, limited amounts of chalcopyrite and associated occasional "D" veins and veinlets.

7.3.2. PAMPACANCHA

SKARN ALTERATION

Pro-grade proximal magnetite-chalcopyrite-pyrite skarn grades to distal less well-mineralized garnet and pyroxene skarn which is locally overprinted by epidote-bearing retrograde skarn.

POTASSIC ALTERATION

Porphyry diorite and diorite show potassic alteration assemblage characterized principally by quartz veining with rims of K-feldspar stockwork and hydrothermal secondary biotite replacing earlier ferromagnesian minerals; the intensity of alterations is variable, ranging from weak to strong, and the best development of potassic alteration is shown in deep diorite very close to the deposit.

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7.4. MINERALIZATION

7.4.1. CONSTANCIA

The Constancia deposit is a porphyry Cu-Mo-Ag system which includes copper-bearing skarn mineralization. This type of mineralization is common in the Yauri-Andahuaylas metallogenic belt where several porphyry Cu-Mo-Au prospects have been described but not exploited.

Five distinct mineral associations are found within the Constancia Project area, namely:

1) Hypogene, porphyry-style mineralization including disseminated, quartz-vein stockwork and fracture-controlled chalcopyrite-molybdenite mineralization in the intrusive;

2) Hypogene chalcopyrite, rare bornite, galena and sphalerite mineralization in skarns;

3) Supergene digenite-covellite-chalcocite (rare native copper) mainly hosted by intrusive rocks lying below the leach cap;

4) Transitional (mixed) including secondary copper sulphides/chalcopyrite in the monzonite (overlap of 1 and 3, above); and

5) Oxide copper, usually cuprite and tenorite appears at this zone associated to malachite-chrysocolla and rarely azurite.

Of these, the hypogene mineralization (Type 1) constitutes the bulk of the deposit. Skarn (Type 2) is volumetrically smaller, but grades are normally higher, and mineralization occurs at or near the surface. At the contact between the intrusive and limestones, magnetite - garnet skarn develops, while the pyroxene-diopside (garnet-epidote) association is more common in calcareous sandstones and of the Chilloroya Formation. Supergene enrichment (Type 3) occurs immediately beneath, and occasionally as remnants within, the leached cap. The highest copper grades in the Constancia porphyry are typically associated with this and with the skarn zone. The transitional or mix zone (Type 4) corresponds to the zone where the supergene and hypogene mineralization co-exist. Finally, oxide copper mineralization (Type 5) occurs in shallow levels and deeper along fractures.

Two areas of porphyry-style mineralization are known within the project area, Constancia and San José. At Constancia, mineralization is deeper than that observed at San José which occurs at surface. The mineralized zone extends about 1,200 m in the north-south direction and 800 m in the east-west direction. At the present time, both areas are part of the Constancia Pit.

Along the Yanak fault located at the northwestern part of the Constancia Pit, in recent exploration drilling, massive to semi-massive bodies of magnetite skarn with high-grade Cu-grade chalcopyrite mineralization have been identified in the footwall of the fault, the mineralization is related to mineralized porphyry dikes which have also mineralized the dioritic host rock and generated these high-grade skarn bodies.

7.4.2. PAMPACANCHA

The Pampacancha mineralized body has been divided into three sectors, according to lithology and mineralization (Figure 7-3).

Sector 1 is characterized by the predominant magnetite skarn over calc-silicate skarn. Mineralization in sector extends from surface down to about 200m (true depth) with occasional deeper lenses. The skarn is intruded by potassic altered dioritic porphyry and fine-to-medium grained diorite (also exhibiting potassic alteration).

Sector 2 is located immediately north of Sector 1 and is characterized by the presence of a limestone bracketed between two main skarn layers, vertically extending from 70 m to about 250m. The shallower segment of the skarn shows alternations with thin limestone layers, all disrupted by the emplacement of diorite porphyry dikes and endoskarn.

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Sector 3 extends immediately NW of Sector 2 and has the same characteristics as Sector 2, with a limestone body intercalated between two skarn layers, which pinch out northward, eastward and westward.

FIGURE 7-3: PAMPACANCHA

Epithermal mineralization of the low sulphidation quartz-sulphide Au + Cu style, accounts for common supergene enriched Au anomalies, along with other features such as hydrothermal alteration and veins typical of near porphyry settings.

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8. DEPOSIT TYPES

The Constancia deposit is a porphyry Cu-Mo-Ag system which includes copper-bearing skarn mineralization. This type of mineralization is common in the Yauri - Andahuaylas metallogenic belt where several porphyry Cu-Mo-Au prospects have been described but not exploited.

The main porphyry deposits in the belt include Antapaccay (586 Mt @ 0.42% Cu, Glencore Resources and Reserves, as of Dec 2020 Report) and Quechua (~350 Mt @ 0.38% Cu, Mitsui Kinzoku press release, Nov 2015). Several other porphyry prospects are also being explored in the district.

Historically, the belt was better known for copper skarn deposits such as Las Bambas (copper skarn-porphyry) where MMG recently reported resources of 1,429 MT @ 0.63% Cu (MMG Mineral Resources and Ore Reserves Statement as of June 30, 2020). and Coroccohuayco (559 Mt @ 0.68% Cu, Glencore Resources and Reserves, as of Dec 2020 Report).

The Pampacancha deposit is a Cu-Mo-Au porphyry related skarn system, with copper-bearing skarn mineralization. As mentioned above, this type of mineralization is common in the Yauri-Andahuaylas metallogenic belt. As the sedimentary formations were intruded by dioritic and granodioritic batholiths, Cu-Au-Ag mineralization associated to magnetite developed in stockworks of seams, veins and lenses

The Constancia project area geology contains the classic characteristics typical of Andean porphyry copper deposits. The following generalized geological characteristics are the typical targets of exploration compilation efforts and support the classic genesis model of a copper porphyry:

Associated to volcanic arc igneous assemblages.
Ore bodies are associated with multiple intrusions and dikes of diorite to quartz monzonite composition with porphyritic textures.
Breccia zones with angular or locally rounded fragments are commonly associated with the intrusives. The sulfide mineralization typically occurs between or within fragments.
The deposits typically have an outer epidote - chlorite mineral alteration zone.
Quartz - Sericite alteration zone typically occurs closer to the center and may overprint.
A central potassic zone of secondary biotite and orthoclase alteration is commonly associated with most of the ore.
Fractures are often filled or coated by sulfides, or by quartz veins with sulfides.
Closely spaced fractures of several orientations are associated with the highest grade ore.
The upper portions of porphyry copper deposits shows supergene enrichment with the metals in the upper portion being dissolved and carried down to below the water table, where they precipitate.

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9. EXPLORATION

Copper and gold exploration in the Constancia area dates back to the early 1990s. The San José prospect (now part of the Constancia Mine) was explored by Mitsui during the 1980s with a focus on delineation of high-grade ore amenable to process at the Katanga facility. Exploration consisted of detailed mapping, soil sampling (1,949 samples), rock chip sampling (1,138 samples), ground magnetic and IP surveys with several drilling campaigns, mainly located in the western and southern sides of the prospect.

Mitsui completed 24 drillholes (4,190.5 m) and Minera Katanga completed 24 shallow, close-spaced drillholes at San José (1,239.8 m). Through a prospecting program carried out in 1995, an area of 1.4 km x 0.7 km (Constancia main) with porphyry-style mineralization was mapped and determined to be open in several directions.

Rio Tinto carried out exploration in the Constancia area between 2003 and 2004 including geological mapping, soil and rock chip sampling, and surface geophysics (magnetic; 20.3 lines-km and IP 12 lines-km). Rio Tinto completed 24 diamond drill holes for a total of 7,484.15 m.

Norsemont exploration activities between 2007 and 2011 in the district included the mapping of 11,444 hectares at the Constancia Project at several scales, including 1:1,000, 1:2,000 and 1:5,000. Of this, 8,905 hectares were mapped on Hudbay's mining concessions, which represent 39% of Hudbay's mining rights in the area. Additionally, 2,595 rock samples and 41 stream sediments samples were collected during this period. Ground magnetic and IP surveys cover all areas of Hudbay's concessions (magnetic; 281 lines-km and IP 319,6 lines-km).

From 2005 to 2010, Norsemont drilled 153,556m for infill, condemnation, metallurgical, geotechnical, hydrogeological and exploration, drilled by core and reverse circulation.

From 2019 to 2020, additional exploration activities were developed to the northwest of Constancia, between the northwest limit of the pit and the area known as Yanajaja, this after a review and interpretation of the results obtained in the initial explorations, detailed mapping at scale. 1: 2,000 was carried out in 185 Ha and 8,112.30 meters were drilled in 24 holes.

The author is of the opinion that the sampling methodologies and drill frequency is sufficient and that they are representative of the geology and deposits described.

In addition to the Constancia and Pampacancha prospects, other peripheral targets have been identified through prospecting, mapping, grab sampling and geophysics, including conventional IP and Titan 24 (38.4 km covering 8 lines, which was completed in 2011). The following Figure 9-1 shows a general view of the current Constancia district targets.

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FIGURE 9-1: NEAR MINE CONSTANCIA EXPLORATION POTENTIAL

In January 2018, Hudbay announced the entering into of the following agreements to acquire mining properties in southern Peru near its Constancia mine (Figure 9-1): (i) an option agreement with a private Peruvian consortium to earn a 100% interest in the Caballito (formerly Katanga) and Maria Reyna mining properties; and (ii) an agreement to acquire from Panoro Minerals Ltd. 100% of the Kusiorcco mining properties.

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CABALLITO

The Caballito property, located approximately three kilometres northwest of Constancia, is a 120-hectare (297-acre) concession block and is the site of the former Katanga mine, which was operated by Mitsui Mining & Smelting Co., Ltd. and Minera Katanga at different times between the late 1970s and early 1990s. The deposit at Caballito consists of narrow skarn bodies developed in the contact between limestone and monzonite porphyries with copper, silver and gold mineralization in hypogene sulfides. Reliable available data over this area is limited to aeromagnetic and radiometric maps.

MARIA REYNA

The Maria Reyna property, located within ten kilometres of Constancia, is a 5,850-hectare (14,456-acre) concession block. In 2010, diamond drilling by a previous optionee of the Maria Reyna property, intersected copper skarn, breccias and porphyry mineralization. Geophysical surveys and geological mapping have also been conducted on the property and Hudbay believes that the area remains very prospective for additional discoveries.

A summary of the historic drill results from Maria Reyna is contained in Table 9-1, however a qualified person has not independently verified this historical data or the quality assurance and quality control program that was applied during the execution of this drill program for Hudbay and, as such, Hudbay cautions that this information should not be relied upon by investors.

QUEHUINCHA NORTE

Quehuincha Norte target is located 11 kilometers to the north of Constancia Mine and 3 kilometers of the Kusiorcco area, within these block of concessions, was identified by the Hudbay exploration team. This area is coincident with the zone of aeromagnetic and radiometric anomaly identified in 2014.

Detail geological mapping at 1:2,000 scale was complete, 471 rock samples was collected for geochemical analisys, brown-green garnet prograde skarn area was mapping in the contact of porphyritic diorite with limestone of Ferrobamba Formation with an extension of 2 km, the mineralization is copper oxides, chalcopyrite in veinlets and quartz-sulfides structures, this area have active mining workers along the skarn area, an area of 500 x 500 meters of diorite porphyry with intense stock work with quartz and quartz-magnetite veinlets was mapped.

Geophysical data includes 56 kilometers of ground magnetic survey and 26,8 kilometers of IP survey in 13 lines, the chargeability sections show IP anomalies at the west along the skarn outcrop.

All permits have been obtained and the project is ready to drill.

KUSIORCCO

The Kusiorcco property is located within seven kilometres of Constancia and is nearby the Caballito and Maria Reyna properties. It consists of 10 concessions totalling 3,962 ha and hosts a large mineralized dacidic porphyry intrusive showing presence of quartz-sericite alteration, quartz stockwork and a leached cap. Geophysical data includes 24 kilometres of ground magnetic survey and 3 IP lines showing a coincident 1-km by 2-km induced polarization and resistivity anomalies overlying a 300m by 500m alteration zone.

Permitting and community relations work is ongoing to support exploration work on the Caballito, Maria Reyna and Kusiorcco properties.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

TABLE 9-1: HISTORIC DRILL RESULTS FROM THE MARIA REYNA PROPERTY

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10. DRILLING

10.1. OVERVIEW

Extensive drilling has been conducted at Constancia since the early 2000's. The most recent drilling programs were completed by Hudbay, with prior drilling campaigns performed by Rio Tinto and Norsemont Mining (Table 10-1).

TABLE 10-1: DRILLING PROGRAMS BY YEAR (IN METRES DRILLED)

Exploration, infill, metallurgical, geotechnical, hydrogeological, condemnation and twin-hole drilling was conducted in order to adequately assess the deposit (Table 10-2).

TABLE 10-2: DRILLING PROGRAMS BY TYPE

A map showing the location of the drill holes by company is provided in Figure 10-1 along with an outline of the mineral resource pit shell limits for the Constancia and Pampacancha deposits.

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FIGURE 10-1: MAP OF DRILL HOLE LOCATION

10.2. RIO TINTO (2003-2004)

Rio Tinto drilled 24 diamond drill holes for a combined length of 7,483 m, using an average azimuth of 110° and average dip of 70°. The average depth is 312 m and the drill spacing is 300 m. Information for recovery is not available. Collar locations were surveyed with a hand-held GPS and certified by Horizons Consulting of Lima, Peru. Down hole surveying is not available for this campaign.

10.3. NORSEMONT MINING (2005-2010)

Norsemont Mining drilled a total of 529 holes for a combined length of 153,550 m, using an average azimuth of 150° and average dip of 78°. The average depth is 287 m and the drill spacing is 70x70 m. The overall core recovery was recorded as 95% in the original logging however further investigations of core pictures revealed that the recovery was measured on a length basis. Random checks conducted over several thousands of samples have shown that in reality half of the samples had a core recovery inferior to 85% when measured on a volumetric basis. This issue is quite common in the Peruvian porphyry deposits with intense alteration.

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Collar locations were surveyed with a hand-held GPS and certified by Horizons Consulting of Lima, Peru. Down hole surveying was conducted on 30 m intervals (before 2007) and 50 m intervals (after 2007) using a combination of Flexit and Maxibor Survey instruments.

10.4. HUDBAY (2011-2020)

Hudbay acquired Constancia and conducted exploration, infill, metallurgical, geotechnical, hydrologeological, metallurgical and condemnation drilling between 2011 and 2020. A total of 742 drill holes were drilled using an average azimuth of 65° and average dip of 85°. Combining the Rio Tinto, Norsemont and Hudbay drilling, the average drill spacing is 80x80 m. Holes were usually collared with larger HQ-sized core as deeply as possible and finished with NQ-sized core if a reduction in core size was required due to ground conditions. For this drilling, it was also found that actual volumetric core recovery was inferior to the linear core recovery that had been recorded in the database.

In 2017, a twin-hole program of 21 holes was specifically conducted to assess the impact of material loss in previous drilling campaigns. Drilling was conducted with triple tube coring while using lubricants and minimizing fluid pressure. Volumetric core recovery significantly improved in this new campaign compared to the twinned historical Norsemont and Hudbay holes with the proportion of samples with core recovery over 85% increasing from 50 to more than 75%. The findings of the twin-hole study are discussed in more details in the 2017 technical report.

The final collar locations were recorded with a hand-held GPS. All measurements were recorded using UTM coordinates based on the Provisional South America 1956 (PSAD56) datum. Collar locations were surveyed and certified by Hudbay Peru.

Before 2014, the drilling azimuths were set up by marking the front and back sights using a Brunton compass. For the latest drill campaigns, the drilling azimuths were set up with a Differential GPS Trible R8 model 4 with a precision of 3 mm. The inclinations were measured with the drill rig inclinometer (typically, a digital inclinometer). Comparisons between the designed and downhole survey measurements at the base of the casings have shown negligibale variations.

Down hole surveys were measured every 30 m intervals using a hand-held GPS which measured inclination/dip and azimuth direction. For the latest drill campaign, down hole surveys were conducted on 30 m intervals using a Gyro Survey instrument which measured inclination/dip and azimuth direction.

At Pampacancha, down hole surveys were measured every 3 m intervals using a Maxibor 2 instrument which measured inclination/dip and azimuth direction.

There are no factors that could materially impact the accuracy and reliability of the drilling results.

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11. SAMPLE PREPARATION, ANALYSES AND SECURITY

The sampling methodology, analyses and security measures used by the previous owners at Constancia have been documented in the 2018 Technical Report produced by Hudbay and available on SEDAR.

During the Hudbay 2011 campaign, the samples were transported to the SGS laboratory in Lima, Peru for preparation and analysis while in 2014, the samples were transported to the SGS Constancia laboratory at the mine site for preparation. Once samples were pulverized, a 250 g subsample pulp was collected and air freighted to the SGS laboratory in Lima, Peru, for analysis. In 2015, the samples were transported to the SGS Constancia laboratory at the mine site for preparation and analysis and in 2017, for the twin hole drilling program, blanks and standards were inserted at site, prior to dispatching the core boxes to the Certimin laboratory in Lima for sample preparation followed by analysis at the Certimin and SGS laboratories in Lima. Since September 2017, operation of the Constancia laboratory has been transferred to Bureau Veritas (Inspectorate) where all samples collected since have been sent for sample preparation. The pulps were then sent the Bureau Veritas laboratory in Lima for ICP analysis. The SGS laboratory in Lima has been retained for external checks.

As part of Hudbay's QA/QC programs, samples were systematically introduced in the sample stream to assess adequate sub-sampling procedures, potential cross-contamination, precision, and accuracy.

Sample preparation, analysis, and security procedures were reviewed by Olivier Tavchandjian, P. Geo. and Qualified Person of this Technical Report. The sampling methodology, analyses and security measures have been documented in great details already in past published technical reports on the Constancia property and will only be summarised in this document. Unless specified in the report, all the laboratories used for samples prepararation, analysis and security are independent of Hudbay.

11.1. CORE LOGGING

The drilling contractors thoroughly cleaned the drill core retrieved from the core tube before piecing all the segments together in the core boxes. Footage marker were inserted after each run to indicate the relative down-hole depth. Core boxes were labelled with the hole name, box number and from - to footage measurements before being closed with a tightly fitted lid and being delivered to the core processing areas at the Constancia. Private 24-hour per day security guards or Hudbay personnel controlled site access and oversaw sample security at each camp and drill site.

Prior to measuring the core recovery parametres and rock quality data (RQD), visual checks were performed for incorrect placement and orientation of core fragments. Discrepancies caused by mislabelled or misplaced block markers were resolved by consulting the drilling contractors. The drill core was marked with cut lines designed to provide the most representative split.

Standard parametres for core recovery and RQD for each drill run were measured by either the trained geotechnicians or geologists. All core logging was performed by experienced geologists. At the start of each drilling campaign, all geologists were provided with three days of training on the rock types, alterations, mineralization and structures found on the property.

Before 2014, all drill holes were logged on paper logs and later transferred into spreadsheets. Later, drill holes were logged directly on PC tablets. Drill cores are divided into sub-intervals based on the rock types observed by the geologists. Each interval was described for alteration, mineralization, and oxidation state of the primary sulfides. Any significant veins found were also logged along with identifiable structures.

11.2. SAMPLE SELECTION

Core samples for assaying were selected by the core logging geologist. The typical sample interval is 1 to 2 m long in mineralized rocks and 3 m in barren samples while being mindful of lithological contacts. Geologists were responsible for filling paper tags for each sample in the sampling book, with the hole name and sample interval. Sample tag numbers along with the sampled intervals were also entered into the core logging database. For core samples, two of the three tags were stapled into the core box at the starting point of each sample, one to remain there, and the second to accompany the sampled split in the sample bag. Lines were drawn on the core using a permanent marker to indicate the beginning and end of each sample. For QA/QC samples consisting of duplicates (analyses generated from the same interval from a coarse reject or a quarter split), two sets of sample tags were stapled into the core box, and a double line was drawn on the core. For other QA/QC samples (blanks, duplicates and standards), a single sample tag was stapled into the core box indicating the QA/QC sample's relative position in the sequence and the sample type.

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11.3. CORE PHOTOGRAPHS

Logged core boxes with the sample intervals marked and sample tags inserted were photographed using a digital camera mounted to a tripod in natural light. All photos taken were loaded on to a laptop computer and reviewed by the on-site database manager. Any photos deemed unacceptable were retaken.

11.4. CORE SPLITTING

Prior to splitting core, the database manager printed a sample list for each drill hole that included the sample identification number, hole name, sample type and the start and end footage of each sample. This list was used to label sample bags. At the core splitting station, a bucket was lined with the correctly labelled sample bag and the corresponding core box was placed on to the work table next to the core saw. The core cutters separated one of the two sample tags stapled in the core box at the start of the sample and placed them in the sample bag. For quarter split, the second QA/QC sample tag was also placed in the same bag. Core was split along the cut line drawn by the geologist. In gouge and rubble intervals, an aluminium sampling scoop was used to separate the gouge into two halves in the core boxes. Completed sample bags were closed using the bag draw strings and secured at the neck using two zip ties. All saws and sampling buckets were rinsed with water after cutting each sample to prevent cross contamination.

11.5. SAMPLE DISPATCHING

A requisition form that listed the range of sample numbers, job order number, requested analytical codes and any special instructions was created. The requisition form were emailed to the preparation laboratory prior to sample shipment. QA/QC samples including blanks, duplicates and standards were prepared by the database manager prior to sample shipment. On the day of sample shipment, sample bags were cross-checked with the sample requisition list before packing.

11.6. SAMPLE PREPARATION

11.6.1. HISTORICAL

Table 11-1 presents the summary of the sample preparation method used for the drill campaigns conducted by the previous owners.

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TABLE 11-1: SUMMARY OF THE SAMPLE PREPARATION METHOD BY COMPANIES

11.6.2. HUDBAY

Drill core samples were picked up at the core processing facilities and transported via from the camp site to Arequipa and ALS Chemex - SGS laboratories were weighed upon arrival, dried at 100°C, and crushed in jaw crushers to 90% passing through #10 mesh (2 mm). The entire crushed samples were homogenized, riffle split, and 250 g subsamples were pulverized to ≥95% passing through #140 mesh (106 μm) using steel/chrome grinding bowls. Jaw crushers, preparation pans, and grinding bowls were cleaned by brush and compressed air between samples. Cleaning with a quartz wash was conducted between jobs and between highly mineralized samples and, and finally from Arequipa to Lima, Peru. Samples. The remaining coarse and pulps rejects were sent for storage to Abil Corporation warehouse, located in Lima, Peru.

ALS Chemex and SGS have a quality system that is compliant with the International Standards Organization (ISO) 9001 Model for Quality Assurance and ISO/IEC 17025 General Requirements for the Competence of Testing and Calibration Laboratories.

As for the 2017 twin hole program, drill core samples were picked up at the core processing facilities and transported page via Transportes ACOINSA from camp site to Certimin laboratory located in Lima, Peru.

Upon arrival at Certimin laboratory, samples were relabelled, weighed and crushed in jaw crushers to 100% passing through #10 mesh (1.7 mm). The entire crushed samples (about 12 Kg) were homogenized and splitted via rotary splitter (about 500 gr). For every sample, one sub sample was kept while the other was pulverized to ≥85% passing through #200 mesh (74 μm) using Steel/Chrome grinding bowls. Jaw crushers, preparation pans, and grinding bowls were cleaned by brush and compressed air between samples. Cleaning with a quartz wash was conducted between jobs and between highly mineralized samples. The remaining coarse and pulps rejects are stored temporarily at Certimin facilities located in Lima, Peru, and after finishing all processes (metallurgical), will be transported to Constancia mine for final storage.

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Pulp samples prepared at Certimin were dispatched to SGS Lima for ICP-EOS analysis. Upon arrival, samples were weighted and 4% were randomly selected from individual batches for granulometry analysis. If the pulp sample's granulometry failed to achieve ≥95% passing through #140 mesh (106 μm), the samples in the batch were pulverized to ≥85% passing through #140 mesh (106 μm) using steel/chrome grinding bowls. Grinding bowls were cleaned by brush and compressed air between samples. Cleaning with a quartz wash was conducted between jobs. The remaining pulps rejects were stored temporarily at SGS facilities, located in Lima, Peru. Once the assay program was completed, all the pulps reject were transported back to the Constancia mine for final storage.

Certimin (Lima, Peru) has a quality system that is compliant with the International Standards Organization (ISO) 9001 Model for Quality Assurance and ISO/IEC 17025 General Requirements for the Competence of Testing and Calibration Laboratories.

SGS (Lima, Peru) has a quality system that is compliant with the International Standards Organization (ISO) 9001 Model for Quality Assurance and ISO/IEC 17025 General Requirements for the Competence of Testing and Calibration Laboratories.

Since 2018, sample preparation has been conducted at the on-site laboratory in Constancia; one samples arrive to lab facilities,samples are weighed and crushed in jaw crushers to 100% passing through #10 mesh (1.7 mm). The entire crushed samples (about 12 Kg) are homogenized and splitted via rotary splitter (about 500 gr). For every sample, one sub sample is kept for QC purposes while the remaining sample is pulverized to ≥85% passing through #200 mesh (74 μm) using Steel/Chrome grinding bowls. Jaw crushers, preparation pans, and grinding bowls are cleaned by brush and compressed air between samples. Cleaning with a quartz wash is conducted between jobs and between highly mineralized samples. Pulps are delivered to Arequipa via Cruz del Sur (The official personnel transporting contractor).Once in Arequipa, samples are deliver to Lima by air cargo. Once samples are received by Bureau Veriats in Lima, they are assayed following the instructions provided by Mine Geology department (HUDBAY). The remaining coarse and pulps rejects are stored at Mine Geology facilities located in Constancia Mine.

The sample preparation, analysis, and security procedures are considered industry standard, adequate, and acceptable.

11.7. BULK DENSITY

11.7.1. HISTORICAL

At the laboratory, samples were dried at 80°C overnight (12 hours) and then allowed to cool to room temperature. The initial weight of the sample is determined using a top loading balance and recorded. Balances are calibrated using a 0.01 mg, and 0.001 mg calibration weight. The sample is immersed in a pan containing molten paraffin, immediately removed from the molten paraffin and shaken a few times to remove excess wax while hardening. The wax coated sample is re-weighed using the top loader and the weight is recorded. The standard water displacement method^1^ is then used to calculate the specific gravity. In addition and following the same methodology, Norsemont re-analyzed 281 samples at SGS, Lima, Peru as part of its QC program. The results have shown that the SG values are accurate.

^1^ Density = W1 / (V1 - (W2 - W1/Density Wax))

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11.8. ASSAY METHODOLOGY

11.8.1. RIO TINTO

Drill core samples were analyzed for 27 elements by Inductively Coupled Plasma - Atomic Emission Spectroscopy (ICP-AES), with four acids digestion. Selected samples above the detection limit were analyzed by Atomic Absorption Spectroscopy (AAS). All the analysis was performed by ALS, Lima, Peru using the method ME-ICP61 and AA62 for Cu over-ranges.

Sequential copper data is not available. According to an exploration report from Rio Tinto Exploration (2004), limited Cu soluble analysis (CN leach) was performed over selected intervals when chalcocite was described as the main Cu-bearing mineral.

11.8.2. NORSEMONT

Drill core and RC samples were analysed by ALS Chemex for 27 elements using ME-ICP61 and by SGS for 41 elements using ICP40B with four acid digestions. Samples above detection limit were analysed by Cu_AA62 method.

11.8.3. HUDBAY

CONSTANCIA

Drill core samples were analyzed for 41 elements by Inductively Coupled Plasma - Optical Emission Spectroscopy (ICP-OES) with HNO3-HClO4-HF-HCl digestion and HCl Leach. Samples above detection limits were analyzed by AAS, while gold was analyzed by Fire Assay with AAS finish. Samples with copper concentrations above 2,000 ppm were systematically re-assayed by a sequential copper method (AAS73B).

As for the 2017 twin-hole program, drill core samples were analyzed for 41 elements by ICP-OES with HNO3-HClO4-HF-HCl digestion and HCl Leach and by AAS, while gold was analyzed by Fire Assay with AAS finish.

Since 2018, the drill core samples are analyzed for 41 elements by ICP-OES with HNO3-HClO4-HF-HCl digestion and HCl Leach and by AAS, while gold was analyzed by Fire Assay with AAS finish.

PAMPACANCHA

Samples were analysed by the SGS laboratory through Inductively Coupled Plasma - Optical Emission Spectrometry (ICP-OES) after multi-acid digestion and gold was determined by fire assay with Atomic Absorption Spectroscopy (AAS). SGS is currently registered with ISO/IEC 17025 and ISO 9002 accreditation.

Samples were analyzed for 41 elements through ICP-OES (ICP40B Lab Code) after multi-acid digestion of a 0.20 g of sample. If silver, copper, iron, molybdenum, lead and zinc values were above upper reporting limit, then the element was re-assayed using AAS after multi- acid digestion of an 0.25 g of sample. Gold was analyzed by AAS after fire assay using 30 g of sample (FAA313 Lab Code) and values greater than 5000 ppb were re-assayed using a gravimetric finish. All analytical balances are certified annually by a third party. Check weights are used daily to verify calibration of balances. All metal standards used to make the calibration standards for the AAS and ICP are certified and traceable.

Samples with copper concentrations above 2,000 ppm were systematically re-assayed by a sequential copper method (AAS73B).

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11.9. DATA VALIDATION

11.9.1. DRILL HOLE DATABASE

Hudbay built an entirely new drill hole database from all pre-Hudbay drilling and assaying information. The following subsections describe the process Hudbay used to build a completely new database of the drilling and assay values and the steps taken to verify the information. All non-Hudbay drill holes will hereby be referred to as "historical drill holes". The author's opinion is that the data verification is adequate for the purposes used in this Technical Report.

DRILL HOLE COLLARS

Drill hole collar coordinates of pre-Hudbay drilling were resurveyed and reported in a single grid system. The coordinates were converted to WGS 84 UTM Zone 19 using ArcGIS Software. The conversion was based on a best fit transformation using drill hole collars and corners from claim boundaries (approximately 800 points in total). This conversion was verified by plotting the converted coordinates against a drill hole collar compilation map prepared by Horizons South America in 2013, and the results are within an acceptable margin of error (+/- 0.07 m).

DOWNHOLE SURVEYS

Downhole survey files exist for 379 of the 463 historical drill holes. The majority of the downhole surveys were conducted by Geotec and Comprobe using survey methods that measured the drift angle and azimuth. The readings were generally recorded every 50 m for Flexit and every 3m for Maxibor. From the record sheets it cannot be determined if the azimuth recorded was adjusted for magnetic declaration, hence no further adjustments were made to these readings. However, of the 463 historical drill holes, 17 were drilled vertically.

Hudbay field-checked the collar positions and visually verified plans and the database for correctness. Down-hole surveys were checked by examining coarse changes in the variables. Check runs were done at regular intervals to check the consistency of the drilling data.

A random revision from the lithology and mineralization log data and assay certificates from 27 drill holes was performed. Discrepancies were not identified between the log data and assay certificates and the drill hole database used for the mineral resource estimate.

An internal validation of the Pampacancha drillhole database against the original drill logs and assay certificate information was carried out by Hudbay. No significant discrepancies were observed within the database.

DATA SECURITY

The historical assay database and the Hudbay assay database are administered by the database manager with working copies kept on the local drive of a secure computer and backups placed on a secure location on a Hudbay server. Any requests for edits to the database are made to the database manager who updates all the copies. All paper copies of the historical assay certificates and logs are available on the Hudbay's server.

11.10. QUALITY ASSURANCE AND QUALITY CONTROL PROGRAMS

11.10.1. HISTORICAL

Table 11-2 displays the standard reference material used while Table 11-3 presents overviews of the quality control used.

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TABLE 11-2: STANDARD REFERENCE MATERIAL

11.10.2. NORSEMONT

The Quality Control (QC) protocol during the Constancia exploration campaigns from 2006 to 2010 included the insertion of blanks, duplicates and standards using an insertion rate of one of each every 20 samples.

The field duplicates, coarse duplicates, coarse blanks and Certified Reference Materials (CRMs) were inserted on the drill site prior to submission to the laboratory. Furthermore, Norsemont used the additional controls listed below:

Pulp duplicates were routinely analyzed by the laboratory to monitor the precision of the laboratory. They were not submitted as "blind" to the laboratory, so therefore their value is diminished. These were inserted every 20 samples.
Coarse duplicates (or preparation duplicates) were taken after crushing to provide information in regards to the sub-sampling variance. They were not submitted to the same laboratory so therefore, their value is diminished. These duplicates were submitted in a proportion of one in every 20 samples.
Check samples (pulp rejects) were submitted to an umpire laboratory (ACME and Inspectorate). These samples were used to estimate the accuracy of the assay results, along with the standards. These pulps duplicates were submitted in a proportion of one in every 20 samples.

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TABLE 11-3: QUALITY CONTROL BY DRILLING CAMPAIGNS

11.10.3. HUDBAY

CONSTANCIA

Blanks

Blank material consisting of barren quartz certified by INS-1078 (Inspectorate) was introduced into the sampling stream (one every 30 samples) to monitor cross-contamination and samples swaps.

Standards

Two standard reference materials from Ore Research & Exploration (Table 11-4) were introduced into the sampling stream (one every 30 samples) to monitor the accuracy of the assay results.

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TABLE 11-4: ORE RESEARCH AND EXPLORATION STANDARD REFERENCE MATERIALS

Reference material	Au ppm	Ag ppm	Cu ppm	Mo ppm	Zn ppm
OREAS-501	0.204	0.84	2710	59.2	-
OREAS-503	0.687	1.63	5660	390	-

Duplicates

Duplicate samples consisting of one in every 30 samples) were introduced into the sampling stream to monitor the precision of the assay results and to assess the homogeneity of the mineralization.

Check Assay

Since 2018 a rerun of randomly selected samples are assayed in two steps:

Coarse rejects: about 5% of coarse rejected samples were recalled (from Constancia Mine Geology facilities), re-labeled and sent to the on site lab for preparation and delivered to Bureau Veritas Lima Lab for assaying under the same methods of "originals".

Pulp rejects: about 5% of remanent pulps were recalled from Bureau Veritas Lima lab; these samples were sent to SGS Lima, which acted as an Umpire Lab;; once samples arrived to SGS Lima facilities, samples were homogenized in order to avoid any segregation of sample. All samples were assayed following the same method used by the Bureau Veritas Lima lab. No check assays (i.e. pulp reruns) were analyzed through an umpire lab.

PAMPACANCHA

Blanks

Blank material consisting of barren quartz certified by three laboratories (INSPEC, CIMM, and SGS) was introduced into the sampling stream (one every 30 samples) to monitor cross contamination and samples swaps.

Standards

Three matrix match standard references consisting of rock from the Pampacancha deposit and certified by Acme labs, were introduced into the sampling stream (one every 30 samples) to monitor the accuracy of the assay results (Table 11-5).

TABLE 11-5: MATRIX MATCH STANDARD REFERENCE MATERIALS

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Duplicates

Duplicate samples consisting of quartered splits were introduced into the sampling stream (one every 30 samples) to monitor the precision of the assay results and to assess the homogeneity of the mineralization.

In the opinion of the Qualified Person, the sample preparation, security and analytical procedures fully meet best industry practices and the requirements of the CIM guidelines

11.11. SITE VISITS

Hudbay geologists have visited the Constancia and Pampacancha deposits to conduct site inspections to become familiar with conditions on the property, to observe the geology and mineralization, to perform core review and to verify the work completed on the property as part of the mineral resource estimation and technical report process since 2012.

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12. DATA VALIDATION

Data verification was conducted under the supervision of Olivier Tavchandjian, a Qualified Person under the terms of the NI 43-101, and it is the opinion of the author that the quality of the data is suitable for use in resource and reserve calculations and that sampling to date is representative of the deposit. Data validation protocols and results already documented in details in previous technical reports issued by Hudbay for Constancia are summarised in this section together with details on new activities conducted since 2017.

12.1. CONSTANCIA

12.1.1. NORSEMONT QUALITY CONTROL PROTOCOL AND RESULTS

The Quality Control (QC) protocol during the Constancia exploration campaigns from 2005 to 2011 included the following insertions prior to dispatching samples to the laboratories:

Duplicate coarse and pulp samples: one in 20 samples;
CRMs: one in 20 samples; four CRMs were inserted in alternate order; and
Blanks: one in 20 samples.

Duplicate Samples

1,952 duplicate samples obtained by splitting half core samples, were inserted in the samples stream between 2005 and 2011. The sampling procedure is considered to be acceptable if the proportion of failures using the hyperbolic method does not exceed 10% using as a failure limit, the y2=m2x2+b2 hyperbola, evaluated for a 30% relative error1 (RE). The failure rates for Cu, Mo, Zn and Ag ranged between 7% and 16%.

Figure 12-1 represents the duplicate analysis scatter plot. This figure shows that several outliers are evidently occuring from sample mix-up which when excluded provide an acceptable rate of failure. Norsemont also inserted a total of 51 field duplicates from RC holes. The failure rates for Cu, Mo, Zn and Ag ranged between 0% and 8% which indicates that the sampling precision was reasonable.

FIGURE 12-1: CU IN DUPLICATE SAMPLES

Relative error: calculated as the absolute value of the pair difference divided by the pair average and expressed in percentage; also known as Absolute Value of the Relative Difference, or AVRD.

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CRMs

Certified Reference Materials (CRMs) were prepared by SG and ACME Laboratories from coarse rejects from previous drilling campaigns at Constancia. The CRM dataset includes 1,807 assays from 18 different CRMs.

The analytical bias was calculates as follow:

Bias (%) = (AVeo / BV) - 1

Where AVeo represents the average recalculated after the exclusion of the outliers (AV±3*SD), and BV is the Best Value calculated as a result of round-robin tests.

For eleven out of twelve CRMs, biases were between -3.9% and 4.5% with low proportions of outliers. The twelveth CRM shows high biases of 9% and 6% respectively for Cu and Mo.

The author concludes that the Ag accuracy at the ALS Chemex during the 2005 to 2011 campaigns were very good.

Blanks

A total of 2,119 coarse blanks, made from barren material, were inserted in the samples streams. Possible contamination issues were investigated if the blank value exceeded five times the detection limit for the element, and/or if a definite, positive rapport was observed between the blank grade and the grade of the preceding sample.

The sample preparation process at ALS Chemex produced no significant Cu, Mo, Zn, and Ag contamination in most of the blanks evaluated. Only one blank showed failure rates which exceeded 10% for Cu and Mo. However, it is deemed that the failure rate was due to an inadequate sample blank preparation.

Check Samples

388 samples submitted for external control to ACME in 2007, which acted as a secondary laboratory for the 2006 and part of the 2007 campaigns were analyzed.

Reduced-to-Major-Axis (RMA) plots we used to assess the check samples. The RMA method offers an unbiased fit for two sets of pair values (original samples and checks sample) that are considered independent from each other. In this case, the coefficient of determination R2 between the two laboratories is determined, and the bias of the primary laboratory for each element as compared to the secondary laboratory is calculated as:

Bias (%) = 1- RMAS

Where RMAS is the slope of the RMA regression line of the secondary laboratory values versus the primary laboratory values for each element.

The RMA plot indicates good fits for Cu (0.988 R2), Mo (0.974 R2), Zn (0.994 R2) and Ag (0.92 R2) after the exclusion of four outliers for Mo (1.0%), two outliers for Zn (0.5%), and eight outliers for Ag (3.8%). Figure 12-2 represents the results.

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FIGURE 12-2: CU IN CHECK SAMPLES - ACME VS ALS

12.1.2. HUDBAY QUALITY CONTROL PROTOCOL AND RESULTS

The Quality Control (QC) protocol included the insertion of the following control samples in the sample batches:

Duplicate samples: Since 2018, 5% of coarse rejects are prepared and assayed at the same lab using the same methods; also 5% of pulp are recalled and reassayed at umpire lab under the same methodology.
CRMs: one in every 20 samples, were four different CRMs are inserted in alternate order; and
Blanks: two in every 20 samples.

The coarse blanks and Certified Reference Materials (CRMs) are inserted at site prior to submission to the laboratory.

Duplicates Samples in 2015

Hudbay inserted 57 duplicate samples during the drill campaign in 2015. This data was processed using the hyperbolic method.

Figure 12-3 represents the precision evaluation for copper and confirms the absence of bias or any significant issue with sample swap.

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FIGURE 12-3: PRECISION EVALUATION CUT - DUPLICATES SAMPLES

CRMs

Hudbay used Certified Reference Materials (CRMs) prepared by Target Rocks (6 laboratories round robin) using material coarse reject from the mine.

These standards were prepared using porphyry material. To evaluate the accuracy of the CRMs, Hudbay excluded the outliers and calculates the analytical bias.All elements present a bias (Bias (%) = (Mean / Best Value)-1) lower than 10%. Only one standard displayed a higher bias for %Mo.

Blanks

A total of 63 samples, prepared using barren material from the mine, were inserted with each batch submission. Their analysis indicated that the percentage of failures was acceptable for all the elements.

12.1.3. 2017 DRILL PROGRAM

A total of 2,132 core samples were sent for assaying at Certimin Lima (Main laboratory) and SGS Lima laboratories. Atomical Absortion (AA) and ICP methods were used at both laboratories. All the results were imported in to the mine geology database. For resource model purposes, Atomical Abortion results were prioritized over ICP result. Blanks and standards had an insertion rate of 25% and 6% respectively. The coarse duplicates, pulp duplicates, blanks and standards analyzed by Certimin were submitted as "blind" to SGS and to the umpire laboratory (Bureau Veritas Lima).

Certimin

Duplicate samples results have shown that all elements were inside a 20% accepted relative error, implying that the results from the laboratory are precise. Standards presented very low failure rates (mean+ 2SD) for all elements including copper (2.7%), implying that the results from the laboratory are accurate. Performance of the blanks using 10 times the detection limit as performance gates also demonstrated that there were no cross contamination or samples swaps issues at the laboratory.

SGS

All elements were within the performance gates of the standards (mean + 2SD) with low failure rates for copper (<3%), implying that the results from the laboratory are accurate. The performance of the blanks using 10 times the detection limit as a threshold were also excellent, implying no cross contamination or samples swaps at the laboratory.

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Bureau Veritas Umpire Lab

Bureau Veritas was used as an umpire lab in order to ensure that the results from Certimin and SGS were precise and accurate. 10% of the total samples analyzed by Certimin and SGS were randomly selected. Half of the samples were re-dispatched to their original labs and the other half was dispatched to Bureau Veritas for ICP-ES. A total of 30% quality control samples (coarse and pulp duplicates, blanks and standards) were inserted into the samples stream.

Duplicate samples results have shown that all elements were inside of 20% accepted relative error. All elements were within the performance gates of the standards (mean + 2SD) with low failure rates for copper (<3%). The performance of the blanks using 10 times the detection limit also did not show any issue with respect to sample swaps or cross-contamination.

12.1.4. 2019 DRILL PROGRAM

A total of 1,887 core samples were sent for assaying at Bureau Veritas Lima laboratory. Blanks and standards had an insertion rate of 5% and 15% respectively. Samples were prepared at Bureau Veritas Constancia laboratory and sent to Lima for for ICP-ES assay. Finally coarse and pulp rejects were reclaimed and re-assay at Bureau Veritas Lima. Selected pulps we also dispatch to the Umpire lab (SGS Lima). The inserted the blanks and standards analyzed by Bureau Veritas and SGS were submitted as "blind".

Bureau Veritas Lima

All elements were within the performance gates of the standards (mean + 2SD) with low failure rates for copper (<3%), implying that the results from the laboratory are accurate. The performance of the blanks using 10 times the detection limit as a threshold were also excellent implying no cross contamination or samples swaps at the laboratory.

After finish the campaign, 6.35% of coarse rejects were randomly selected, re-labeled and resent to Bureau Veritas Lima Lab for ICP-ES. A total of 10% quality control samples (blanks and standards) were inserted into the samples stream.

SGS Umpire Lab

SGS Lima was used as an umpire lab in order to ensure that the results from Bureau Veritas were precise and accurate. 6.35% of the total samples analyzed by Bureau Veritas were randomly selected. These samples were re-dispatched to SGS Lima Lab. A total of 45% quality control samples (blanks and standards) were inserted into the samples stream.

Duplicate samples results have shown that all elements were inside of 10% accepted relative error. All elements were within the performance gates of the standards (mean + 2SD) with low failure rates for copper (<1%). The performance of the blanks using 10 times the detection limit also did not show any issue with respect to sample swaps or cross-contamination.

12.1.5. 2020 DRILL PROGRAM

This program was executed by the Hudbay Peru exploration team which collected 3,734 core samples (there are 3 last holes not considered in this report); all the samples collected were prepared at Bureau Veritas Constancia onsite laboratory and sent to Lima for for ICP-ES assaying. All the blanks and standards analyzed by Bureau Veritas Lima were submitted "blind". to Bureau Veritas Lima laboratory. Duplicates, Blanks and standards were inserted at a rate of 3.43%, 3.05% and 3.43% respectively.

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Bureau Veritas Lima

All elements were within the performance gates of the standards (mean + 2SD) with low failure rates for copper (<3%); for Ag values are inside of mean +2SD calculated from all samples, implying that the results from the laboratory are accurate. The performance of the blanks using 10 times the detection limit as a threshold were also excellent, implying no cross contamination or samples swaps occurred at the laboratory with an exeption for Zn for which some samples presented some indications of cross contamination. This issue is under investigation by Hudbay but would not have any material impact on the mineral resource and mineral reserve estimates presented in this report.

Duplicate core samples results have shown that all elements were inside of 20% accepted relative error. All elements were within the performance gates of the standards (mean + 2SD) with low failure rates for copper (<5%).

12.1.6. CORE RECOVERY

Prior to the 2017 DDH campaign, core recovery was measured via a linear method and the average core recovery was 95%. Further investigations and random checks of the core box pictures revealed that a significant proportion of these core samples had a volumetric recovery inferior to 85%. During the 2017 twin-hole program, core recovery was measured and compared via the volumetric method. For the 17 pairs of holes, the proportion of 2m composites with a volumetric core recovery higher than 85% increased from less than 50% to more than 75%. Since 2017, great care has been exercised during drillin campaigns to maximise volumetric core recovery.

12.1.7. DRILLING DATA BASE

An internal validation of the drillhole database against the original drill logs and assay certificate information was carried out by Hudbay in 2020. 24 holes (randomly selected) from 492 holes used for the construction of the 2020 resource block model, representing approximately 5% of the total, were randomly selected to be checked against the original laboratory certificates. This validation was conducted for all the elements of importance for mine planning including: Cu, Ag, Au, Mo, Pb, Zn, CuCn and CuSS. From the 4089 samples tested, only 4 samples were found to have different values than in the original certificates representing 0.09% of the total and therefore the database can be considered very reliable. However, a comparison with previous versions of the resource modelling database used between 2014 and 2019 highlighted that precision had been truncated to the second decimal place in the MInesight database in the past. Table 12-1 below presents the differences between the 2020 drillhole database and the 2019 database, confirming a negative bias on gold in the 2019 database and minimum differences for the other metals of economic significance.

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TABLE 12-1: COMPARATIVE SUMMARY OF THE 2019 AND 2020 DRILLHOLE DATABASES (BIAS RESULTS FROM TRUNCATION DURING IMPORTING PROCESS BETWEEN THE DATASHED AND MINESIGHT SOFTWARES)

Based on the work conducted during this review, the 2020 drillhole database used to construct the long-term model LT20-V2 is deemed very reliable and following all NI 43-101 requirements to support resource and reserve estimation and reporting. The correction of a truncation during the importing process has resulted in the elimination of a bias in gold content previously experienced and that had not been detected since at least 2014.

12.2. PAMPACANCHA

12.2.1. QUALITY ASSURANCE / QUALITY CONTROL

Hudbay submitted a total of 15,932 samples from 110 drill holes to the SGS's laboratory for analysis. In addition to these samples, 739 duplicate samples, 471 coarse blanks, 336 reference standards were also submitted. The author's opinion is that the results of the QAQC program at Pampacancha validate the use of the assaying information for resource modeling.

The rate of failure observed on Figure 12-4 is deemed to be related to the fact that quartered core samples were used as duplicates. Given the high variability of the mineralization, it is recommended that future duplicates from core samples be taken from crushed, homogenized and split samples, rather than using quartered split core.

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FIGURE 12-4: PRECISION CU (PPM) - DUPLICATE SAMPLES

Standards

A total of 548 copper standards were submitted as a mean to monitor the accuracy of the laboratory results. Out of these, 113 of the assays fell outside the average +/- two standard deviations. However the bias in all cases is acceptable: 36 failures from 548 in copper (7% of failures), 31 from 548 in silver (6% of failures), 17 from 548 in gold (3%), 18 from 548 in molybdenum (3% of failures) and 11 from 308 in zinc (2% of failures).

Blanks

A total of 720 barren samples prepared from mined rocks were inserted in the submission batches in order to monitor cross-contamination or samples swaps. Results are acceptable for all blanks except for Molybdenum. Failures in Molybdenum indicate a high level of contamination. While this is a concern, it is not considered a material impact to the resultant resource statement.

Drilling Database

An internal validation of the drillhole database against the original drill logs and assay certificate information was carried out by Hudbay. The validation included 100% of the assay values from the Pampacancha drilling. No significant discrepancies existed within the database and it is believed to be accurate and suitable for mineral resource estimation.In the opinion of the Qualified Person, the data validation procedures followed for both the Constancia and Pampacancha database to support the mineral resource and mineral reserve estimates presented in this report fully meet best industry practices and the requirements of the CIM guidelines

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13. MINERAL PROCESSING AND METALLURGICAL TESTING

The Constancia mill is now an operating plant running at steady state and, as a result, several of the initial metallurgical testing and assumptions are no longer relevant. As such, the metallurgical forecasts have been updated to reflect the operating experience and performance of the plant since commissioning and ramp-up.

This section will first briefly describe some of the optimization tests that have been conducted while processing the Constancia ore types to adjust their milling recovery formulae followed by a recap in more details of the metallurgical test work done during the pre-feasibility study for the Pampacancha deposit that is planned to start feeding the mill in 2021Q2.

It is believed that the metallurgical test samples collected are representative of the various types and styles of mineralization found at the Constancia and Pampacancha deposits as a whole".

Table 13-1 summarizes the annual average copper and molybdenum recoveries estimated for the life of the mine between 2021 and 2037 for the combined feed from the two deposits.

TABLE 13-1: COPPER AND MOLYBDENUM RECOVERY

Section 13.1 describes the metallurgical testing and optimization work used to derive the recovery estimates above.

13.1. PLANT TREATMENT CAPACITY

Since 2012, considerable efforts were made by Ausenco and Hudbay to re-evaluate the actual throughput capacity of the processing facilities. The plant has now demonstrated its capacity to process approximately 90 ktpd of ore (31 Mtpa at 94% plant availability) from the Constancia and Pampacancha pit. Its processing capacity is higher than the original Ausenco design based on the following:

(i) While the process design criteria were 76,000 t/d at 91.3% availability, it has been demonstrated that 94% total run time (utililization * availability) can be achieved with good maintenance and operating practice.

(ii) It was thought that the SAG mill throughput would initially be constrained by the hardness of the hypogene ore, however further reviews showed that most of this domain had very low rock quality (RQD). Therefore, SAG milling requires less energy per tonne, thus increasing the overall throughput.

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(iii) The grind size was increased for the hypogene ore, from P80 of 106 to 150 µm. There has been an impact on the copper recovery; however, this will be mitigated by implementing a series of flotation upgrades that together comprise the recovery optimization program described in Section 17.

(iv) There was no extra capital expense required to increase the throughput as it was already within the considered design factors.

Mill throughput predictions for future ores are derived based on a combination of mill sampling campaigns and hardness test work on drill core samples representing the ores that will mined and processed in the future. The is:

1) Perform sampling and mass balance studies around grinding circuit

2) Measure the hardness parameters of the ore processed during the time of the surveys

3) Compare the model-predicted specific energy with measured specific energy

4) Calibrate the model to the plant

5) Apply the calibrated model to future ore hardness parameters to estimate plant capacity

Table 13-2 shows the SAG mill circuit surveys collected at the time of reporting and the calculated correction factors for the SAG mill specific energy model. Table 13-3 shows the ball mill survey rsults to date and the calculated correction factors for the ball mill specific energy model. Test work on future ores is presented in Section 17.

TABLE 13-2: SUMMARY OF SAG MILL SURVEYS AND MODEL CALIBRATION

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TABLE 13-3: SUMMARY OF BALL MILL SURVEYS AND MODEL CALIBRATION

13.2. PLANT RECOVERY

The original Constancia plant feasibility study characterized the metallurgical response of the five ore types-hypogene, supergene, skarn, mixed, and high zinc-in terms of throughput capacity, metallurgical recovery and concentrate quality. Locked cycle testing achieved copper grades in the final concentrate higher than 24%, with relatively low levels of contaminant elements, while the molybdenum concentrate produced grades were over 50% Mo with low contents of copper, lead and iron.

Since commissioning of the plant, various changes have been made to the originally designed operating parameters. For example, the annual throughput capacity was increased from 25.3 milltion to 31.4 million tonnes per year, while the primary grind size was increased from 106 µm to 155 µm (nominal). Changes were also made to the mill liner and lifter configurations, ball sizes, flotation reagents, flocculants, and other parameters as part of the normal optimation and continuous improvement program. To quantify the impact of these changes, numerous sampling campaigns and process audits have been (and continue to be) conducted on both the grinding and the flotation circuits. The audits provide a metric of the plant performance at the time of sampling, and also allow for the calibration of phenomenological models for use in debottlenecking and metallurgical optimization. The calibrated phenomenological models can then be applied against detailed metallurgical test programs undertaken on samples representing the future ores of Constancia and Pampacancha. The results, together with the historical performance of concentrator facility, are used to derive the throughput and recovery models described below.

The following sections describe the procedure. Section 13.2.1. provides a concise description of the grouping criteria applied to the ore zones. Section 13.2.2 describes the empirical recovery models derived for each ore type, and Section 2.2.2 describes the simulation models developed for extrapolation to future ores and plant configurations.

13.2.1. ORE TYPES

The Constancia deposit is a Cu-Mo-Ag porphyry with some of the copper occurring as mineralized skarn. For metallurgical predictive purposes, five ore types were identified based on their dominant metallurgical responses:

Hypogene zone with primary porphyry mineralization. It represents the largest portion of the deposit, including disseminated, quartz veins in stock work and fractures containing chalcopyrite-molybdenite that run below the 3,900 masl. Molybdenum is present as molybdenite and is mostly un-oxidized.
Supergene enrichment zone which occurs lower, made up by secondary ores of digenite copper, covellite, chalcocite (rare native copper) deposited under a leached layer.

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Mineralized skarn zone, smaller in volume, but with higher grades of chalcopyrite, hypogene, rare bornite, galena and sphalerite with mineralization occurring close to the surface. The skarn contains lower levels of molybdenum. The zinc, as sphalerite, concentrates higher in skarn and in the hypogene adjacent to skarn. Lead is present as galena in close relation to sphalerite.
Mixed, or transition zone ore corresponds to the zone where the supergene and hypogene mineralisation mix or co-exist.
High zinc skarn zone, which corresponds to skarn zones with higher total zinc to copper ratios.

13.2.2. EMPIRICAL RECOVERY MODELS

The empirical recovery models are based on plant performance between June 2018 and June 2019. Production data was grouped by period during which ore from each of the ore types was processed and statistical regression was performed to derive equations relating the recovery and concenterate grades to the observed feed grades. The equations were then checked against data through June 2020 and found to be valid. Note that the equations are baseline estimators of plant performance for the current plant configuration and ore characteristics. They have been used to forecast the near-term plant performance when processing Constancia ores; i.e., for 2021 and 2022. The equations are summarized below.

Plant recovery forecasts for alternative plant configurations, including when processing Pampacancha ores, are based simulations as described in Section 13.2.2.

The following abbreviations are used in the equations presented below:

CuT is the total copper grade of the feed, in units of percent
CuSS is the total copper soluble in sulfuric acid grade of the feed, in percent
Zn is the total zinc grade of the feed, in units of percent
Pb is the total lead grade of the feed, in units of percent
Fe is the total iron grade of the feed, in units of percent
P80 is the 80% passing size of the flotation feed, in micrometers

13.2.2.1. HYPOGENE ORES

The following formula describe the copper recovery and concentrate grades:

Cu Recovery (%)=88.73 -0.02*P80 +13.19*(CuT)-30.87*(Zn +Pb)
Cu in Conc (%)=22.84+9.40*CuT-17.48*Zn-0.30*Fe

13.2.2.2. MIXED AND SUPERGENE ORES

The following formula describe the copper recovery and concentrate grades:

Cu Recovery (%)=88.73 -0.02*P80 +13.19*(CuT)-30.87*(Zn +Pb)
Cu in Conc (%)=24.35+5.05*CuT-8.45*Zn-0.42*Fe

13.2.2.3. SKARN AND HI ZINC ORES

The following formula describe the copper recovery and concentrate grades:

Cu Recovery (%)=88.73 -0.02*P80 +13.19*(CuT)-30.87*(Zn +Pb)

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Cu in Conc (%)=23.06+4.60*CuT-5.64*Zn-0.17*Fe

13.2.3. FLOTATION CIRCUIT SIMULATION

Flotation circuit simulations are conducted using state-of-the art phenomenalogical modeling methods. The models are based on mineral-specific flotation kinetics measured in the laboratory and validated with plant sampling campaigns and mass balances. The following plant surveys were used for model validation:

14 Apr 2016, Hypogene mineralization [Aminpro Survey]
16 Apr 2016, Skarn mineralization [Aminpro Survey]
18 Dec 2017, Hypogene mineralization [Aminpro Survey]
18 Jan 2018, Hypogene mineralizaiton [Aminpro Survey]
29 Oct 2020, Hypogene mineralization [Constancia Survey]

The results of the flotation circuit survey work, including mineralogy and flotation test work, model validation, and results are described in the following reports and presentations:

Aminpro Peru S.A.C., 2015, "Optimización de Circuito de Flotación Colectivo." Document P2015-087 Informe Final Rev 0.pdf
Aminpro Peru S.A.C., April 2018, "Constancia Benchmarking." Document P2017-105 Informe Final rev5.docx
Hudbay Corporate Technical Services, Dec 2020, "Constancia Bulk Survey Results and Recommendations." Document Bulk Survey Results.pptx

Note that the flotation circuit simulation studies are used to derive the net effect of a significant change in flotation equipment type or operation, circuit configuration, or ore type. The net effect, or delta, is then overlayed on the empirical recovery model in order to produce the recovery estimated shown in Table 13-1.

13.3. TEST WORK

13.3.1. CONSTANCIA ORE BODY

13.3.1.1. DFS TEST WORK

The original DFS test work was performed prior to 2009. It consisted of extensive laboratory testwork to evaluate metallurgical performance of the Constancia deposit at a DFS level of detail. In addition, a 25t pilot campaign was conducted to provide concentrate for molybdenum flotation evaluation, settling, regrind and filtration tests. The sample collection and testwork were directed by GRD Minproc with assistance from Transmin Metallurgical Consultants (Lima). The results and interpretation of the test program are described at length in the publicly available DFS Technical Report dated September 28, 2009 and for brevity will not be reviewed here.

13.3.1.2. POST-PRODUCTION TEST WORK

Following startup, various test programs were undertaken by both the internal metallurgy and operations team and by external consultants and third-party laboratories. Between 2016 and 2018, two hardness testing programs were commissioned at two different Peruvian laboratories, and approximately 150 to 200 samples were tested (depending on the test type) for SAG Grindability Index (SGI) and Bond Work Index (BWI).

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In 2020 a hardness testing program was completed, with 300 samples representing current and future Constancia ores submitted for SGI and Bond Work Index tests at a Chilean lab. As part of this study, a quality control and assurance program was implemented, in which 15 standard duplicate samples were sent to each of the three labs (the two from the 2016-2018 work, plus the Chilean lab) to ensure that all labs used for prior studies were validated. The results of the 2020 study are shown by the cumulative distribution graphs shown in Figure 13-1, indicate that the Constnacia ores will be progressively harder, particularly after 2023.

FIGURE 13-1: CUMULATIVE DISTRIBUTION OF SGI (LEFT) AND BOND WI (RIGHT), CONSTANCIA

In 2021, a hardness test program was commissioned to improve the understanding of future ore hardness and to further refine the timing of the pebble crushing project (see Section 3.5.1), currently scheduled for implementation in 2023

The most significant flotation study is the recent rougher variability test program performed on samples of future ore from Constancia. The test program was performed in 2020 and early 2021 using the standard Constancia pulp chemistry scheme and new, more selective reagent scheme.

The results, presented below, were used to derive the collection zone kinetics parameters for each test. These were related to head grade and grind size using regression curve fitting techniques. The resulting models were then used to derive the recovery forecast present in Table 13-1using the methodology described in Section 13.2.2

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TABLE 13-4: HEAD GRADES OF VARIABILITY SAMPLES TESTED FOR ROUGHER FLOTATION KINETICS

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TABLE 13-5: ROUGHER RECOVERY OF VARIABILITY SAMPLES

13.3.2. PAMPACANCHA ORE BODY

Pampacancha is a resource that was delineated during the Constancia feasibility study phase but with the metallurgical test work at a pre-feasibility study stage of engineering. It was planned that the Constancia plant, process and reagents would be used to treat the Pampacancha ore. However, the mineralization and rock types at Pampacancha are different and consequently, it was necessary to develop a different regime for this deposit. Rougher Recovery of Variability Samples

13.3.2.1. ORE TYPES

Pampacancha is a skarn deposit, and hence the same broad ore type definitions of the Constancia deposite were applied to Pampacancha. These are:

Skarn Supergene, which constitutes approximately 17% of the deposit
Skarn Mixed, approximately 42% of Pampacancha
Skarn Hypogene, 34% of the deposit, and
Skarn Hi Zinc, the least abundant, approximately 7% of the deposit.

For descriptions, refer to Section 13.2.1.1.

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13.3.2.2. METALLURGICAL TESTS

The metallurgical test programs are divided into three phases, these are:

Phase 1 - Scoping level tests conducted before startup of Constancia
Phase 2 - Prefeasibility level level tests, also conducted prior to Constancia startup.
Phase 3 - Feasibily level tests, conducted during 2020 and early 2021, with some work ongoing.

The results of the test programs are summarized below.

13.3.2.2.1. PHASE 1 - SCOPING TESTS

Preliminary metallurgical tests on the Pampacancha samples were performed by SGS in 2012. The objective of this Phase 1 testwork was to provide comminution and flotation design information for the Pampacancha resource. Phase 1 included mineralogical analysis through SGS Chile and comminution and flotation tests at SGS Peru. Phase 1 involved head assay, specific gravity, natural pH determination, preliminary rougher flotation reagent screening and pH on flotation performance. Locked cycle tests were also conducted using the standard Constancia regime with some optimization tests.

The core samples selected for the Pampacancha metallurgical testwork were fresh quarter core samples collected from diamond drill holes drilled from 2008 to 2011. Samples with skarn and hypogene lithologies were considered representative of the Pampacancha deposit. There was some concern about the age of the samples as oxidation of the samples may impact flotation results. Three master composites for comminution testing and two additional master composites for flotation testing were produced.

Comminution

Preliminary comminution test, SAG Power Index (SPI), Abrasion Index (Ai) and Bond Ball Work Index (BWi), were conducted by SGS Peru, results are shown in Table 13-6

TABLE 13-6: PHASE 1: PAMPACANCHA COMMINUTION RESULTS

Locked Cycle Test

Revised conditions for two composites at a primary grind of P80 of 106 µm, used a collector dosage of 30 g/t A-3302 and 10 g/t A-404 at a pH of 9.5 for a rougher flotation time of 10 min, a regrind P80 of 25 µm and three stages of cleaning at a pH of 11. Locked cycle test results are shown in Table 13-7. Concentrate grade from 19.6 to 21.6% Cu with a recovery from 78.6 to 79.2% Cu were achieved.

TABLE 13-7: PHASE 1: PAMPACANCHA LOCKED CYCLE TEST RESULTS - REVISED CONDITIONS

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Flotation Variability Test

Using the optimal rougher conditions, fifteen samples were floated to determine the effect of the variability on rougher flotation. The head grades ranged from 0.36 to 1.89% Cu. The variability flotation test was a kinetic flotation for 0, 1, 2, 4, 8, 12 and 18 min. The results of the variability test are shown in (Table 13-8). Rougher concentrate recoveries were in the range from 81.2 to 93.5% Cu, 7.85 to 21.1% Fe, 75.8 to 93% Au and 65.9 to 88.4% Ag.

TABLE 13-8: PHASE 1: ROUGHER VARIABILITY TESTING RESULTS

Mineralogy

A PMA QemSCAN, analysis was conducted on the Locked Cycle Test (LCT) head stream for the two composites to determine the mineral assemblage. The copper occurrence in both cases was dominated by chalcopyrite (70-73%), bornite (22-28%) with minor chalcocite (1%) and enargite - tennantite (2%). Table 13-9 reports the QemSCAN modal analysis results.

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TABLE 13-9: PHASE 1: MINERALOGICAL ANALYSIS (MINERAL MASS %)

13.3.2.2.2. PHASE 2 - PREFEASIBILITY TESTS

During 2012, the SGS Laboratory in Lima, Peru conducted a test program using drill core rejects from the 2011 program. The main objective of Phase 2 was to obtain the metallurgical response of mineralized material from the Pampacancha deposit to the Constancia mill flotation conditions. Phase 2 involved head assay, specific gravity, natural pH determination, a locked cycle test to approximate plant metallurgical results and variability flotation conditions. Samples from the skarn and hypogene lithologies were considered representative for the Pampacancha deposit. Three master composites were produced and designated for each sector of the Pampacancha deposit.

Locked Cycle Test

The Constancia project flotation conditions for the locked cycle tests are shown in Table 13-10

TABLE 13-10: PAMPACANCHA LOCKED CYCLE TEST CONDITIONS

Locked cycle test results in Table 13-11 show that a concentrate grade from to 14.9 to 25.3% Cu with a recovery from 67.8 to 80.6% Cu was achieved.

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TABLE 13-11 PAMPACANCHA PHASE 2: LOCKED CYCLE TEST RESULTS

Flotation Variability Test

A variability test was performed on 30 skarn and hypogene samples (Table 13-12). The sample head grades were in the range from 0.21 to 1.29% Cu. The variability flotation test was a kinetic flotation for 0, 1, 2, 4, 8, 12 and 18 min. The results of the variability test including, cumulative recovery and grade until flotation completion (18 min), are shown in Table 13-12. Rougher concentrate recoveries obtained were in the range of 75.3 to 95.7% Cu, 8.6 to 36.5% Fe, 54.7 to 91.8% Au and 54.1 to 95.1% Ag.

TABLE 13-12: PHASE 2: ROUGHER VARIABILITY TESTING RESULTS

13.3.2.2.3. PHASE 3 - FEASIBILITY TESTS

The Pampacancha feasibility study is ongoing at the time of reporting. It consists of

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100 variability samples submitted for hardness tests
10 Rougher flotation kinetics tests conducted on composites of each of the four major ore types
9 Cleaner flotation kinetics tests conducted on a composite of the Pampacancha deposit
40 Rougher variability kinetics tests on samples from each of the four major ore types

The SGI results confirm the Phase 1 comminution results, and show that Pampacancha is moderately softer than Constancia with a mean SGI of approximately 85 minutes.

FIGURE 13-2: DISTRIBUTION OF SAG GRINDABILITY INDEX IN PAMPACANCHA

Rougher flotation kinetics tests were performed under a range of pH, percent solids, and P80 conditions. The results show that the copper recovery is expected to be above 90% at nominal 150-micron grind size, with the exception of HI Zinc, which is only 7% of Pampacancha ores, approximately. Gold recovery is expected to be above 80% for all ore types.

FIGURE 13-3: ROUGHER RECOVERY OF GOLD AND COPPER

Cleaner tests were also conducted on laboratory-produced sample of rougher concentrate floated from a master composite of all ore types. The tests were conducted at a range of typical cleaner percent solids, pH, and regrind size. The results, shown in Table 13-13, indicate expected cleaner recoveries of 99% for copper and 94% for gold, very similar to current plant performance for these elements.

Given the above rougher and cleaner recoveries, there are no major concerns related to the metallurgical performance of the Pampacancha ores; indeed, the fixed recoveries of 85% Cu and 70% Au appear to be conservative in light of these results.

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TABLE 13-13: CLEANER RECOVERY AS A FUNCTION OF % SOLIDS, PH, AND P80

discussion of the Zn/Pb issues we have from time to time based on ore types. If this is not in tech report, consider including.

13.4. CONCENTRATE QUALITY

Although concentrate quality is adequate when treating most of the ores associate with the Constancia deposit, post-start up some ores have demonstrated higher levels of zinc and lead in the concentrates, primarily a result of the feed grade characteristics. These ores constitute a very small percentage of the Constancia deposit and have, to date, been mitigated by a combination of factors, including:

ores that are uneconomical due to higher zinc and lead, but low grades of economic minerals are sent to waste stockpiles, and
ores that are marginally economical due to zinc and lead ratios are sent to stockpiles and, subsequently, fed to the mill in a controlled manner in order to manage the concentrate quality.

The impact of the foregoing on Constancia's operating costs and reserves is considered in the scope of this document. Not considered is the ongoing work aimed at improving the metallurgical deportment of these ores in the process plant, in order to reduce the costs associated with the additional handling and/or the impact on reserves estimation. These include:

use of a more selective primary collector to reduce zinc and lead hydrophobicity in the flotation circuit, currently in Phase 2 plant trials.
Active control of cleaner pulp potential, such as by aeration/oxygenation of the cleaner feed, currently in phase 1 lab trials
use of amine-based depressants in the cleaning circuit, currently in phase 1 lab trials, and
reconfiguration of the column flotation cells to operate in serial configuration instead of parallel configuration, as currently configured, currently in conceptual engineering.
Pampacancha ores do not have significant lead and zinc, but have low to moderate arsenic, although not at levels that would trigger penalties.

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

14. MINERAL RESOURCES ESTIMATES

14.1. CONSTANCIA MINE

Hudbay prepared an update of the Constancia mine resource model using Leapfrog® version 4.5 and MineSight® version 13.10, two industry standard commercial geological and mining softwares. The construction of this 3D resource model and the estimation of mineral resources were performed by Hudbay personnel following Hudbay procedures in compliance with best industry standards and the CIM guidelines. The work was supervised and validated by Olivier Tavchandjian, P.Geo., Vice-President Exploration and Geology of Hudbay.

There are no known factors that could materially affect the mineral resource estimates for the Constancia mine.

MODELING DATABASE

As shown in Table 14-1, 492 drillholes totaling approximately 143,817 m were included in the Constancia database to support the mineral resource estimate. Amongst these, 34 holes were drilled in 2019 and 2020 to define extension of skarn and porphyry mineralization to the North of the Constancia deposit along the Yanak fault.

TABLE 14-1: DRILLING DATA BY YEAR

The drillhole database was exported from MineSight Torque® and provided in Microsoft Excel® format with a cut-off date for mineral resource estimate purposes of October 26th, 2020. In parallel, the drillhole database was also imported in Leapfrog® for additional and independent comparison and validation. No significant differences in the projection of the drillhole traces were found between the two datasets.

MODELING OF THE MINERALIZED DOMAINS

The Constancia deposit trends approximately along an azimuth of N150° with a general dip of 70° to the East. Geologically, Constancia is a Cu-Mo porphyry deposit with small bodies of Cu-Au skarn located at the contact between the porphyry intrusions and the sediments. The Constancia deposit is continuous along a strike length of 3.0 km east-west, 3.0 km north-south and to a vertical depth of approximately 900 m.

Constancia displays the typical porphyry alteration zonation: argillic in the upper part, phyllic and potassic in the centre, with propylitic alteration surrounding the deposit. High grade mineralization is well correlated with the phyllic and potassic alteration zones. Oxidation levels of the deposit have been modelled using the total copper to acid soluble copper (sulphuric and cyanide) ratio following the formulas presented in Table 14-2.

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TABLE 14-2: OXIDATION ZONE FORMULATION

Oxidation Levels	Criteria
Oxide	(CuSS / CuT) > 40%
Supergene	(CuSS + CuCN / CuT) > 70%
Mix	70 ≥ (CuSS + CuCN / CuT) > 30%
Hypogene	(CuSS + CuCN / CuT) ≤ 30%

Wireframes were constructed via 3D explicit modelling in Leapfrog as follow:

Domain 1: a hypogene domain dominated by primary sulphide (mostly chalcopyrite) and including most of the skarn mineralization
Domain 2: a supergene domain dominated by secondary sulphide (mostly chalcocite)
Domain 3: a barren domain that includes sediments, intrusive stocks and intrusive dykes
Domain 4: skarn bodies identified at Constancia North Zone related to Yanak fault
Domain 5: copper oxides and leached zone.

In a final validation step, these domains were loaded into MineSight® in order to ensure proper tagging of the solids to actual drillhole locations. The mineral domains were used as hard boundaries in all cases for grade interpolation purposes. Wireframes of the mineralized domains are shown in Figure 14-1 and Figure 14-2

FIGURE 14-1: 3D VIEW OF CONSTANCIA DOMAINS

Note: Domain 1 in green, Domain 2 in orange and domain 4 in red. Domain 3 and 5 (waste material) are not shown

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Constancia Mine, Cuzco, Peru<br>Form 43-101F1 Technical Report

FIGURE 14-2: CROSS SECTION OF CONSTANCIA DOMAINS

Note: Domain 1 in green, Domain 2 in orange and domain 4 in red. Domains 3 and 5 (waste material) are not shown

CORRECTION FACTORS

As discussed in the previous NI 43-101 technical report, correction factors for copper in the supergene zone were applied based on the 2017 twin-holes study in order to circumvent the apparent low bias in the diamond drillholes. Since then, a significant number of blastholes have been drilled, providing additional information on the low bias affecting the diamond drillholes.

A paired analysis between the blastholes and the drillholes has been performed in 2020 in order to develop a correction factor as close to reality as possible. Figure 14-3 presents the QQ plot of copper in the supergene domain.

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FIGURE 14-3: PAIR ANALYSIS BETWEEN BLASTHOLES AND DRILLHOLES IN SUPERGENE DOMAIN

Based on this analysis, the low bias of copper in the diamond drillholes can be corrected by applying a correction factor of +12% when the copper grade is between 0.4% and 1.3%.

A similar low-grade bias was also identified for lead which constitutes a deleterious elements for the mill. Using the same methodology; the following correction factors were developed to correct the issue:

A correction factor of +8% in ENVLP = 1 when Pb% < 0.05
A correction factor of +7% in ENVLP = 1 when 0.05 ≤ Pb% < 0.5
A correction factor of +8% in ENVLP = 2 when Pb% ≥ 0.1

These corrections were applied to the drillhole database prior to drillhole compositing and grade interpolation.

SPECIFIC GRAVITY

Density values generated from multi regression formulas are based on 1,849 measurements. When geochemical results were not available, baseline values were attributed based on the average value of the domain. The predicted density values were validated through comparisons against the measured density to ensure the predictive models performed adequately.

COMPOSITING

Assay intervals were regularized by compositing drillhole data within each interpreted domain. Although most of the drillholes were assayed on 2m interval, a composite length of 7.5 m was selected as more appropriate to conduct interpolation into the 20m x 20m x 15m block size.

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Compositing over a longer interval than the original assay results in a loss of details and resolution in the downhole grade distribution but does not impact the quality of the block estimates with a 15 m height. Composites with a length lower than 3.75 m were added to the previous interval. Compositing was weighted by specific gravity which has a material impact only for iron, zinc and arsenic grade.

The compositing process was validated by comparing total length, density and length weighted average grade for each metal of the 7.5m composites to the original assays.

EXPLORATORY DATA ANALYSIS

Exploratory data analysis (EDA) includes basic statistical evaluation of the composites for copper (Cu), molybdenum (Mo), silver (Ag), gold (Au), zinc (Zn), lead (Pb), and density (SG). The EDA was conducted separately for each mineralized domain. Within each mineral domain, the EDA also aimed to identify sub-domains that did not support the use of hard contacts for grade interpolation purposes but that justified being used for the purpose of block model validation, smoothing assessment and correction and resource classification. Capping analysis and smoothing assessment were also conducted separately in each domain.

The 7.5 m composite statistics for Cu, Mo, Ag, Au, Zn, Pb and calculated SG are summarized in in the grade estimation validation section. Overall, the statistics of the composites indicate high coefficient of variation and skewness suggesting that the kriging interpolation will be expected to introduce smoothing even in areas of dense drilling coverage and that a thorough smoothing assessment and post-processing of the initial interpolation is warranted to produce unbiased grade-tonnage curves and mineral resource and mineral reserve estimates.

GRADE CAPPING

Capping of gold and silver was applied via the Parrish method due to the skewed nature of the frequency distributions. This method considers capping when the last decile of the population contains more than 40% of the metal and that the last percent contains more than 10% of the metal. Based on this method, gold was capped at 0.8 PPM in ENVLP=1, 0.35 PPM in ENVLP=2 and 1 PPM in the northern skarn, while silver was capped at 18 PPM in ENVLP=1, 25 PPM in ENVLP=2 and 40 PPM in the northern skarn.

VARIOGRAPHY

Down-hole and pairwise directional variograms were created for each individual domain using the MineSight data analysis software. The major, semi-major and minor axis were built in order to fit the attitude of each mineral domain. A linear combination of a nugget and two nested spherical models were adjusted in all cases. Once generated, a systematic visual check was conducted to ensure that the search ellipsoid would be correctly oriented with respect to the geometry of the mineral domains. The variogram parameters modelled by envelope are shown in Table 14-3.

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TABLE 14-3: VARIOGRAM PARAMETERS (SPHERICAL MODELS)

GRADE ESTIMATION AND INTERPOLATION METHODS

The block model consists of regular blocks (20 m along strike by 20 m across strike by 15 m vertically). The block dimensions were selected to match the long-term smallest mining unit (SMU) at Constancia mine. To date, the mine has been using a more selective approach with a SMU size of 10mx10mx15m and this higher level of selectivity is planned to continue until the operation reaches an elevation of 3,930m. The change in SMU size over time is handled through the smoothing assessment/correction step as a post-processing of the kriged estimates. Where a block was intersected by two domains, majority assignment was performed.

Grade interpolation used a strict composite and block matching system based on the domains codes. For example, only composites coded as domain 1 were used to estimate the grade of the block that were assigned to domain 1. Both nearest neighbour (NN) and ordinary kriging (OK) grade interpolations were completed on the uncapped and capped grades, using three passes with increasing minimum information requirements (Table 14-4). The search passes were selected to ensure best local estimates recognizing that OK has a smoothing effect but making no attempt during interpolation to reduce this smoothing as it would negatively impact the quality of the local estimates. Over-smoothing is addressed through the post-processing of the model described in the smoothing assessment sub section of this report.

TABLE 14-4: SEARCH ELLIPSE PARAMETERS

The NN interpolation is solely used to validate that there is no global bias and to assess the level of grade smoothing in the OK interpolation for the selected SMU dimension.

GRADE ESTIMATION VALIDATION

The grade estimation process was validated for each domain to ensure appropriate honouring of the input data and subsequent unbiased resource reporting and use of the model for reserve estimation through the following steps:

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Visual checks of appropriate honouring of the input data but acknowledging that some natural smoothing should occur between samples as the grade of a sample in the middle of a block is not the average grade of the block;
Absence of global bias by comparing the mean grade estimated by kriging to the original composite average grade and to a declustered grade obtained from a nearest neighbour interpolation; and
Assessment of the level of smoothing in the model for the selected selective mining unit (SMU) dimensions considered for resource and reserve reporting.

VISUAL INSPECTION

Visual inspection of block grade versus composited data was systematically conducted in section and plan views. This check confirmed a good reproduction of the data by the model. As an example, a cross-section (looking North) of the copper grade is presented in Figure 14-4.

FIGURE 14-4: VISUAL VALIDATION OF THE COMPOSITES VS THE BLOCK GRADE

GLOBAL BIAS CHECKS

A nearest neighbour interpolation is equivalent to the declustered statistics of the composites based on weighting each composite by its polygon of influence. The average grade obtained from this method is a useful benchmark but not a perfect one as it fails to incorporate the nugget effect measured by the variogram. The higher the nugget effect, the closer the average grade should be to the mean of the composites. Ordinary kriging is in fact the best method of declustering.

A global check was performed to verify that the kriged mean block estimate was located between the mean of the composites and the mean of the nearest neighbour model. Differences between the 7.5 m composites, the NN and OK grades are acceptable in all the sub-domains within each domains. The comparison of the mean and variance for each metal between the 7.5 m composites, the NN and OK models. No significant issues for any of the interpolated variables were observed.

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SMOOTHING ASSESSMENT

The visual validation conducted in section and plan views confirmed that the block grade interpolation was consistent with the supporting composite data. The larger number of composites used for grade estimation in the block model significantly improves the individual block grade estimates but, at the same time results in a much smoother model requiring a careful assessment and in many cases a post-processing of the OK estimates.

The extent of grade 'over-smoothing' in the model was investigated separately by sub domains based on material differences in grade distribution and/or drilling density as illustrated in Figure 14-5. The mean and variance of the kriged estimates were compared to the variance of the composites after declustering (obtained from a nearest neighbour interpolation). The expected true variance between SMUs was calculated from the variogram models.

FIGURE 14-5: EXAMPLE OF SMOOTHING SUBZONES

Over-smoothing is a normal outcome of a sound interpolation method when the drill spacing is not sufficient to address the short-range variability in the metal grade distribution. Smoothing will gradually reduce as additional infill drilling is performed during the definition drilling phases.

In a few cases, the smoothing ratios were high (>3), reflecting the erratic nature of the distribution of the metals in these sub-domains usually characterized by highly skewed histograms and outliers contributing to a large proportion of the total variance.

SMOOTHING CORRECTION

Using the smoothed OK estimates results in an erroneous grade-tonnage curve and reporting resources or reserves at a cut-off grade different than 0% would produce biased estimates, usually over-estimating tonnes and under-estimating grade.

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For some of the elements, the correction did not fully attain the targeted variance reflecting that the log-normal model does not perfectly fit these elements. However, the targeted variance was reached within very close limits in most cases, as illustrated in Table 14-5

TABLE 14-5: SUMMARY OF THE APPLIED SMOOTHING CORRECTIONS FOR COPPER

RECONCILIATION

In order to validate the resource model, tonnes and grade reported at the operating cut-off for the areas that have been mined between July 2017 2018 and October 2020 were compared against the credited mill production. The results presented in Figure 14-6 confirm that the quarterly production is almost always predicted within +/-15% on a quarterly basis both for tonnes, grade and metal content which constitutes the threshold used corporately by Hudbay to report measured resources/proven reserves. Overall the model present no significant bias for neither tonnes, grade or metal content and as such constitutes a reliable to develop a life of mine plan for Constancia mine. The SMU dimensions adopted for this reconciliation work are 10mx10mx15m as per the current operating practice.

FIGURE 14-6: MONTHS ROLLING AVERAGE RECONCILIATION BETWEEN THE RESERVE MODEL AND MILL CREDITED PRODUCTION OF THE MINE COMPARED AT SAME OPERATIONAL CUT-OFF

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CLASSIFICATION OF MINERAL RESOURCE

Classification of mineral resource estimates at Constancia was conducted in full compliance with the 2014 CIM definition standards.

The regression slope values obtained from the kriging of copper grade estimates was used as the primary criteria for resource classification with 80% and 60% regression slope thresholds used respectively to separate measured from indicated and from inferred resources. From detailed reserves to mill reconciliations exercises conducted by Hudbay at its operating mines, including Constancia, this criteria was found to be a reliable first pass measure of quarterly and annual performance in tonnes and grade prediction.

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FIGURE 14-7: RESOURCE CLASSIFICATION (RAW & SMOOTHED)

Note: the top insert presents the block by block resource classification (Measured in green, Indicated in yellow and Inferred on orange) based on the regression slope of copper while the insert at the bottom presents the final classification.

The reconciliation results reported in the previous section from the rolling quarterly estimates confirm the validity of the classification scheme used by Hudbay at Constancia. The error on metal, tonnes and grade is within the +/-15% range on a quarterly basis with a 90% level of confidence as a mining industry benchmark to report measured resources. Measured resources constitute the vast majority of the material mined during the period.

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REASONABLE PROSPECTS OF ECONOMICS EXTRACTION AND MINERAL RESOURCE ESTIMATES

The component of the mineralization within the block model that meets the requirements for reasonable prospects of economic extraction was based on the application of the Lerchs-Grossman (LG) algorithm. The mineral resources are therefore contained within a computer generated open pit geometry.

Mineral Resource Statement

Mineral resource estimates are reported exclusive of reserve estimates for Constancia and include the inferred mineral resource estimates located inside the pit design used to report mineral reserve estimates as well as the measured, indicated and inferred mineral resource estimates located outside of the mineral reserve pit and inside the mineral resource pit shell. Resource estimates are reported by applying the same NSR formula as the one used for reserve reporting and described in Section 15 of this Technical Report. This NSR formula models the combined benefit of producing copper, molybdenum and silver in addition to mine operating, processing and off-site costs.

The mineral resources, classified as Measured, Indicated and Inferred, are summarized in Table 14-6.

TABLE 14-6: CONSTANCIA MINERAL RESOURCE ESTIMATES EXCLUSIVE OF MINERAL RESERVES

Notes:

Totals may not add up correctly due to rounding.
Mineral resources are exclusive of mineral reserves and do not have demonstrated economic viability.
Mineral resource estimates do not include factors for mining recovery or dilution.
Metal prices of $3.10 per pound copper, $11.00 per pound molybdenum, $1,500 per ounce gold, and $18.00 per ounce silver were used to estimate mineral resources.
Constancia mineral resources are estimated using a minimum NSR cut-off of $6.14 per tonne and assuming metallurgical recoveries (applied by ore type) of 85.8% on average for the life of mine.
Mineral resources are based on resource pit designs containing measured, indicated, and inferred mineral resources.

14.2. PAMPACANCHA

The following summarize the method used for Pampacancha mineral resource model developed by Hudbay's in 2017. There have been no changes to the resource model of Pampacancha since 2017.

There are no known factors that could materially affect the mineral resource estimates for the Pampacancha deposit.

MODELING DATABASE

There have been no changes to the Pampacancha resource model since the 2018 Technical Report, which is filed on SEDAR.

140 drill holes totalling approximately 38,240 metres were used for the construction of the resource model. All drill holes have full chemistry obtained from ICP analysis while approximately 14% were also analysed for sequential copper, i.e. sulphuric (CuSS) and cyanide (CuCN) soluble copper.

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MODELING OF THE MINERALIZED DOMAINS

The skarn mineralization is located at the contact between a limestone, a semi-circle shaped diorite and a monzonite porphyry (MP2) and strikes approximately N160. Porphyry diorite dikes are interpreted as the source of the copper mineralization. A smooth and continuous envelope of the skarn mineralization generally grading above 0.1% Cu was used for grade estimation purposes. The diorite and limestone interpreted bodies were used as a guide to limit the skarn mineral envelope on all sides (Figure 14-8). A high-grade skarn envelope was also modelled but was used only for the smoothing correction only.

FIGURE 14-8: SIMPLIFIED GEOLOGICAL FRAMEWORK FOR THE 2017 LONG TERM MODEL

SPECIFIC GRAVITY

Specific Gravity (SG) was measured for 633 samples or approximately 2% of the total sampled length. A stochiometric formula used to assign SG to the samples 2013 resource model was checked and validated against the actual SG measurements in the skarn mineralized domain, which is the only zone hosting significant mineralization. The SG calculated from the stochiometric formula presents a very strong linear correlation with %Fe and a simple linear regression from the iron grade, which can be easily be used to ensure a SG value was assigned to all the samples located within the mineralized envelope used for grade interpolation.

COMPOSITING

In order to normalize the weight of influence for each sample, assay intervals were regularized by compositing the drill hole data into equal length intervals using domain boundaries to break composites. The compositing length was set to 7.5m when considering the envisaged mining block height of 15 m. Compositing was weighted by specific gravity which has a material impact for most of the elements in the skarn mineralized envelope. Weighting by specific gravity has a very significant impact on the Cu, Fe and As grade with a reduced impact on most of the other elements.

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GRADE CAPPING

Capping was considered due to the skewed nature of the frequency distributions. The approach to define an appropriate amount of metal at risk was based on Constancia, on defining a population of outliers as the highest grading of 1% for each metal. For these outliers, the mean and median values were compared and the natural breaks in the cumulative histogram of the outliers were checked to identify natural slope changes. For all the elements, the outliers represent less than 10% of the total metal content and the metal at risk was deemed low and no capping was applied for the modeling of Pampacancha.

VARIOGRAPHY

Pairwise relative variograms were calculated for each element within the mineral envelope. A linear combination of a nugget effect and two exponential (spherical, in the case of Fe) structures were fitted.

GRADE ESTIMATION AND INTERPOLATION METHODS

A block model prototype with cells 20 m x 20 m x 15 m was created in order to entirely fill the mineralized domain. The block size was chosen such that geological contacts are reasonably well reflected and to support a large-scale mining scenario while taking into consideration the DDH spacing. The interpolation plan was completed on the composites, 7.5 m in length, via ordinary kriging (OK). All the OK grade estimates were weighted by specific gravity. Grade estimation used a composite and block matching scheme based on the envelope code which is treated as a hard boundary for grade interpolation purposes.

The interpolation process used a nested search strategy with three passes of increasing ellipsoid radius. The composite selection parametres for grade estimation in each domain (minimum, maximum and, maximum number of composites per hole) were selected to minimize bias.

This first interpolation pass is restricted to a minimum of 16 composites and a maximum of 32 composites (without a maximum of six composites per hole) with quadrant declustering. For the blocks that did not meet this criteria, a second interpolation pass is restricted to a minimum of 16 composites and a maximum of 32 composites (with a maximum of six composites per hole) and quadrant declustering but with a search radius increased by 50% in all directions. Finally, in a third pass, the blocks not yet estimated were interpolated inside a search ellipsoid with a radius doubled from the second search (3 x the initial search) with a minimum of one composite and a maximum of 32 composites (with a maximum of six composites per hole) without quadrant declustering.

GRADE ESTIMATION VALIDATION

The final block model was validated to ensure:

appropriate honouring of the input data but acknowledging that some natural smoothing should occur between samples and that the grade of a sample in the middle of a block is not the average grade of the block;
absence of global bias by comparing the mean grade estimated by kriging to the original composite average grade and a declustered grade obtained from a nearest neighbour interpolation; and
assessment of the level of smoothing in the model for the selected Selective Mining Unit (SMU) dimensions considered for reserve reporting.

VISUAL INSPECTION

Systematic visual inspection of block grade versus composited data was conducted in section and plan view. This check confirmed a good reproduction of the data by the model.

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GLOBAL BIAS CHECKS

This validation step consists of comparing the global average grade of each between the original composites, the kriged block estimates and the nearest neighbour estimates.

Global checks for each element revealed no significant issues for any of the interpolated variables. The impact of declustering has a significant impact on the mean grade for most of the economic metals associated with the sulphide mineralization and also for CuSS and CuCN. This is a fairly common result as higher grade zones tend to be over sampled compared to lower grade zones.

Smoothing Assessment and correction

The extent of grade 'over-smoothing' in the model was investigated separately in two separate zones in order to take into consideration domains with material differences in grade distribution. The mean and variance of the kriged estimates were compared separately in each zone with the variance of the composites after declustering. This exercise was repeated for all the elements used for mine planning including Cu, Au, Ag, Mo, Fe, Zn, Pb, As, CuSS and CuCN.

An indirect log-normal correction was used to perform a change of support on the kriged model in order to obtain unbiased grade tonnage curves within the future pit. For some of the elements, the correction did not fully attain the targeted variance reflecting that the log-normal model does not perfectly fit these elements. However, for Cu which is the main element of economic interest, the targeted variance was reached within a very close limit

CLASSIFICATION OF MINERAL RESOURCE

Similarly to Constancia, the regression slope values obtained from the kriging of the copper grade estimates were used as a basis for resource classification. 90% and 80% regression slope thresholds were used respectively to separate measured from indicated and indicated from inferred resources. Since no mining has yet occurred at Pampacancha, the criteria for defining measured and indicated resources is higher than at Constancia which used 80% and 60% thresholds respectively on the kriging regression slope. In a second pass, a smoothing algorithm was applied to remove isolated blocks of measured within areas of mostly indicated category or isolated indicated blocks within areas of mostly measured category blocks. Proportions of measured and indicated category blocks were not changed significantly by this smoothing process.

REASONABLE PROSPECTS OF ECONOMICS EXTRACTION AND MINERAL RESOURCE ESTIMATES

The component of the mineralization within the block model that meets the requirements for reasonable prospects of economic extraction was based on the application of the Lerchs-Grossman (LG) cone pit algorithm. The mineral resources are therefore contained within a computer generated open pit geometry.

1) The following assumptions were applied to the determination of the mineral resources:

2) Economic benefit was applied to measured, indicated and inferred classified material within the resource cone.

3) A constant 45 degree pit slope was used for the resource estimate.

4) A haulage increment or bench discounting was applied to the blocks above (+ $0.004) and below (+ $0.01) RL 4305.

5) The resource estimate was not limited by any property or permit constraints.

The cost and recovery assumptions summarized in Table 14-7 and Table 14-8, and are specific for Pampacancha. The reporting of the mineral resource by Net Smelter Return (NSR) within the LG pit shell reflect the combined benefit of producing copper, molybdenum, silver and gold as per the following equations based on ore grades, recoveries, grade of concentrate, payable metals, deductions and royalties, in addition to mine operating and processing costs.

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TABLE 14-7: COST INPUTS FOR PIT SHELL CONSTRUCTION

TABLE 14-8: RECOVERY VS MATERIAL ASSUMED FOR PIT SHELL CONSTRUCTION

14.3. MINERAL RESOURCE STATEMENT

Mineral resources for the Pampacancha deposit were classified under the 2014 CIM Definition Standards for Mineral Resources and Mineral Reserves2 by application of a NSR that reflects the combined benefit of producing copper, molybdenum and silver in addition to mine operating, processing and off-site costs.

The mineral resources, classified as Measured, Indicated and Inferred are summarized in Table 14-9. The QP for the mineral resource estimate is Olivier Tavchandjian, P. Geo, Vice-President Exploration and Geology of Hudbay. The mineral resources are reported using the long-term metal price assumptions and have an effective date of January 1st, 2021.

TABLE 14-9: PAMPACANCHA MINERAL RESOURCE ESTIMATES EXCLUSIVE OF MINERAL RESERVES

Notes:

Totals may not add up correctly due to rounding.
Mineral resources are exclusive of mineral reserves and do not have demonstrated economic viability.
Mineral resource estimates do not include factors for mining recovery or dilution.
Metal prices of $3.10 per pound copper, $11.00 per pound molybdenum, $1,500 per ounce gold, and $18.00 per ounce silver were used to estimate mineral resources.

^2^ Ontario Securities Commission web site (http://www.osc.gov.on.ca/en/15019.htm)

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Mineral resources are estimated using a minimum NSR cut-off of $6.14 per tonne and assuming metallurgical recoveries (applied by ore type) of 85.8% on average for the life of mine.
Mineral resources are based on resource pit designs containing measured, indicated, and inferred mineral resources.

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15. MINERAL RESERVES

The Mineral Reserves are contained within two open-pit deposits (Constancia and Pampacancha) and, to a minimum extent, surface stockpiles. Proven and Probable Mineral Reserves for the two deposits as of December 31st, 2020 are estimated to be 532.5 Mt grading 0.31% Cu with a 17 year mine life. A complete mine plan has been developed to update the mineral reserve estimates starting with the resource models described in Section 14. Detail pit designs were optimised based on the application of Net Smelter Return values incorporating mining, processing, and economic parameters.

These Mineral Reserve estimates are reported following NI 43-101 and the classifications adopted by CIM (2014). NI 43-101 defines a Mineral Reserve as "the economically mineable part of a Measured and/or Indicated Mineral Resource".

There are no other mining, metallurgy, permiting or other relevant factors that could materially affect the mineral reserve estimates presented in this section.

15.1. PIT OPTIMIZATION

Pit optimization of multi-element revenues at Constancia and Pampacancha was performed and nine pit stages were selected for the Constancia pit (San José is the third Stage) and two pit stages for the Pampacancha pit.

RESOURCE MODELS

The resource models used for the Mineral Reserve Estimation are described in details in Section 14, with a Selective Mining Unit (SMU) ranging from 10x10x15 meters in the upper part of the deposit to 20x20x15 meters below 3,930 m where the mine plans to operate in a less selective manner due to the more consistent and uniform nature of the mineralization.

The approach used to construct the resource model describes in section 14 ensures all necessary internal and external dilution has been already included in the grade estimation process within each SMU. There is not additional mining dilution or mining recovery factor applied when the mineral resource estimates are included in the life of mine plan and production profile to be converted to mineral reserve estimates. These assumptions have been validated through on-going mineral reserve estimates to mill credit reconciliation.

NET SMELTER RETURN

Mine Design and Reserve estimation for Constancia and Pampacancha pits used an NSR value calculation stored in each block in the resource models. The NSR calculation takes into account the Cu, Mo, Ag, and Au grades, deleterious element concentration such as Zn, Pb, Fe, and Cu oxide content and mill recoveries, contained metal in concentrate, deductions, and payable metal values, metal prices, freight costs, smelting, refining and royalty charges. The following points detail the NSR calculation:

In-situ NSR is the net value of metals (in dollars) contained in a concentrate produced from an ore block after smelting and refining. Using the concentrator recovery of the metals into the concentrate and the grade of the concentrate produced the mass pull of each block in the resource model expressed in terms of tonnes of concentrate per tonne of ore processed is first estimated.
The value of the payable metals in the concentrate is calculated based on copper, gold, silver and molybdenum metal content in the concentrate subject to deductions with smelters and roasters.
From the value of the payable metals, the selling costs, which include the marketing costs, transport costs, port charges, insurance costs, shipping costs, and smelting charges expressed in $/dmt concentrate and other deductions like the refining charges and price participation defined in $/payable metal are taken out as well as applicable royalties.

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The in-situ NSR of each block in the resource model is the sum of the in-situ NSR value from the copper concentrate and the molybdenum concentrate to be sold.
Only measured and indicated resource blocks with NSR values greater than their processing costs are considered to be potential ore. Inferred resources are reported as waste.
Copper, silver, gold, molybdenum, zinc, and lead grades in the copper concentrate and in the molybdenum concentrate are estimated based on the test work detailed in Section 13.

The operating cost, and metal recoveries used for pit optimization are showed in Table 15-1 and Table 15-2.

TABLE 15-1: OPERATING COST

TABLE 15-2: METAL RECOVERIES

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OPTIMIZATION RESULTS

The Whittle software was used to create incremental economic pit-shells using the Lerchs-Grossman (LG) algorithm to calculate the optimum pit limit. This optimization relies on the block's NSR value to find each mining block that can be mined with a profit. The optimum economic shell generated by Whittle is then used as a guide for mine design.

Pit shell for Constancia was generated at a 0.73 revenue factor and contained approximately 473 Mt of Ore and 432 Mt of Waste. The pit shell captures close to 99.0% of the Net Cash flow of the 1.00 revenue factor pit shell.

Pit shell for Pampacancha was generated at a 0.85 revenue factor and contained approximately 42 Mt of ore and 75 Mt of waste. The pit shell captures about 100% of the net cash flow of the 1.00 revenue factor pit shell.

PIT DESIGN CRITERIA

The pit design parameters for Constancia and Pampacancha are summarised in Table 15-3, and the final pit extents are illustrated on Figure 15-1 and Figure 15-2.

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TABLE 15-3: PIT DESIGN PARAMETERS

FIGURE 15-1: CONSTANCIA ULTIMATE PIT DESIGN

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FIGURE 15-2: PAMPACANCHA ULTIMATE PIT DESIGN

15.2. MINERAL RESERVES

The mine plan is based on the capacity of the process plant, which in turn relies on the grinding circuit throughput.

The plant has the capacity to process 31 Mtpy (90ktpd at 94% availability), with a mining rate of 81.0 Mtpy as maximum (ore plus waste). Table 15-4 presents Constancia and Pampacancha Mineral Reserves at December 31st, 2020.

TABLE 15-4: CONSTANCIA AND PAMPACANCHA MINERAL RESERVE ESTIMATES AS AT JANUARY 1, 2021

Notes:

Totals may not add up correctly due to rounding.
Metal prices of $3.10 per pound copper, $11.00 per pound molybdenum, $1,500 per ounce gold, and $18.00 per ounce silver were used to estimate mineral reserves.
Mineral reserves are estimated using a minimum NSR cut-off of $6.14 per tonne and assuming metallurgical recoveries (applied by ore type) of 85.8% on average for the life of mine.

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A comparison of the December 31st, 2019, and December 31st, 2020 Mineral Reserve estimate is summarised in Table 15-5 for the Constancia pit for the Pampacancha pit. The December 31st, 2020 Constancia Reserve estimate shows an increase of 10% in the copper contained after subtracting the 2020 mining depletion in the Mineral Reserves estimates at Constancia. The increase in tonnes and metal content at Constancia results from the extension of the deposit to the North through successful exploration efforts, improvements in the resource modeling process and a re-optimization of the pit design and mining sequence. There has been no change between 2019 and 2020 the Pampacancha reserve estimates.

TABLE 15-5: COMPARISON OF THE MINERAL RESERVE ESTIMATES AT CONSTANCIA AS AT DECEMBER 31, 2019 AND AS AT JANUARY 1, 2021

Notes:

Totals may not add up correctly due to rounding.
Re-evaluation of economic viability.
Mineral resources are exclusive of mineral reserves and do not have demonstrated economic viability.
Metal prices of $3.10 per pound copper, $11.00 per pound molybdenum, $1,500 per ounce gold, and $18.00 per ounce silver were used to estimate mineral reserves and resources.
Mineral reserves and resources are estimated using a minimum NSR cut-off of US$6.14 per tonne.
Metallurgical recoveries are applied by ore type and assumed to be 85.8% on average for the life of mine.

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16. MINING METHODS

16.1. MINE OVERVIEW

The Constancia mine is an open pit mining operation relying on conventional trucks and shovels. The Constancia ultimate pit design will measure approximately 1.6 km east to west, 1.7 km north to south, and have a maximum depth of around 705 m. The Pampacancha ultimate pit design will measure approximately 0.6 km east to west, 1 km north to south, and have a maximum depth of about 300 m. A primary waste rock facility (WRF), which is located to the south and east of the Constancia pit, is intended to be used for both deposits.

The processing facility is located approximately 1 km west of the Constancia Pit. The NAG waste rock is deposited south of the Constancia pit, while the tailings management facility (TMF) is located 3.5 km southwest of the Constancia pit. The general layout of the Constancia Operations site is shown in Figure 16-1

FIGURE 16-1: CONTSTANCIA SITE LAYOUT

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16.2. MINE PHASES

16.2.1. DESIGN CRITERIA

Final pit limit designs have been created for Constancia and Pampacancha based on the optimized pit shells. Constancia is mined in nine stages and a two stage pit is planned at Pampacancha. The minimum mining width for each phase is 60 meters; this will allow a shovel, trucks (in two lines) and a drill to work safely and simultaneously.

The haul roads at Constancia and Pampacancha have been designed with a 10% grade and 32 m width for double lane haul roads and 24 m for single lane haul roads. Figure 16-2 presents the ultimate pit footprint for Constancia and Pampacancha.

FIGURE 16-2: CONSTANCIA AND PAMPACANCHA ULTIMATE PIT DESIGN

PIT SLOPE GUIDANCE

The pit slope angles used for pit optimization for Constancia and Pampacancha are based on engineering studies conducted at pre-feasibility or feasibility levels of engineering. The pit slope design was updated by Golder in December 2019 for Constancia and designed by TWP/Itasca in August 2013 for Pampacancha. The following figures and tables present the design parameters for Constancia (Figure 16-3, Table 16-1, Figure 16-4 and Table 16-2) and Pampacancha (Figure 16-5 and Table 16-3).

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FIGURE 16-3: CONSTANCIA PIT STRUCTURAL DOMAINS

TABLE 16-1: RECOMMENDED PIT SLOPE ANGLES BY STRUCTURAL DOMAINS

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FIGURE 16-4: CONSTANCIA PIT SLOPE SECTORS

TABLE 16-2: CONSTANCIA, RECOMMENDED PIT SLOPE ANGLES

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FIGURE 16-5: PAMPACANCHA PIT SLOPE SECTORS

TABLE 16-3: PAMPACANCHA, RECOMMENDED PIT SLOPE ANGLES

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16.2.2. MINE PHASES AND ULTIMATE PIT

The extraction sequence is defined by nine mining phases at Constancia and two mining phases at Pampacancha. In parallel, seven pushbacks are planned for ore exposure purposes and optimal sequencing at Constancia with one pushback for waste and construction purposes. The phase development strategy consists of extracting the highest metal grades along with minimum strip ratios during the initial years to maximize the economic benefits while enabling smooth transitions in waste-stripping throughout the life of the mine to ensure sufficient ore exposure. The NSR cut-off grade to report Mineral Reserves has been defined at a 6.14 $/t cut-off for both Constancia and Pampacancha. Marginal material is treated as waste. The extraction sequence by phase is illustrated for Constancia on Figure 16-6 and for Pampacancha on Figure 16-7.

FIGURE 16-6: CONSTANCIA PHASE DESIGN, SECTION VIEW SN

FIGURE 16-7: PAMPACANCHA PHASE DESIGN, SECTION VIEW NS

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CONSTANCIA - MINE PHASE 1

Phase 1 of mining has been completed in 2018 and is not discussed in this Technical Report as it does not impact the future sequence of extraction.

CONSTANCIA - MINE PHASE 2

Phase 2, mined since 2014, corresponds approximately to the Lerchs-Grossman pit shell defined by a $0.60/lb Cu price (20% of base metal price sensitivity case). This pit is located about 2,000 meters east of the primary crusher and ranges in elevation from 4,440 to 4,035 masl. The phases were approximately 1,100 meters wide east-west and 1,170 meters north-south. This phase is actually being completed. Remainings of Phase 2 will produce around 1.1 million tonnes of ore at a stripping ratio of 0.38:1 (tonnes of waste per tonnes of total ore). Figure 16-8 presents Phase 2.

FIGURE 16-8: CONSTANCIA PIT PHASE 2

CONSTANCIA - MINE PHASE 3

Phase 3 of mining has been completed in 2016 and is not discussed in this Technical Report as it does not impact the future sequence of extraction.

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CONSTANCIA - MINE PHASE 4

Phase 4, mined since 2018, corresponds approximately to the Lerchs-Grossman pit shell defined by a $1.20/lb Cu price (40% of base metal price sensitivity case). This pit ranges from 4,410 to 3,990 masl in elevation. The phase is approximately 1,400 meters wide east-west and 1,250 meters north-south. Phase 4 will produce approximately 53.8 million tonnes of ore at a stripping ratio of 0.52:1 (tonnes waste per tonnes of total ore). An illustration of the Phase 4 pit is shown in Figure 16-9.

FIGURE 16-9: CONSTANCIA PIT PHASE 4

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CONSTANCIA - MINE PHASE 5

Mining in Phase 5 starts in 2020 and will correspond approximately to the Lerchs-Grossman pit shell defined by a $1.38/lb Cu price (46% of base metal price sensitivity case). This pit will range from 4,440 to 3,885 masl elevation. The phase is approximately 1,600 meters wide east-west and 1,300 meters north-south. Phase 5 will produce about 140.3 million tonnes of ore at a stripping ratio of 0.68:1 (tonnes waste per tonnes of total ore). An illustration of the Phase 5 pit is displayed in Figure 16-10.

FIGURE 16-10: CONSTANCIA PIT PHASE 5

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CONSTANCIA - MINE PHASE 6

Mining in Phase 6 will start in 2025 and will correspond approximately to the Lerchs-Grossman pit shell defined by a $1.68/lb Cu price (56% of base metal price sensitivity case). This pit will range from 4,320 to 3,825 masl in elevation. The phase is approximately 1,650 meters wide east-west and 1,700 meters north-south. Phase 6 will produce about 59.2 million tonnes of ore at a stripping ratio of 1.66:1 (tonnes waste per tonnes of total ore). An illustration of the Phase 6 pit is displayed in Figure 16-11.

FIGURE 16-11: CONSTANCIA PIT PHASE 6

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CONSTANCIA - MINE PHASE 7

Phase 7 will finish mining in 2021. The waste material from this phase will be primarily used for the Tailing Management Facility (TMF). The phase is approximately 800 meters wide east-west and 400 meters north-south. Remainings of Phase 7 will produce approximately 0.8 million tonnes of waste. An illustration of the Phase 7 pit is displayed in Figure 16-12. The pit entrance is at 4,230 masl.

FIGURE 16-12: CONSTANCIA PIT PHASE 7

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CONSTANCIA - MINE PHASE 8

Mining in Phase 8 will start in 2027 and will correspond approximately to the Lerchs-Grossman pit shell defined by a $1.92/lb Cu price (64% of base metal price sensitivity case). This pit will range in elevation from 4,410 to 3780 masl. The phase is approximately 1,300 meters wide east-west and 1,300 meters north-south. Phase 8 will generate about 50.7 million tonnes of ore at a stripping ratio of 2.24:1 (tonnes waste per tonnes of total ore). An illustration of the Phase 8 pit is presented in Figure 16-13.

FIGURE 16-13: CONSTANCIA PIT PHASE 8

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CONSTANCIA - MINE PHASE 9

Mining the final pushback, Phase 9 will start in 2028 and will correspond approximately to the Lerchs-Grossman pit shell defined by a $2.19/lb Cu price (73% of base metal price sensitivity case). This pit elevation will range from 4,440 to 3,735 masl. The phase is approximately 1,500 meters wide east-west and 1,400 meters north-south. Phase 9 will produce about 180.6 million tonnes of ore at a stripping ratio of 0.84:1 (tonnes waste per tonnes of total ore). An illustration of the Phase 9 pit is shown in Figure 16-14.

FIGURE 16-14: CONSTANCIA PIT PHASE 9

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PAMPACANCHA - MINE PHASE 1

Mining in Phase 1 will start in 2021 and correspond approximately to the Lerchs-Grossman pit shell defined by a $0.24/lb Cu price (24% of base metal price sensitivity case). This pit is located about 5,700 meters east of the primary crusher, and the elevation ranges from 4,320 to 4,080 masl. The phase is approximately 570 meters wide east-west and 800 meters north-south. Phase PC01 will produce about 23.4 million tonnes of ore at a stripping ratio of 1.71:1 (tonnes waste per tonnes of total ore). An illustration of the Phase 1 pit is shown in Figure 16-15.

FIGURE 16-15: PAMPACANCHA PIT PHASE 1

PAMPACANCHA - MINE PHASE 2

The mining of the final pushback during Phase 2 will start in 2023 and correspond approximately to the Lerchs-Grossman pit shell defined by a $2.55/lb Cu price (85% of base metal price sensitivity case). This phase will range in elevation from 4,350 to 4,050 masl. The phase is approximately 650 meters wide east-west and 1000 meters north-south. Phase 2 will produce about 16.5 million tonnes of ore at a stripping ratio of 2.44:1 (tonnes waste per tonnes of total ore). An illustration of the Phase 2 pit is shown in Figure 16-16.

FIGURE 16-16: PAMPACANCHA PIT PHASE 2

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Total ore reserves included in the final two pits are estimated to be 532.5 million tonnes and 569.4 million tonnes of waste material. Approximately 6.8 million tonnes of medium and low grade ore have been hauled to the temporary stockpiles at the end of 2020. This material will be sent to and reclaimed from the stockpiles through the life of mine. At the end of the mine's life, the ore from these stockpiles will be reclaimed and processed. Table 16-4 presents the Mineral Reserves estimates breakdown by Phases.

TABLE 16-4: CONSTANCIA AND PAMPACANCHA BREAKDOWN BY PHASES AS OF DECEMBER 31ST, 2020

16.3. MINE PRODUCTION SCHEDULE

16.3.1. PRODUCTION SCHEDULING CRITERIA

The operating and scheduling criteria used to develop the mining sequence plans are summarized in Table 16-5. Pit and mine maintenance operations are scheduled around the clock. Allowances for downtime and weather delays have been included in the mining equipment and workforce estimations.

TABLE 16-5: MINE PRODUCTION SCHEDULE CRITERIA

16.3.2. MILL FEED AND CUT-OFF GRADE STRATEGY

The mine plan criteria are summarized in Table 16-6.

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TABLE 16-6: MINE PLAN CRITERIA

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A variable cut-off grade strategy was implemented to bring forward the higher grade ore from the pit. Delivering higher-grade ore to the mill in the early years will improve the net present value and internal rate of return of the operations.

Priority plant feed consists of high grade material (NSR grade above 10 $/t). The lower grade material (NSR below 10 $/t) is being fed as needed or otherwise sent to the long term stockpile to be reclaimed at the end of mine's life.

The blending of ore is needed for plant fed in order to reach contaminants levels below the maximum allowed, for Zinc <5% and Lead <2%. Rejected Ore, by contaminant levels, also sent to long-term stockpile to be reclaimed in further years with fewer contaminants.

16.3.3. OVERBURDEN STRIPPING REQUIREMENTS

Mineral reserve tabulations by bench, phase, and mine production scheduling program (MSSO and MSReserves, modules from MineSight® software) are used to analyze long-term stripping requirements for the Constancia Operations. Elevation and phase order dependencies and sinking rate controls are used in conjunction with mill ore production targets and an internal NSR cutoff of 6.14 $/t to simulate open pit mining. The program, through successive iterations, allows the user to examine waste stripping rates over the life of the mine and their impact on ore exposure and mill head grades.

The analysis determined a production stripping requirement of approximately 32 - 48 million tonnes of waste and 32 million tonnes of ore by year. About 75.8 million tonnes of ore will also be mined and stockpiled during this period and will be reclaimed when the plant feed requirements exceed the daily mine production.

16.3.4. MINE PLAN

The mine production plan includes 569.4 Mt of waste and 532.5 Mt of ore (from both pits and stockpiles), yielding an average stripping ratio (waste/ore) of 1.1. An average yearly mining rate of 77.0 Mtpy, through the first 13 years, with a maximum of 81 Mtpy, is required to provide a nominal ore process feed rate of 31.3 Mtpy based on a variable throughput by ore type (90-94 ktpd and 94% availability). The ore production schedule for the operation presents step down trends with the Cu grade averaging 0.44% Cu from 2022 to 2024 when Pampacancha is at its peak production, then reducing to 0.34 % Cu from 2025 to 2029 with the preferential scheduling of higher grade mineralization from the Contancia pit and finalizing decreasing to 0.25% Cu to the end of the mine life. The average LOM average grade is 0.311% Cu, 0.009% Mo, 0.065 g/t Au and 3.04 g/t Ag for the remaining 17 years of the mine life.

The destination flowsheet of Constancia operations is shown in Figure 16-17 while Table 16-7 presents the distances in km between the Mine Phases and the main facilities. The mine schedule drawings for the life of mine are shown in Figure 16-18 to Figure 16-23.

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FIGURE 16-17: DESTINATION FLOWSHEET

TABLE 16-7: DISTANCES BETWEEN PHASES AND FACILITIES

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FIGURE 16-18: MINE PLAN OF PERIODS 2021 AND 2022

FIGURE 16-19: MINE PLAN OF PERIODS 2023 AND 2024

FIGURE 16-20: MINE PLAN OF PERIODS 2025 AND 2026

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FIGURE 16-21: MINE PLAN OF PERIODS 2027 TO 2030

FIGURE 16-22: MINE PLAN OF PERIODS 2031 TO 2034

FIGURE 16-23: MINE PLAN OF PERIODS 2035 AND 2037

The Mine schedule details for the total annual movement by destination and by phases, stripping ratios, mill feed by ore type and by phases are presented in Table 16-8 and Figure 16-24 to Figure 16-26.

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TABLE 16-8: CONSTANCIA OPERATIONS MINE PLAN AT JANUARY 1, 2021

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FIGURE 16-24: TOTAL MOVEMENT BY DESTINATION

FIGURE 16-25: TOTAL MOVEMENT BY PHASE

FIGURE 16-26: ORE TONNAGE MILLED BY PHASE

The processing schedule corresponding to the combined LOM plans of Constancia and Pampacancha is illustrated in Figure 16-27 and in Table 16-9 to Table 16-11.

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FIGURE 16-27: TOTAL COPPER PRODUCTION (2021 - 2037) IN CONCENTRATE

TABLE 16-9: CONSTANCIA ORE PROCESSING PLAN 2021 - 2037

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TABLE 16-10: PAMPACANCHA ORE PROCESSING PLAN 2021 - 2037

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TABLE 16-11: ORE PROCESSING PLAN 2021 - 2037 (TOTAL)

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16.4. MINE FACILITIES

16.4.1. WASTE ROCK FACILITY (WRF)

The final design geometry for this facility incorporates the same slope profile as the original WRF: 20 m high benches with 1.4H:1V (36°) of bench slopes and 32 m wide benches. The overall slope of the stockpile will be about 3.0H:1V (18°). The current remaining capacity of the WRF is estimated at 413 Mt. The waste material balance is shown in Table 16-12 and the Waste Rock Facility is represented in Figure 16-28.

TABLE 16-12: CONSTANCIA, WASTE MATERIAL BALANCE

FIGURE 16-28: WASTE ROCK FACILITY

16.4.2. ORE STOCKPILES

The Constancia mine is planned to have four operational stockpiles (Figure 16-29), which will segregate the ROM material by ore type and grade range. The lift height of these stockpiles ranges between 12 and 20 meters. The planned maximum stockpiles capacities are:

Stockpile 01 : 600 Kt

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Stockpile 02 : 500 Kt
Stockpile 03 : 600 Kt
Stockpile 04 : 5,000 Kt (stock located inside WRF, lift 4240, and used until 2025)
Stockpile 05 : 2,000 Kt (a finger from stock 04 destinated for Pampacancha ore)
Stockpile 06: 70,000 Kt (low-grade stock located at south from WRF)

FIGURE 16-29: LOCATION OF ORE STOCKPILES

16.4.3. TAILINGS MANAGEMENT FACILITY (TMF)

The Tailings Management Facility (TMF) area is where mill tailings are placed behind waste rock containment buttresses. The buttresses' material will come throughout the life of the mine, from mining operations and quarries. The TMF area is located southwest of the mining facilities area. The design for the TMF includes ramps access to the lower and upper levels for the regular mining fleet (240 tonnes trucks) and a smaller fleet called HCW, which has been used since October 2015 to send NAG material from the Constancia pit to the TMF area.

At the end of the impoundment, the final dam crest elevations will be:

Tailings Dam equal to 4,190 masl
Saddle Dam equal to 4,190 masl, where the height of the saddle dam will be 12 m, with 4 m of minimum freeboard.
The average dry density of deposited tailings for this expansion was considered 1.45 gr/cm3 and will be evaluated at the next level of study.

Hudbay is managing the planning, construction and operation of the Tailings Management Facility (TMF) and the Waste Rock Facility (WRF) with design and engineering support from Knight Piésold, also an Engineer of Record (EOR) of those structures.

The TMF Stacking Plan is illustrated in Table 16-13 and the final TMF design is presented in Figure 16-30.

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FIGURE 16-30: TMF - CONSTANCIA - FINAL CONFIGURATION

TABLE 16-13: TAILING MANAGEMENT FACILITY (TMF) STACKING PLAN

16.5. MINE EQUIPMENT

EQUIPMENT OPERATING PARAMETER

Mine equipment was selected based on the production requirements as shown in Table 16-15 and Table 16-16. At the mine site, three crews operate with two twelve-hour shifts per day, 365 days of the year. No significant weather delays are expected and the mine will not be shut down for holidays.

The parameters used to determine productivity are listed in Table 16-14. There are several different rock types at Constancia; however, for production estimation, the weighted average of all rock types was used. Significant loading and haulage equipment is equipped with electronic monitors which optimize loading time. The production is reported in dry metric tonnes, which is consistent with the reserve model. Moisture content is expected to range between 3.5 and 4.5 percent. Haulage calculations relied on a 4.0% moisture content.

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TABLE 16-14: MATERIAL CHARACTERISTICS

MINE EQUIPMENT CALCULATION

Mine equipment requirements were developed based on annual tonnage movements from the mine production schedule based on 15 meter bench heights, two twelve hour shifts per day, 365 days per year operation, manufacture machine specifications and material characteristics specific to the deposit (Figure 16-31 Truck Study Results).

A summary of fleet requirements by period is shown in Table 16 13 and Table 16 14. This equipment is considered necessary to perform the following tasks:
production drilling;
loading and hauling ore to the primary crusher (located on the east side of the pit), waste rock facility (WRF) and to the tailings dam (TMF);
maintaining mine haulage and access roads; and
maintaining the waste rock facility (WRF) areas, ore stockpiles, tailings management facility (TMF), and berms and regrading of slopes and final surfaces.

TABLE 16-15: MINE EQUIPMENT 2021 - 2029

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TABLE 16-16: MINE EQUIPMENT 2030 - 2037

FIGURE 16-31: TRUCK STUDY RESULTS

16.6. MINE OPERATIONS

Open pit mining at the Constancia operation is based on conventional open pit mining techniques. The Constancia operation consists of two pits, Constancia and Pampacancha. The Constancia pit operation started in 2014, while Pampacancha will begin in 2021.

The mine production plan contains 569.4 Mt of waste and 532.5 Mt of ore (from pit and stockpiles), yielding a stripping ratio (waste/ore) of 1.1 to 1. An average yearly mining rate of 77.0 Mtpy, through the first 13 years, with a maximum of 81 Mtpy, is required to provide a nominal ore process feed rate of 31.3 Mtpy based on a variable throughput by ore type (90 to 94ktpd and 94% available). LOM average grades are 0.311% Cu, 0.009% Mo, 0.065 g/t Au and 3.04 g/t Ag, where the mine's life is 17 years.

The priority to feed the process plant will involve optimizing the net value based on NSR ($/t), where high value material (HG) is provided first. The low value material (LG) will be fed as needed, sent to stockpiles, or will otherwise be sent to the waste rock facility (WRF).

16.6.1. DRILLING

For primary drilling production, single-pass 270 mm diameter (or 10 5/8") drills are used. After evaluating equipment dimensioning, three drill rigs PV-271 were selected to achieve the production rate. The drilling ratio of the drill rig is approximately 45 m/h. Pre-split drilling is done with a Smart Rock D65 drill (downhole). Equipment used to improve the stability of the slopes and optimize the design of the phases.

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DRILL PATTERNS AND PARAMETERS

Drill patterns were designed only for ore and waste material, based on geomechanical characterization, Bond Work Index (BWI) and Sag Power Index (SPI) range. Table 16-17 presents the parameters for the drilling production.

TABLE 16-17: PARAMETERS FOR DRILLING PRODUCTION

16.6.2. BLASTING

For wet holes, Heavy ANFO 73 (70% of Emulsion and 30% of ANFO) and dry hole, Heavy ANFO 55 (50% emulsion and 50% of ANFO) is used. Based on the evaluation of productivity, powder factors are estimated to be about 0.40 Kg/t for ore and 0.30 Kg/t for waste. The estimated blast design, explosive type, and powder factors for the Constancia operation are based on the mine operation results. The overview of blasting practices includes the supply and transportation of explosives to the hole, loading, stemming, tie-in, evacuation of work and blast areas, and the ignition of the blast.

16.6.3. LOADING

Two 27 m3 (Hitachi EX5600-6) shovels and a 19 m3 (CAT 994H) loader are used to excavate blasted materials. The loader provides flexibility for blending purposes. Pampacancha is expected to enter into production in 2019, and will require a loader at the start of the mining activities.

All phases have been designed to achieve high productivity, taking advantage of double-side loading and working at faces around ~60 meters in width.

16.6.4. HAULING

For life of mine, ore and waste will be transported by 240-tonne capacity haul trucks (CAT 793F). The use of this class of trucks minimizes road congestion, labor requirements, and operating costs. 240-tonne trucks require a minimum haulage road width of 32 meters. Haulage routes were designed to improve productivity and the operating hours needed per year. Table 16-18 and Figure 16-32 show parameters used in the truck & shovel study.

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TABLE 16-18: TRUCK FIXED TIME

FIGURE 16-32: TRUCK SPEED BY GRADIENT

16.6.5. SUPPORT EQUIPMENT

Major support equipment includes mine equipment that is not directly responsible for production but which is used on a regular basis to maintain pit and ex pit haul roads, pit benches, WRF and TMF and to perform miscellaneous construction work as needed. Mine support fleet equipment includes:

Crawler dozers, D10T2 class;
Rubber-tired dozers, 824K - 854K class;
Motor graders, 32M - 16M class; and
Water truck, 777G class.

In general, the rubber-tired 854 and 824K-class dozers are used in the pit to clean up around the primary loading units, while the track dozers are used for haul road construction, pit development, WRF and TMF management, and final re-grading requirements. The graders and water trucks are used respectively to maintain roads and control dust.

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16.7. MINE ENGINEERING

16.7.1. HYDROGEOLOGICAL PARAMETERS

A regional map of the estimated hydraulic conductivities covering both the Constancia and Pampacancha pits is presented in Figure 16-33.

FIGURE 16-33: PREDICTED HYDRAULIC CONDUCTIVITIES IN CONSTANCIA AND PAMPACANCHA PITS

CONSTANCIA PIT

The Constancia pit will operate from 2021 to 2037. At the end of mine life, the pit footprint at the surface will be approximately 15% larger, and the ultimate pit floor will be around 30 m deeper than the previous pit design (3,750 m elevation versus 3780 m elevation).

Predicted total dewatering rates for the Pit Expansion Design and the influence area should be close to 180 L/s as approved in the Groundwater Use Licence. The design was partially driven by the requirement for mine water supply. However, the main goal of dewatering is to depressurize the pit wall to assure the overall pit slope stability. Although the pit is slightly larger in the expanded design, the depressurization targets are expected to be less than the 100 m distance. Assuming the bedrock hydraulic conductivity at depths below 250 m is greater than 1 x 10-8 m/s, pit dewatering using a combination of vertical dewatering wells and/or horizontal drain holes should be feasible. Since limited hydraulic conductivity tests have been conducted below this depth, vertical wells and horizontal drain holes could be less effective if the hydraulic conductivities are inferior to10-8 m/s.

Suppose horizontal drain holes are used to achieve depressurization in targeted sections of the pit. In that case, it is expected that the drain hole spacing will range between 10 and 15 m (assuming a bedrock hydraulic conductivity of 1 x 10-8 m/s to 1 x 10-7 m/s). In the weak upper zones where higher hydraulic conductivities were observed, the spacing could be somewhat larger (~30 m) and the installation of vertical dewatering wells could be more applicable. The final spacing and length of the drain holes would be established and refined during operation by the phased installation of the drain hole system.

The number of vertical dewatering wells required to achieve depressurization would likely range between 15 and 50, based on the 2013 hydrogeological assessment and the subsequent recognition of reduced depressurization requirements. The actual number of wells will depend on the amount of depressurization required for slope stability and the hydraulic conductivity of the bedrock below 250 m depth. It is assumed that horizontal drain holes will likely be required below depths of 250 m. If in-pit wells are used, depending on the mine plan, wells may need to be progressively moved and replaced as the pit is widened and deepened. As mining progresses, optimization of the dewatering system will likely be necessary based on hydrogeological data collected from dewatering wells and the network of piezometers.

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In practice, a staged approach to dewatering is applied where geological, structural, and hydrogeological information that is gathered at each stage of mine development is used to refine and optimize the dewatering activities. The staged approach focuses on dewatering efforts in areas of identified concern (i.e., poorly draining slopes with low factors of safety), and selecting the most appropriate method to mitigate high pressures, if present. This staged approach requires detailed monitoring of the dewatering progress, dewatering well-pumping rates, and pit inflows.

Dewatering progress is typically assessed by observations gathered from multi-level piezometers and monitoring wells that are installed at critical locations behind the pit walls. Monitoring of water levels in pumping wells is typically not sufficient for monitoring dewatering progress since well inefficiencies can cause the water level in the pumping well to be different from correct groundwater elevations near these installations.

PAMPACANCHA PIT

The hydrogeological model enabled the planning of the pit drainage strategy. As a result, the construction of five perimeter wells, which are to be implemented before mining, is required. These wells will be adapted to the requirements and geometry of the pit design.

Besides, in the first year of mining (2021), the two existing wells, constructed during the Feasibility Study and located inside the pit, will be used, contributing to the start of the draining process.

The expected flow rate will be increased as mining and pit exploitation progresses until a maximum speed of approximately 30 liters per second (L/s) is achieved. Based on the modeling undertaken, it is estimated that each well will have approximate flow rates varying between 5 - 12 L/s, depending on the local geology of each well.

To monitor the advancement of pit dewatering activities, it is recommended that three additional piezometers be installed to complement the network of existing piezometers already within the pit area.

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17. RECOVERY METHODS

17.1. INTRODUCTION

17.1.1. GENERAL LAYOUT AND FACILITY DESCRIPTION

The process plant and facilities have been designed by Ausenco (Australia, Canada and Peru regional offices). The design was based on the GRDMinproc DFS design, with one phase of optimization of the comminution circuit, followed by front end engineering and design (FEED), and detailed engineering as part of the EPCM contract.

The mine is located on mountainous terrain. The plant site was selected as being the closest site available to the mine that is both economical and practical to develop, but significant earthworks were required to establish a multi-tiered pad. The crushing facility is located on the hillside in order to minimize expensive ROM pad construction. The layout allows the crushing facility to be free-draining, hence limiting the amount of ground work maintenance during the rainy season.

The primary crusher, belt conveyors, thickeners, tanks, flotation cells, mills and other equipments are located outdoors without buildings or enclosures. To facilitate the appropriate level of operation and maintenance, the molybdenum concentrate bagging plant, copper concentrate filters and concentrate storage are housed in clad steel buildings. A 3D view of the plant in 3D is shown on Figure 17-1.

FIGURE 17-1: 3D VIEW OF THE CONCENTRATOR

Ausenco designed the processing facilities according to the process design criteria presented in Table 17-1.

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TABLE 17-1: KEY PROCESS DESIGN CRITERIA

The process flowsheet for the Constancia recovery process, as represented in Figure 17-2, is similar to other major Copper-Molybdenum plants in Latin America.

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FIGURE 17-2: CONSTANCIA PROCESS FLOWSHEET

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17.1.2. BUILDINGS

The following buildings were constructed as part of the facilities:

1) Office building - originally the EPCM offices on site. These were converted for general administrative offices.

2) Warehouse and workshop - the building shell was constructed as part of the original construction with services running to the building; however, completion was deferred to a later date. Process plant maintenance is working out of the construction shops originally set up for the SMP contractor. Warehousing is managed within a fenced platform near the process plant using construction structures put in place for commissioning.

3) A plant office building with a control room, server room, a DCS engineering station and a shift boss office.

4) A combined chemical and metallurgical laboratory building.

5) A crusher control room.

6) A combined copper concentrate filtration and storage building.

7) A molybdenum concentrates packing shed.

17.2. PROCESSING PLANT

17.2.1. GENERAL

The general layout of the processing plant is shown in Figure 17-3.

FIGURE 17-3: PROCESSING PLAN LAYOUT

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The processing plant has been laid out in accordance with established industry best practices for traditional grinding and flotation plants. The major objective was to make the best possible use of the natural ground contours to minimize pumping requirements by using gravity flows and to reduce the height of steel structures.

To optimize the cost of the major footings for the SAG mill, the height of the SAG mill above grade was minimized by situating the cyclone feed sump as low as possible. The mill cyclones have been located so that the cyclone overflow can gravitate into the rougher conditioning tank.

At the tailing end of the flotation bank, the copper tailings thickener has also been situated to facilitate gravity flow and eliminate the requirement for another large set of pumps.

Due consideration has been given to the layout of the molybdenum plant to facilitate good housekeeping and occupational health and safety requirements. The potentially hazardous nature of the molybdenum bearing froth very often results in difficult housekeeping conditions within the work area. To this effect, the use of sodium hydrogen sulphide (NaHS) for copper depression and separation from molybdenum concentrate has the potential to generate hydrogen sulphide (H2S) gas.

The copper and molybdenum plants are independent of each other and separated by a reasonable distance. Furthermore, the molybdenum concentrator is located downwind of major high occupancy plant areas (prevailing wind on site is from the north). Similarly and given the fact that reagents can release hazardous gases, the storage facility is also located downwind of high occupancy plant areas, but close enough to the flotation area to avoid long runs of piping.

Due to these considerations the molybdenum plant is fenced out to strictly control personnel access. Similarly, a thorough training program has been put in place to educate anyone entering the plant about the hazards and safety procedures.

Rainwater runoff and general spillages from the processing plant are fully contained and directed to the process plant sediment pond. Again, the natural topography of the site is used to direct the flow to the pond, located at the western end of the site to avoid the need for pumps.

Wherever possible, the major substations and motor control centre (MCCs) have been located as close as possible to the electrical drives to minimize cable run lengths.

17.2.2. PRIMARY CRUSHING

A 63" x 118" gyratory crusher receives ROM of up to 1 m in size and reduces it to less than 125 mm (Figure 17-4). A variable speed belt feeder delivers ore from the crusher chamber to the coarse ore stockpile feed conveyor. The site mobile crane is used for major crusher maintenance. A drive-in dump pocket makes it possible to clear the crusher with a small FEL or bobcat.

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FIGURE 17-4: PRIMARY CRUSHER ISOMETRIC VIEW

The crusher is an FLSmidth 63" x 118" gyratory crusher with a 1000 kW motor. This is likely to be under-utilized for most of the highly fractured ore, but will ensure throughput when processing an area of competent rock. The ROM bin and crusher surge pocket has the volume to hold 1.6 truck loads

17.2.3. CRUSHED ORE STOCKPILE AND RECLAIM SYSTEM

The grinding circuit requires two SAG mill feed conveyors (Figure 17-5). Dual reclaim chambers (near- parallel) are used to house a total of four apron feeders that draw ore from the crushed ore stockpile to the SAG mill feed conveyors. A secondary egress or emergency tunnel is provided through the connection of the two reclaim chambers.

FIGURE 17-5: COARSE ORE STOCKPILE ISOMETRIC VIEW

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17.2.4. GRINDING

The grinding circuit consists of two trains of grinding mills, each train identical, treating 1875 t/h (Figure 17-6). Each train consists of a single FLSmidth semi-autogenous grinding (SAG) mill, a ball mill and a potential addition of a pebble crusher (SABC). The ball mills are operated in closed circuit with a FLSmidth (Krebbs) cyclone cluster to produce a product from the grinding circuit of 80% passing until 130 µm.

The SAG mills are fitted with a dual pinion drive system and are driven by a variable-speed drive comprised of a SER hyper-synchronous drive. The SAG mills are 11 m (36') dia, 7.3 m EGL (26.5'). Each SAG mill is driven by two 8 MW motors. The 16 MW, twin pinion, SAG mills were selected over the equivalent GMD to reduce both capital and commissioning costs.

The two 16 MW ball mills selected are fitted with twin pinion drives operated at fixed speed. The mills are 7.9 m (26.5') dia, 12.4 m EGL (41'). Like the SAG mills, each ball mill is driven by two, 8 MW motors. Each of the eight mill motors are interchangeable, and started using conventional liquid resistant starter circuits (although once started, the SAG mill SER system takes over control enabling limited variable speed control). The commonality of motors reduces the number of maintenance spares required to be held on site.

As a result of the processing of softer ores from Pampacancha during the 2021 through 225 period, the installation of the pebble crushers, originally contemplated for year 6, has been deferred until at least 2023. The pebble crushing system originally considered two hydrocone or shorthead pebble crushers with 600 kW drives each; however, the exact type and sizing of the equipment is under review based on current plant performance and may change during the early works engineering period.

Each grinding train was originally configured with a single (duty only), 1.5 MW Warman 650 MCR M200 variable speed cyclone feed pump but since commissioning a larger-capacity Excellence ESH-650MU was installed as a trial unit on grinding line 1. Pending successful results, the other line will be upgraded. A full spare assembly is held on site and changed out when required. The layout allows for the flexibility to recycle cyclone underflow to the SAG mills; a single "drive-on" grinding floor; a simple cyclone tower; clear access to the wet end pumps; and clear access to scats clean-up from the ball mill.

FIGURE 17-6: GRINDING CIRCUIT ISOMETRIC VIEW

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The capital cost was reduced by not installing standby cyclone feed pumps. Spare units can be installed in approximately six hours. Easy access by mobile cranes is provided and it is possible to only change the wet end of the pump which minimized downtime.

17.2.5. CU-MO BULK FLOTATION

The function of the copper flotation area (Figure 17-7) is to recover copper and molybdenum into a bulk concentrate, while rejecting coarse non-sulphide gangue to the rougher tailings and rejecting zinc, lead and iron sulfides to the cleaner scavenger tailings stream.

Slaked lime is added in the grinding circuit to ensure the rougher circuit pH is maintained at its set value. Cyclone overflow from each of the primary grinding hydrocyclone batteries, comprised of Weir-Vulco Cavex CVX13 hydrocyclone, reports via gravity to the corresponding bulk rougher circuit. The bulk flotation consists of two copper rougher flotation banks, each consisting of seven Outotec 300 m3 forced air mechanical flotation cells, operated at a pH of approximately 10. These produce a low-grade rougher concentrate containing copper, silver, gold, molybdenum and other nonvaluable sulfide and gangue minerals, which reports to the regrind hydrocyclone feed sump. The bulk tailings flow by gravity to the tailings thickener for water recovery and eventual disposal at the TMF. Spillage and clean-up water are removed by two rougher flotation area floor sump pumps.

The bulk rougher concentrate is pumped to a hydrocyclone cluster. The coarse hydrocyclone underflow is pumped to the bulk cleaner regrind mill, consisting of a 3 MW Glencore Technologies M10,000 IsaMill.

The regrind mill discharge is combined with the hydrocyclone overflow and pumped to the first bulk cleaner flotation cell. In the bulk cleaner circuit, the pH is increased to 11.0 with slaked lime.

The copper cleaner circuit consists of three stages of cleaner and one stage of cleaner-scavenger flotation cells. The first cleaner stage consists of four Outotec 130 m3 forced air flotation cells. The second consists of three Outotec 130 m3 forced air flotation cells, and the cleaner-scavenger consists of six Outotec 130 m3 cells. The third cleaners are comprised of two Eriez CPT column flotayion cells, each 4.88m diameter by 12m tall and equipped with wash water.

The flowsheet is as standard counter-current configuration with scavenging. Intermediate concentrate from the first cleaner reports to the second cleaner feed sump, while the tails report to the cleaner-scavenger stage. Cleaner-scavenger concentrate returns to the regrind and cleaner-scavenger tails combines with the rougher tails and reports to the tailings thickener. Concentrate from the second cleaner reports to the third cleaners, and concentrate from the third cleaners reports to the 24.0m diameter bulk concentrate thickener. Third cleaner tails report to the second cleaner and second cleaner tails report to the first cleaner.

The column cells are supplied with dedicated high-pressure forced air and the mechanical flotation cells are supplied with low pressure air from dedicated low pressure air blowers.

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FIGURE 17-7: CU-MO FLOTATION CIRCUIT ISOMETRIC VIEW

17.2.6. CU-MO FLOTATION ARRANGEMENT

The cleaning circuit, consisting of 1st, 2nd and scavenger flotation stages is arranged in a counter current arrangement, whereby the largest flow, the tails, flows by gravity from one stage to the next. All cells are the same size. The concentrate launders and pumps are all arranged on the outside of the plant to allow for easy maintenance access. This arrangement also allows for the number of cells in each stage to be modified, with some minor pipework changes, during operations.

To reduce capital costs and improve the overall plant, the volume held in each sump has been minimized.

Under normal operating conditions, the slurry that is collected under the flotation areas is composed of concentrate overflowing from the top of the cells. This is collected and pumped from the sump pump area back into the process.

During routine shutdowns or power failures, the mechanical cells can stop and do not need to be drained. For maintenance, it will be occasionally necessary to dump the contents of the flotation cells.

Since the value of the slurry in each rougher cell is low, the sump is not designed to contain an entire cell. In a typical maintenance shutdown the slurry dumped from the rougher cells is expected to be pumped to tails by the rougher area sump pump. In exceptional cases, i.e. emergencies, if this pump fails or if there is no power, then the rougher slurry will overflow the sump and runs along a channel to the plant run off pond.

In the case of the 3rd cleaner columns, the sump has enough volume to hold one of the two columns. When the plant is down for more than three to four hours then the columns will need to be drained. If the sump pump is down then the sump will over flow into the 2nd cleaner sump.

17.2.7. MOLYBDENUM SEPARATION

The function of the molybdenum flotation area (Figure 17-8) is to recover molybdenum into a molybdenum concentrate and reject copper to the tailings as a copper concentrate. The stages of molybdenum flotation are: rougher, 1st to 5th stage cleaners and cleaner scavengers.

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Underflow from the copper-molybdenum concentrate thickener is pumped to a 38 m3 agitated molybdenum rougher conditioning tank. NaHS is added to the conditioning tank to inhibit the flotation of copper, with promoter (diesel emulsion) also added to enhance the flotation of molybdenum. From there the overflow flows to the ten 10 m3 molybdenum rougher flotation cells. The tailings from the rougher cells form a portion of the copper concentrate and are pumped, with the molybdenum cleaner scavenger tailings, to the copper concentrate thickener. The molybdenum rougher concentrate reports to the molybdenum 1st cleaner flotation cells.

All of the molybdenum cells, other than the final cleaner, will be WEMCO type self aspirating cells from FLSmidth.

The molybdenum cleaner circuit consists of five stages of cleaning and one bank of cleaner scavenger flotation cells. The molybdenum 1st stage cleaner consists of six 28.3 m3 mechanical flotation cells. From these, a concentrate containing molybdenum values is recovered and flows to the molybdenum 2nd stage cleaner flotation cells for further upgrading. The tailings from the molybdenum 1st stage cleaner flow to the four 14.2 m3 molybdenum cleaner scavenger flotation cells.

The molybdenum 2nd stage cleaner flotation consists of four 14.2 m3 flotation cells. From the flotation cells the concentrate reports to the molybdenum 3rd stage cleaner flotation cells for further upgrading. The tailings from the molybdenum 2nd stage cleaner flotation cells reports to the molybdenum 1st stage cleaner flotation cells.

The molybdenum 3rd stage cleaner flotation consists of five 8.5 m3 flotation cells. From the cells, the concentrate reports to the molybdenum 4th stage cleaner flotation cells for further upgrading. The molybdenum 3rd stage cleaner tailings report to the molybdenum 2nd stage cleaner flotation cells.

The molybdenum 4th stage cleaner flotation consists of four 4.3 m3 flotation cells. The cells produce molybdenum concentrate which reports to the molybdenum 5th stage cleaner flotation cells. The tailings from the molybdenum 4th stage cleaner cells reports to the molybdenum 3rd stage cleaner flotation cells.

The molybdenum 5th stage cleaner flotation consists of a single 4 m3 Xstrata Jameson flotation cell. The cell produces the final molybdenum concentrate. The tailings from the molybdenum 5th stage cleaner cell reports to the molybdenum 4th stage cleaner flotation cell.

The molybdenum cleaner scavenger consists of four 14.2 m3 mechanical flotation cells. A low-grade concentrate is recovered and is pumped back to the copper/molybdenum feed conditioner tank. The tailings from the molybdenum cleaner scavenger are pumped to the copper concentrate thickener. Spillage and water are removed by a sump pump in the flotation building.

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FIGURE 17-8: CU-MO SEPARATION CIRCUIT ISOMETRIC VIEW

17.2.8. COPPER CONCENTRATE THICKENING AND FILTRATION

The tailings from the molybdenum flotation roughers and molybdenum cleaner scavengers report to a 24m diameter copper concentrate thickener (Figure 17-9), via a static screen to remove tramp material. Flocculent is added to enhance settling. The clarified thickener overflow reports to the process water pond, to be pumped to the process water tank. The thickener underflow is removed at 60% solids via duty/standby peristaltic pumps.

The thickened and de-tramped slurry is stored in two 860 m3 agitated tanks. These provide a 24 hour surge capacity, allowing filter maintenance to be conducted without affecting mill throughput. The filter feed is pumped to one of two LAROX horizontal plate pressure filters with a combined maximum filtration rate of 91.2 t/h. The filters operate with a total cycle time of 9.8 min and are designed to produce a filter cake of less than 8.5% moisture. The filter cake is dropped to the floor and then transferred to the concentrate storage pile by a front-end loader. Filtrate, cloth wash, and flushing water are returned to the copper concentrate thickener.

FIGURE 17-9: CU CONCENTRATE THICKENING AND FILTRATION AREA ISOMETRIC VIEW

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17.2.9. COPPER CONCENTRATE STORAGE AND LOADOUT

Annual concentrate production rates are showed at Figure 17-10

FIGURE 17-10: LOM COPPER CONCENTRATE PRODUCTION

After filtering, the copper concentrate is transferred to a copper concentrate product storage stockpiles by front end loader inside the copper concentrate and load-out building. The building contains a live stockpile with capacity for seven days production (12,000 t).

A fleet of approximately 140 tractor trailers, each with 36 t hauling capacity, are used to transport the concentrate to the port. A single front-end loader is used to collect concentrate as it falls from the filters and either load it directly onto the trucks or temporarily store it in the same building. The trailers are fully enclosed units with pneumatically operated lids. Locks and security seals are applied at site. A truck scale weighs each truck before leaving site and upon arrival in Matarani. One primary contractor is used for the base load with surge capacity being made up with several smaller contractors.

The trucks have a two day round trip from site to the port and back, depending upon road conditions. Rigorous control procedures are applied during transport including separate escorts, GPS tracking and monitoring, controlled resting periods for the drivers, and secure stopping points en route. The position and velocity of every convoy is monitored continuously from the security control center in Constancia.

17.2.10. MOLYBDENUM CONCENTRATE THICKENING, FILTRATION AND LOADOUT

The molybdenum concentrate gravitates to a 4 m diameter thickener where it is thickened to 60% solids. The thickener overflow reports to the molybdenum process water tank, where it is used as process water in the molybdenum flotation circuit. The thickened concentrate is pumped to an agitated 48 m3 molybdenum surge tank, which has a residence time of 24 hours and acts as surge capacity for filtration.

The slurry is then pumped to a pressure filter to produce a filter cake of approximately 15% moisture, the next stage is the drying stage where we can obtain concentrate of 5% moisture. Filter cake is conveyed by a bobcat to the molybdenum bulk bagging plant. Molybdenum concentrate is bagged into 1 m3 bulk bags. The bulk bags are loaded onto trucks using forklifts.

The filtrate from the filter reports to the tailings thickener. The thickener overflow flows by gravity to the process water tank.

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17.2.11. TAILINGS DISPOSAL

The copper tailings flow by gravity from the two copper rougher tailings lines, and by pumps from the cleaner scavenger tailings line, to a 75 m diameter tailings thickener, where it is thickened to 58% solids. Thickened underflow is then pumped to the tailings dam. The thickener overflow gravitates to the process water tank where it is used in the grinding and copper roughing circuits.

The tailings pumping system consists of five centrifugal pumps in series with an HDPE lined steel pipe that crosses the Chilloroya valley. The pumping station at the plant site is designed to take tailings thickener underflow and dilution water and pump the mixture to the tailings dam on the other side of the valley. The pipeline follows the haul road route to save costs, increase maintenance access and reduce land disturbance. The road is built at low slope angles, typically less than 10%. The tailings pipeline has a maximum inclination of 12%. This is less than the angle of repose of the settled solids in the slurry, which is 15°. Consequently, if there is an un-programmed shutdown, for example a power outage, the slurry can sit in the line without having to be discharged or flushed. To reduce the risk of plugs during startup, all of the pumps in the pump station are fitted with variable speed drives.

The tailings pumping train can operate with as few as three pumps in operation allowing for extended maintenance times and coordinating with mill liner changes to minimize downtime. Under a normal scheduled stoppage, process water will be used to flush the solids from the line into the tailings dam.

17.2.12. REAGENTS

The reagents area is located to the west of the grinding and flotation areas.

Coarse lime and flocculent are delivered as solids to the reagents area by road trucks and stored in the reagents building.

A packaged flocculent mixing plant prepares flocculent for storage in an agitated tank for distribution. Other powdered reagents are mixed in agitated tanks before storage in tanks for distribution.

Liquid flotation collector and frother are received in one cubic meter iso-containers that are stored in the reagent store and taken by forklift to the reagent dosing area as required. The use of iso-containers also allows for placement in the grinding or flotation areas for flexibility in reagent dosing and plant scale testing.

Reagents are distributed by duty/standby pumps or multiple peristaltic metering pumps along flexible small diameter tubing (or by gravity in the case of temporary dosage by placing iso-containers within the grinding and flotation areas). Spillage and water are removed by the reagents area sump pump to tailings.

A diesel/fuel oil emulsion preparation facility allows addition of small doses of this reagent for molybdenum promotion while assuring even dosing in the slurry.

Lime silo and attrition mills prepare milk of lime for pH adjustment. Lime is distributed by a lime ring to the dosing points. Spillage and water are removed by a lime slaking area sump pump.

Nitrogen for blanketing the molybdenum separation cells is supplied by a pressure swing adsorption plant on site.

Carbon dioxide that will be used for pH regulation in the molybdenum plant is supplied in tanks.

NaHS is delivered as a concentrated liquid by tanker. It is diluted with water onsite for dosing. The decision to not have solid NaHS preparation facilities onsite was made to reduce CAPEX, reduce safety risks and reduce the working capital and management issues of storing solid NaHS.

Hydrogen sulphide gas that comes from the NaHS is extremely toxic. Sensors and suitable engineering controls have been designed in order to use NaHS safely in the plant.

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17.3. TYPICAL OPERATING PARAMETERS

The current level of reagent, power, water and grinding media consumption rates are shown in Table 17-2

TABLE 17-2: OPERATING PARAMETERS

17.4. METALLURGICAL ACCOUNTING AND CONTROL

The plant normally runs in automatic mode whereby all major units are monitored and maintained by the DCS system logic.

The plant operations are monitored and controlled from a central control room that is located between the grinding, flotation and regrinding areas. This control room has CCTV views of different plant areas and DCS views of all the plant equipment, valves and instrumentation. DCS interfaces have also been placed in the crusher control room, molybdenum plant control room, and the flotation deck. In 2019, a centralized control room was installed at the main camp, and control room operators were moved there. The onsite control rooms remain operational for redundancy.

Plant operations are facilitated by two online stream analysers (Outotec Couriers with stream multiplexers). These units constantly measure the principal streams for elements of interest and transmit the results to the operations and metallurgical staff for plant control.

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Metallurgical accounting is achieved by shift composite sample analysis in the site laboratory. The samples are taken from the cyclone overflow, rougher tailings, cleaner scavenger tailings and overall tailings.

Flows are accounted for by weightometres on the SAG mill feed conveyors and flow metres on the concentrate and tailings flows.

Final concentrate tonnage is controlled by weigh bridges that weigh trucks as they leave site and arrive at the port. A weightometer on the ship loading conveyor measures the concentrate weight loaded onto the ships.

17.5. OPTIMIZATION PROJECTS

17.5.1. PEBBLE CRUSHING

Prior to operations, it had been identified that the ore hardness would progressively increase, and the installation and commissioining of the pebble crushing system was originally scheduled for 2021. However, with the addition of softer Pampacancha ores, the installation of the pebble crushing systems is now scheduled for 2023 (commissioning in 2024). The timing will be further refined pending completion of ongoing geometallurgical studies. The pebble crushers, shown in Figure 17-11, are expected to cost approximately $23 million (Table 17-3).

FIGURE 17-11: 3D RENDER OF THE PEBBLE CRUSHER INSTALLATION

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TABLE 17-3: CAPITAL COST SUMMARY FOR PEBBLE CRUSHING

17.5.2. GRINDING

Optimization work around the grinding circuit is ongoing, focused primarily on throughput and grind optimization, process control integration, and SAG and ball mill configuration. Current or near-future optimization projects include:

Geometallurgical hardness characterization of the Constancia deposit, including 500 samples tested for SAG Grindability Index and Bond Work Index. The 500 samples represent the near term (5 years) ores and will be used to optimize the engineering and installation of the pebble crushing systems and better inform the decisions around SAG mill discharge system configuration and ball size.
Geometallurgical hardness characterization of the Pampacancha deposit, including 100 samples tested for SAG Grindability Index and Bond Work Index.
Conversion of the SAG mill discharge system to fully curved pulp lifter design with 60mm grates, including Growth recommendation of liner and discharge cone configuration. The proposed discharge system was modelled extensively by Molycop, Growth, and third-party consulting firms.
Completion of CiDRA PST system install and tuning and the extension of the advanced process control system to include smart cyclone selection.
Integration of Advanced Process Control system (Expert System) in the ball mill circuit, including hydrocyline feed flow and pressure control and SAG mill throughput permissives based on ball mill circuit constraints.
Implementation of the shell-mounted LoadIQ smart mill load system on the second grinding line (it has been installed succesfuly on one grinding line).

17.5.3. BULK FLOTATION

17.5.3.1. COPPER RECOVERY

The flotation optimization project derives from the simulation work (described in Section 13.2.2) and consists of a series of upgrade projects that will allow for an increase in the volumetric flow capacity of the of the rougher concentrate. The higher rougher mass recovery will allow for an increase the copper recovery equivalent to almost 3% at nominal feed grade and tonnage/grind conditions of the plant, as shown in Figure 17-12. To achieve this nominal recovery uplift, rougher mass recovery will be increased from approximately 8.5% to approximately 12% for nominal head grades (mass recovery fluctuates with head grade, grind size, and mineralogy).

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FIGURE 17-12: ROUGHER MASS RECOVERY VERSUS COPPER RECOVERY (LEFT) AND VERSUS ROUGHER CONCENTRATE MASS FLOW (RIGHT)

The upgrade projects include:

Installation of froth crowders and center launders in the first four rougher cells of each row
Upgrade or twin the concentrate pumping system as required and permittable
Upgrade the regrind cyclone feed pump as required and permittable
Upgrade the regrind cyclone to allow for higher treatment capacity
Increase the power draw of the regrind mill
Bypass cleaner-scavenger concentrate to 1st cleaner
Evaluate and, if required, upgrade the 1st cleaner and cleaner-scavenger dart valves
Installation of froth crowders and center launders on the cleaner-scavenger cells
Conversion of the column cells to operate in serial configuration with concentrate cleaning
Install froth dearation system ahead of the column feed pump to eliminate air locking

The estimated capital budget for the recovery projects is $7.85 million, Table 17-4.

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TABLE 17-4: CAPITAL BUDGET FOR RECOVERY PROJECTS

17.5.3.2. COPPER CONCENTRATE QUALITY

Current work on concentrate quality is focused on reducing levels of zinc and lead contanimants and increasing copper concentrate grades. This includes:

Evaluation of alternative, more selective reagents that will improve the effectiveness of depressants in the cleaners, reduce the pH and lime consumption, and increase the molybdenite recovery. At the time of writing, the study is currently in the Phase 2 plant trial stage.
Reduce the regrind size by increasing the Isamill power draw, improving the separation efficiency of the regrind hydrocyclones, and balancing the mass split to the regrind mill and hydrocyclones. This will reduce the mass of contaminants that are recovered due to locking with chalcopyrite.
Evaluation of lead and zinc depressants in the cleaning circuit. At the time of writing, the work is in the Phase 1 laboratory testing phase.

17.5.4. COPPER/MOLYBDENUM SEPARATION

Current work in the moly flotation circuit is focused on increase molybdenum recovery from copper concentrate, improving molybdenum concentrate quality, and reducing operating costs by minimizing the consumption of sodium hydrosulfide, energy, and other consumables. Projects include:

Installation of improved forced-air agitation mechanisms in the roughing cells, completed early 2021. The new mechanisms will allow for a much greater degree of fine level control at the low operating gas rates necessary in moly separation circuits.
Installation of improved forced-air agitation mechanisms in the 1st cleaning and cleaner-scavenging stage, scheduled for completion during the first half of 2021
Implementation of automatic density control in bulk, copper and molybdenum thickening stages to improve water recovery and usage in the moly plant and thereby reduce sodium hydrosufilde consumption
Installation of a new, upgraded nitrogen plant to reduce sodium hydrosulfide consumption
Review flowsheet and potential flowsheet modifications, pending survey and simulation program after completion of agitator upgrade project.

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17.5.5. OTHER PROJECTS

Other ongoing projects focused on improving the economic performance of the Constancia operation include:

Shovel-based XRF ore sorting pilot trial program to evaluate potential of bulk ore sorting. Phase 1 and Phase 2 have been completed and the technology has shown that, in principal, it has the potential to significantly improve the economic returns by upgrading the head grade to the . Ongoing studies are focused on developing a pilot-scale technical demonstration onsite and further quantifying the additional capital and operating costs associated with the face and truck sorting operations.
Various ongoing geometallurgical characterization projects, including flotation charactization and optimization of the Constancia and Constancia Norte deposits

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18. PROJECT INFRASTRUCTURE

The project infrastructure facilities include the Waste Rock Facility (WRF), Tailings Management Facility (TMF), the water management facility, electric power supply and transmission, and improvements to the roads and the port. Hudbay is in charge of these facilities' planning and operation (excepting the roads and port) and is working with different consultants for the engineering, construction and quality control/quality assurance (QA/QC) for their expansion.

An overall site layout is provided on Figure 16-1 showing the various storage facilities' locations relative to the Constancia and Pampacancha mines and the mill.

18.1. TAILINGS, WASTE AND WATER STORAGE

Hudbay is managing the planning, construction and operation of the Tailings Management Facility (TMF) and the Waste Rock Facility (WRF) with design and engineering support from Knight Piésold , also as Engineer of Record (EOR) of those structures.

18.1.1. TAILINGS MANAGEMENT FACILITY (TMF)

The TMF has been developed behind an east to west aligned, zoned landfill and rockfill cross-valley embankment dam that is expected to be raised in stages over the mine's life and has sufficient capacity to accommodate the tailings generated over the current LOM plan. The tailings are expected to be impounded behind the south side of the embankment in a basin that contains two natural valleys running north to south that is separated by a central ridge. The TMF is located on the south side of the Chilloroya River and has a final design elevation of 4,190 masl. The TMF has two stages. The first was designed by KP (engineering detail) at 4,160 masl and the second stage design by Golder (pre-feasibility study) that rises from 4,160 to 4,190 masl without moving the previous footprint.

Tailings have been deposited from designated off-take points from a distribution pipeline located along the upstream crest of the dam and around the perimeter of the facility, and are delivered through drop-bar pipes running down the upstream face of the embankment and down the valley perimeter slopes into the TMF. The points of active deposition have been frequently rotated to form a thin layered, drained, and well-consolidated beach that will slope away from the dam towards the south side of the TMF basin. Initially, the surface pond was located against the embankment in the east valley and has been displaced progressively upward and to the south by the development of the tailings beach. The surface pond will vary in size throughout the life of mine depending on the season, precipitation, and operational requirements.

The design of the TMF includes a LLDPE geomembrane liner and underdrains. The liner is intended to cover the base of the entire eastern valley and a majority of the western valley to provide containment in areas where the surface water pond will contact the bottom at any time over the life of mine. The liner will also be extended up the embankment's upstream face to the top of the downstream constructed raise. Underdrain systems include groundwater underdrains beneath the liner, tailings underdrains, and an embankment toe drain.

Water collected by the underdrains is conveyed to sumps located immediately downstream of the embankment in both the east and west valleys. The groundwater underdrain and tailings underdrain outlet pipes in each valley were installed in reinforced concrete encasements under the embankment. At the sumps, monitoring and control systems allow for automated water quality and flow rate determinations to be made prior to pumping back to the surface water pond. Basin roads are constructed to support basin grading activities, liner installation and anchorage, and tailings pipeline construction and access.

Instrumentation was installed in the TMF to monitor the structure throughout the life of the facility. Settlement and deformation monuments have been installed on the downstream embankment face and on the embankment crest during each major stage of construction. Additionally, vibrating wire piezometres have been installed within the embankment. The vista data vision system (VDV) and the Sensemetric system to monitor the main instruments on-line and in real time were both upgraded in 2020.

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According to the LOM plan of the TMF, and as per water balance, construction periods will occur in 2021, 2023, 2025, 2027, 2029, 2031, 2033 and 2035. Optimization of the LOM to ensure sufficient quality rock from the pit for the TMF construction has been completed, such that the mining fleet is utilized for hauling, placement and compaction of structural fill over the downstream and upstream zone of the embankment.

18.1.2. WASTE ROCK FACILITY (WRF), TOPSOIL AND PONDS

The Potentially Acid Generating and Non Acid Generating Waste Rock Facilities (PAG & NAG WRF) are located in the Cunahuiri Valley east of the Constancia Pit and provide storage for the PAG and NAG waste based on the updated mine plan developed in 2017. The facilities receive mine waste material from the operation of the Constancia pit and from the Pampacancha pit. Underdrain systems include groundwater underdrains to collect the water to WRF Retention

The Waste Rock Facility Retention Pond is located downstream from the WRF in the Cunahuiri Valley. The pond provides energy dissipation for surface water reporting to the downstream WRF Containment Pond via overland flows and the contact diversion channels and protects the WRF Containment Pond from possible falling waste rock and sedimentation. A Waste Rock Facility Containment Pond is located downstream of the WRF Retention Pond in the same Valley. The pond provides storage capacity for surface runoff, direct precipitation and seepage collected from the WRF and its contact diversion channels. The pond also stores Acid Rock Drainage (ARD) water, which is pumped to a lime neutralization facility at the process plant prior to its eventual use in the process plant. A foundation grout curtain has been incorporated into the embankment foundation to minimize seepage. In order to accommodate the maximum operational volume and additional storage to accommodate the 100-yr/24-hr storm event. The capacity of these ponds may be reassessed during the last years of the operation due to the potential expansion of the WRF.

Non-Contact Channels have been sized to accommodate the 200-year/24-hour (200-yr/24-hr) storm event with an additional freeboard for a 500-yr/24-hr storm event with no allowance for a freeboard. There are three Non-Contact Channels on operations: NC1, NC2 and NC3.

A pumping system between the WRF Containment Pond and the treatment plant is designed to pump ARD water from within the pond to the lime neutralization facility. This pumping rate is such that the WRF Containment Pond water level will be maintained at or below the maximum operational water level.

A topsoil stockpile is located in the Cunahuiri Valley downstream of the WRF Containment Pond and has a design storage capacity of 4.7 milion m3. This stockpile stores topsoil excavated during construction activities. Topsoil is used throughout operations for progressive closure of various facilities.

The Non-PAG Waste Rock Facility has a design storage capacity of 27 Mt (14 Mm3 at 1.9 t/m3). Non-PAG waste rock is used for the construction of various access roads, and haul roads as well as the Tailings Management Facility (TMF) embankment. Excess Non-PAG waste can also be placed within a temporary stockpile located 1 km southwest of the Constancia Pit.

18.1.3. WATER MANAGEMENT

Hydrologic and hydraulic analyses, including a detailed water balance, were completed to determine the frequencies and quantities of flows as well as the types of flow, location, and resulting sizes of the water management structures required to contain or convey them. Water was divided into process and non-process water, i.e. water that is not associated with the process plant and typically consists of stormwater runoff from areas within the mine site. The non-process water was further divided into contact or potentially acid generating (PAG) runoff water and non-PAG runoff water.

Structures directly associated with the site water management include the following:

Diversion Channels - These structures divert both contact and non-contact water around the mine site to the sediment or water management ponds, or to natural drainages down-gradient of the planned mine facilities (Constancia pit, WRF, etc.).

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Sediment Ponds - These structures typically accept diverted contact water and control the release of sediments down-gradient of the mine site by settling the sediments out of the discharged water. The sediment ponds for the Constancia site are identified as Main Sediment Pond, Road Sediment Pond No.1, Road Sediment Pond No.2, TMF Sediment Pond No.1, TMF Sediment Pond No. 2, Plant Construction Sediment Ponds (which work as sediment ponds just during construction phase), and Crusher Sediment Pond (construction phase as well).
Water Management Ponds - These structures typically deal with temporary water storage and may include both process and non-process water structures. The water management ponds for the Constancia mine are identified as the TMF Ponds (east and west), Plant Process Pond, Plant Contact Pond, Crusher Pond, Main Sediment Pond, and WRF Containment Pond.
Water Balance - Hudbay Peru performed a site-wide process water balance for the final design phase of the Constancia mine based on the current Mine Plan with the support of Piteau on the basis of the original KP design. Based on the analysis of the water balance model, enough water exists in the system for average climatological conditions to operate at the proposed mining and milling rates and no additional water sources are required for operation. As required, water accumulating in the sediment pond can be sent to the TMF for use in the process.

18.2. ON SITE CAMPS

Constancia is a remote site with restricted road access from the nearest significant town 1 ½ hours away. Currently there are two camps; Constancia; a permanent camp and Fortunia that was a pioneer camp and served as an overflow camp during the construction, Currently Fortunia is for practical purposes mainly being used for construction and specific contractors such as Heavy Civil Works, Security and Community-Relations. The camp facilities have been designed with reference to the IFC standards for camp construction for mining activities.

This camp was the previous construction camp. It's located at the north side of the mine. This permanent camp was developed to provide accommodations for operating personnel and necessary support personnel. The camp is divided into three pads. The first has the dining room, laundry, parking zone, warehouse and administrative offices. The second and the third pads have the housing modules with capacity for 2,335 people. The camp has a potable water treatment plant with 365 m3 per day capacity, a residual water treatment plant with 600 m3 per day capacity, a power house with three (standby) gen-sets of 500 kW each for emergency power and a gas storage capacity of 20,000 gallons. The main medical attention center is across this camp.

18.3. ELECTRIC POWER SUPPLY AND TRANSMISSION

18.3.1. POWER SUPPLY

Power supply for the Constancia mine is brought from the new 220kV transmission line from Tintaya to Constancia that was built for and owned by Hudbay but is operated and maintained by a third party.

18.3.2. POWER TRANSMISSION

This transmission line is connected to the Constancia electric substation that transforms 220kV to 23 kV. The switching area (220kV / 23kV) with its transmission line arrival cells, harmonic filters, power transformers, and an electrical control room are located within this substation.

The electric control room manages the 23kV system that provides energy to the entire mine. In case of emergency, there are three generators that provide energy in a semi-automatic way. The diesel powered engines generate 1,500 kW of electricity at 23 kV and are connected to 23 kV system.

18.3.3. CURRENT STATUS

While the Constancia Substation and the 220Kv power line was built, and is maintained and operated by a contractor, this infrastructure and the land on which it was built will remain under the ownership of Hudbay The historical average and peak power consumptions are shown in Figure 18-1.

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FIGURE 18-1: AVERAGE POWER CONSUMPTION

18.4. ROADS AND PORT

18.4.1. ROADS

The primary road to the site, consists of a 6,516 km compacted dirt road and 49 km of a bitumen surface roadway (National Route PE-3SG) from Coporaque to the Puente Bailey/Constancia (Figure 18-2). These roads (and bridges) have been upgraded, as necessary, to meet the needs of construction and life of mine use.

1) The access road from Coporaque to Constancia consists of:

2) Coporaque District Road running from Coporaque to Puente Kero (4.97 Km)

3) Provias National Road running from Puente Kero to Desvio Livitaca (48.7 Km)

4) Velille District Road running from Desvio Livitaca to Constancia (10.83 Km)

The section of road between Yauri and the Mine is classified as a National Class 3 road (less than 400 vehicles per day). In order to maintain the access road in good condition, it is necessary for Hudbay to intervene and maintain the road as necessary due to the heavy truck traffic - concentrate & consumables transport. The most difficult section is the National road passing over Huallyapacheta pass at an elevation of 4,689 masl. Due to heavy rains and traffic counts, the bitumen surface quickly failed soon after the start of operations and the road was converted into a compacted dirt roadway. This was done to improve transitability and lower costs to Hudbay whom has assumed 100% of the routine maintenance. Hudbay has engineered an alternative 7.4 km road that will connect Constancia traffic to the Las Bambas HHR within 20 km of Constancia. This will allow for safer passage of the Constancia heavy haul traffic on a roadway designed & constructed in anticipation of this type of traffic. The plan is for Hudbay to continue maintaining this current roadway until the connection to the Las Bambas HHR is permitted & constructed. Hudbay has a checkpoint at Yauri for control of trucks travelling to and from site.

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FIGURE 18-2: PERU - CONSTANCIA SITE TO PORT ACCESS ROADS

18.4.2. PORT

Copper concentrate is shipped from the Constancia mine via road (~485km) and arrives at the Matarani port by trucks (Figure 18-3).

This receiving system is designed to handle 350 tph, (2) trucks can dump simultaneously the concentrate in a covered building. The system includes, screen hoppers, feeders and conveyor system that feeds the concentrate into an existing stock pile stacking system inside the warehouse. This covered installation also includes a dust control system, inspection and sampling platforms and a truck wash system. This installation has been commissioned in October 2020 and is for the exclusive use of Hudbay. (Figure 18-4)

Pier C has been assigned to Hudbay and has a 75 Kt capacity. A chute from the shed feeds a tubular conveyor into a ship loader system. The same conveyor and ship loading equipment is shared with other copper concentrate exporters.

The ship loading feeder conveyor is tubular to avoid concentrate losses and to comply with Peruvian regulations. The design, installation and operation of the port is by TISUR. Hudbay pays for its use on a per metric tonne basis.

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FIGURE 18-3: MATARANI PORT

FIGURE 18-4: RECEIVING SYSTEM, WAREHOUSE C AND SHIP LOADER (AERIAL VIEW)

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19. MARKET STUDIES AND CONTRACTS

19.1. CONCENTRATE MARKETING

COPPER CONCENTRATES

Constancia copper concentrate is a clean, medium grade concentrate containing small gold and silver by-product credits. It is a highly desirable feedstock for copper smelters in China which is the most geographically appropriate freight destination but is also suitable for processing by smelters in Europe, India and South America. Constancia copper concentrate is an average quality concentrate containing gold and silver by- product credits which is delivered and sold directly to a variety of copper smelters in Asia and Europe as well as to internationally recognized trading companies. Up to 60% of sales are made pursuant to longer term frame contracts, which typically reference annual benchmark agreements between major concentrate producers and smelters for the purposes of fixing key terms such as treatment and refining charges. The balance of projected annual concentrate production has not been committed for sale for a number of reasons including: (1) uncommitted production may be a potential source of financing and (2) to provide flexibility in the event of fluctuations in annual production. Unsold production will be marketed and sold into the spot market each year at then-current market terms.

Constancia copper concentrate is trucked from the Constancia mine to the Matarani port facility in Southern Peru where it is loaded onto ocean vessels in lots of 10,000 or 20,000 mt for delivery to customers.

MOLYBDENUM CONCENTRATE

The Constancia metallurgical facility is equipped with a molybdenum flotation plant which produces a Mo concentrate which are delivered to customers in Asia and South America, in 1.5 mt bags in lots of approximately 20 mt. Roasting fees are generally estimated to be up to 13% of the contained value of molybdenum in concentrate.

19.2. CONTRACTS

Production at Constancia is subject to a precious metals streaming agreement with Wheaton consisting of 50% of payable gold and 100% of payable silver. Hudbay will receive cash payments equal to the lessor of the market price and US$400 per ounce for gold and $5.90 per ounce of silver, subject to a 1% annual escalation starting 3 years after the completion date in 2016. Under the terms of the stream, gold recovery for the purposes of calculating payable gold will be fixed at 55% for gold mined from Constancia and 70% for gold mined from Pampacancha.

For operations and the assured supply of goods and services, standard contracts are place as well as policies and procedures for contracts execution and closure.

Hudbay has a marketing division that is responsible for establishing and maintaining all marketing and sales administrations of concentrates and metals. As well, Hudbay conducts ongoing research of metal prices and sales terms as part of normal business and long range planning process to achieve market terms. Contract terms used in the Constancia financial evaluation are based on this research and the author has reviewed these results and they support the assumptions made in this technical report.

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20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

20.1. ENVIRONMENTAL

20.1.1. LEGAL FRAMEWORK

The mine was designed to comply with all legislation applicable to the development of mining projects in Peru including mines, roads, port and transmission lines. In addition, the mine complies with regulation regarding archaeological areas of significance, endangered and protected species as well as community relations and public disclosure programs. The main applicable regulation and international standards are detailed below:

Peruvian Political Constitution, 1993
Environmental General Law, 2005
Private Investment Growth Law, 1991
Environmental Impact for Works and Activities Evaluation Law, 1997
Sustainable Use of Natural Resources Organic Law, 1997
Environmental Impact Assessment National System Law, 2001
Law on the Right of Indigenous Peoples to Prior Consultation, 2011
Environmental Protection and Management Regulation for mining, Benefit, General Labor, Transportation and Mining Storage Activities, 2014
Water Resources Law, 2009
Water Quality National Standards, 2017
Air Quality National Standards, 2017
Noise Quality National Standards, 2003
Soil Environmental Quality Standards, 2017
Health Law, 1997
Environmental Management National System Law, 2004
Cultural Heritage Protection Law, 2004
Solid Waste Law, 2016
Forestry and Wildlife Protection Law, 2011
Mines Closure Law, 2003
Public Participation Regulations, 2008
Maximum Permissible Levels for Liquid Effluents from Mining - Mineral Processing Activities, 2010
IFC/World Bank Social and Environmental Performance Standards.
Towards Sustainable Mining (TSM) Protocols of the Mining Association of Canada (MAC)

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20.1.2. LOCAL AND REGIONAL CONTEXT

The Constancia mine is located in the districts of Chamaca, Livitaca, and Velille - all of which are in the Province of Chumbivilcas, Region of Cusco. It includes 5,452.41 ha of land purchased from private owners and the community of Chilloroya (District of Velille), as well as 256.50 ha of surface land granted in use by the community of Uchucarco (Chamaca District) and 473.7 ha by the community of Chilloroya (Livitaca District).

The area has a mining history that dates back 50 years to when Mitsui operated a copper-zinc mine in Katanga near the community of Uchucarco. Mitsui sold the mine to a Peruvian company which abandoned the mine in the late 1970s, leaving a legacy of environmental contamination. Tailings were discharged into the river basin, and the open pits and waste dumps were never reclaimed.

Near the community of Chilloroya, the local population has been traditionally engaged in artisanal and informal mining of superficial gold and copper. Concerns have been raised by both Hudbay and the community to Peruvian authorities regarding environmental and labour conditions related to this activity and levels of mercury found in nearby water bodies exceeding international standards for the protection of aquatic life.

20.1.3. ENVIRONMENTAL SETTING

The surface water quality of the Chilloroya River and its different tributaries (micro watersheds) are generally within the established national standards for water quality. However, due to natural factors (mineralization of the area, water erosion, water runoff) and anthropogenic factors (environmental liabilities, artisanal mining and community activities) some values are above national standards. These were identified during the Environmental Baseline.

The pH of the Chilloroya River and its main tributaries varies in the wet and dry season; during the wet season the pH varies from neutral to slightly alkaline (6.5 to 8.5) and in the dry season the pH varies from slightly alkaline to alkaline (8.0 to 9.5). In most of the monitoring stations, pH showed characteristics with alkaline tendency, even exceeding national standards. There were also some isolated values with slightly acidic characteristics, below the lower pH limit.

The total metals concentrations in ravines and the Chilloroya River are normally below the national standards However, some total metal concentrations exceeded national standards, these exceedances usually happen during the wet season. Total metals that registered high concentrations were copper, lead, manganese, molybdenum nickel, aluminium and zinc. On the other hand, isolated values of aluminium, mercury and iron also exceeded national standards. These exceedances would be associated with the sediments dragging, environmental liabilities and several activities of informal mineral extraction.

The microbiological parametres (Escherichia coli, fecal and total coliforms). along the Chilloroya River increased compared with the Environmental Baseline, reporting values above the national standards. This increase is directly related to livestock activities and non-controlled discharges of domestic wastewater treatment systems from Juan Velasco Alvarado village.

Regarding air quality, maximum concentrations of particles (PM-10 and PM-2.5), were measured in the towns of Uchucarco, Juan Velasco Alvarado and Urazana. All values were below national standards for 24 hours, except for some cases. Also, maximum concentrations of lead, arsenic and copper were recorded below national standards. Gas concentrations are mainly due to their own local emissions from light and heavy vehicle traffic surrounding the monitoring stations. None of the recorded values for gases exceeded the national standards. In general, air quality parametres, have been consistent with seasonality except for gases. Results showed higher concentrations during the dry season (August) than during the wet season (February and April).

342 different species of flora and 7 vegetation types were identified within the mining unit area (grassy brush, rock vegetation, very wet meadow, bog, rushes, bushes and high land vegetation). In total, there are 13 species that are within the conservation concern category in Peruvian legislation and 17 endemic species from Peru (4.97%). In addition, 256 species of fauna have been recorded including 21 species of mammals, 59 species of birds, 4 species of amphibians, 2 species reptiles and 170 species of arthropods. In total, 15 species are of conservation concern: four mammal species, 9 species of birds, and 2 species of amphibians, including only one endemic species, the mammal calomys sorellus (reddish evening mouse).

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A detailed hydrobiological evaluation of aquatic environments located within the mine area, was conducted. 6 species of fish were identified, 1 exotic (trout) and 5 native fish belonging to the "catfish" family, from these the trichomycterus rivolatus is in the category of "near threatened" as determined by the IUCN (2013).

An archaeological survey was completed in the areas directly affected by Constancia mine facilities and activities. A total of 46 archaeological sites were identified in the area of the mine. The certificate of non-existence of archaeological remains of significance (CIRA) was approved by the MC (Ministry of Culture) in September 2011. Hudbay has obtained a CIRA for (i) all the area comprised by the Constancia mine (excluding the archaeological sites as established in that same Certificate); (ii) for the route covered by the power transmission line (excluding the identify archaeological sites); and (iii) for an area for agricultural and forestry training. Later, as part of investigations into the Pampacancha sector, three archaeological sites and five cultural isolated elements were identified on the surface of Pampacancha. Following that, the corresponding Certificate of Nonexistence of Archaeological Remains was obtained in December 2015.

20.1.4. ENVIRONMENTAL CERTIFICATIONS

Constancia Mine Environmental Impact Assessment (ESIA) was approved by the Ministry of Energy and Mines (MINEM) on November 24, 2010, by Directoral Resolution (DR) No 390-2010-MEM-AAM, to conduct mining activities, with an average ore processing capacity of 55,000 tons per day for a period of 15 years.

The first amendment to the ESIA (EISA MOD I) was approved by MINEM on August 20, 2013, through DR. 309-2013-MEM/AAM, to increase the processing capacity to 81,900 tonnes per day, and to reduce the useful mine life to 11 years. The purpose of this amendment was to present changes stemming from engineering development. Additionally, Hudbay obtained the approval of an Environmental Technical Report by DR. No. 454-2013-MEM-AAM for the inclusion of auxiliary components necessary to complete construction activities on schedule.

The second amendment to the ESIA (EISA MOD II) was approved by MINEM on April 17, 2015, through DR. 168-2015-MEM-DGAAM, to increase reserves through the expansion of the Constancia pit, inclusion of the Pampacancha deposit, and the Waste Rock Facility (WRF) and the Tailings Management Facility (TMF) expansion, among others. A 17-year useful life of the mining operations was approved through this amendment.

Additionally, in 2015 and 2016, Hudbay obtained the approval of three Environmental Technical Reports (ITS) for the MOD II. The first one, approved by DR. No. 516-2015-MEM-AAM, included the relocation of auxiliary components. The second one, approved by DR. N° 63-2016-SENACE/DCA, included the modification, replacement and addition of auxiliary components, considering the optimization and usage change of existing components, and the use of areas enabled during mine construction. The third approved by DR. N° 120-2019-SENACE-PE/DEAR, included the modification of TMF design parameters and the type of material to be used in the dam construction, as well as other modifications to existing auxiliary components.

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20.1.5. ENVIRONMENTAL MANAGEMENT

Environmental and social impacts have been assessed and appropriate mitigation measures have been implemented. The EISA and amendments comply with national regulation and have adopted the Equator Principles and International Finance Corporation's (IFC) Performance Standards. The Towards Sustainable Mining (TSM) Standards of the Mining Association of Canada (MAC) are currently being implemented.

The hydrogeological report (Golder 2013) and its update, made by Amphos in 2019, which led to a more refined hydrogeological model and plant water balance, addressed three principal concerns related to the detailed design of the mine and plant water systems: (1) to confirm the relation between the pit dewatering area and reservoir with the springs in Cunahuiri area ; (2) determine the impact of slope stability on the pit wall design; and (3) to confirm availability of groundwater supply through life of mine.

Hudbay has engaged an Independent Project Review Board (IPRB) to conduct periodical reviews of the major earth structures on the mine, with regular visits to the site, as well as, for monitoring the progress of construction of the TMF. The plan for such review was developed during the construction stage, and is now in operation. Design reviews and site visits have been carried out at critical junctures throughout the course of construction and mine operations to observe the performance of the structure and development of subsequent raises to the dam and the impoundment.

The water management strategy includes the following:
Zero water discharge to the environment from the plant, and PAG waste rock facility.
Discharge from the TMF supernatant from the end 2023 in the rainy or dry season
Our prioritization of water sources for the plant is as follows:

1) Use of process water or water in contact with PAG infrastructures as pit sinkwater, WRF drainage water and TMF supernatant water in the process plant.

2) Use of water in contact with Non-PAG Infrastructure main sediment pond (MSP), and four other small sediment ponds. This type of water with sediments can be discharged in compliance with the Peruvian maximum allowable limit (LMP).

3) Use of pit dewatering groundwater for the process plant and dust control.

The MSP and the NC2 channels are used for water flows mitigation purposes, in particular they use water supply to mitigate any reduced river flow or water supply potential impacts of pit dewatering during dry season (27.4L/s), and the 10 L/s based on an agreement with the Chilloroya community.

The Constancia environmental monitoring program was updated in January 2018 Environmental Monitoring Plan (PLA-AMB-04: Plan de Monitoreo Ambiental). This plan covers monitoring of air, noise and vibration, surface water quality and flow, surface lake/reservoir levels, groundwater levels and quality, domestic water consumption and quality, effluent monitoring, topsoil, soil and sediment, biological (flora and fauna) monitoring, hydrobiological and meteorological data collection.

During the operation phase, Hudbay is maintaining a joint environmental monitoring committee with the communities of direct and indirect influence.

From a biodiversity stand point specific management plans for wetlands, flora and fauna are being executed; Hudbay has also developed a Biodiversity Action Plan (BAP), as generally anticipated and described in the EISA and as required in the MAC TSM protocol and IFC performance standard N° 6. The BAP considered the consolidation of commitments and effort regarding biodiversity.

During 2017, Hudbay obtained the certification of the ISO 14001: 2004 and OHSAS 18001: 2007 standards. In March 2017 the certification audit process was carried out in compliance with the International Organization for Standardization (ISO) ISO 14001: 2004 standard and the OHSAS 18001: 2007 standard, resulting in zero non-conformances and 15 observations that have already been corrected. In 2018, the External Audit for migration process to ISO 14001: 2015 and the first Follow-up Audit of compliance with OHSAS 18001: 2007 will be carried out.

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In 2020, the recertification of ISO 14001:2015 was obtained and the migration from the OHSAS 18001_2007 standard to ISO 45001:20018 of our Integrated Management System was achieved

20.1.6. MINE CLOSURE PLAN

The Constancia closure plan, describing the closure activities for the mine components aligned with the EISA, was approved through Resolution No. 286-2012-MEM-AAM dated September 4th, 2012.

Constancia's closure plan was updated to incorporate recent studies and technological changes that will reduce costs and provide financial guarantees that Hudbay must provide annually to the Peruvian government. The updated plan was approved by the Ministry of Energy and Mines in June of 2015 by Directorial Resolution No. 255-2015-MEM-AAM. This document updated measures of closure and post closure of the mine, according to the EISA MOD I and EISA MOD II.

The approved updated plan estimates a closure cost of $209.2M at the end of mine life. The cost guarantees provided to the Peruvian Government from Hudbay Peru SAC account for the discounted cost as calculated per Peruvian legislation for the reclamation of actual disturbed mining activity.

The Ministry of Energy and Mines started the evaluation of the Second Update of the Mine Closure Plan in September 2020. It includes new components, the closure schedule updating and amount guarantees updating

20.2. PERMITTING

Hudbay has obtained several permits for the construction and operation of the Constancia mine. All of these permits were supported by the environmental background provided by the ESIA and amendments.

The Constancia mine has secured all necessary permits and authorizations to conduct construction and operation activities at the mine. A summary of these permits is detailed below:
ESIA. - Besides being an environmental background and requirement for other permits, ESIA allows the construction and operation of certain facilities that do not have a specific construction and operation permit, for instance, camps and stockpiles. As mentioned in previous sections, an ESIA for the Constancia mine was approved in 2010, and to date it has two amendments and three Environmental Technical Reports (ITS).
Certificate of Non-Existence of Archeological Remains (CIRA).- On October 6th, 2011, Hudbay obtained a CIRA certificate for the area comprised by the Constancia mine, excluding identified archeological sites. In 2014 and 2015, Hudbay obtained five more CIRA certificates.
Beneficiation concession - A beneficiation concession allows its titleholder to process, purify and refine minerals by using chemical and/or physical procedures in process plants. In June 2012, Hudbay was granted the authorization to construct the process plant, including the TMF and water collection pond. Subsequent amendments have allowed Hudbay to raise the elevation of the TMF dam. After the construction of the process plant, Hudbay obtained the operation licence. This resolution allows Hudbay to operate the two lines of the process. Additionally, Hudbay has two Mining Technical Reports (ITM) approved in 2017 and 2018 regarding auxiliary components.
Underground Water License (pit dewatering). - In 2014, Hudbay obtained the underground water license. This permit authorizes the use of groundwater from wells installed for the pit dewatering. This license was later modified to authorize the use of up to 5,676,000 m3 of groundwater from the pit dewatering.
Surface Water License. - In 2015, Hudbay obtained a surface water license, which authorizes the use of surface water in the area of the mine for mining purposes, for an annual volume of up to 22,780,000 m3.
Mine Permit. - In 2012, Hudbay obtained authorization to start mining exploitation activities. This Resolution approved the mining plan and authorizes the exploitation activities in the Constancia and San Jose pits. Subsequent amendments were approved between 2013 and 2014 to authorize the development and subsequent expansions of the waste rock facility and auxiliary services.

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20.3. COMMUNITY RELATIONS

Between February and April 2012, Hudbay reached agreements with the neighboring communities of Uchucarco and Chilloroya for the land required for the Constancia mine. These agreements were validated in a meeting with support from two thirds of the community members. Both of these agreements have been recorded in the public land registry.

A compensation plan was prepared for the 36 landholding families from the lchuni area of Chilloroya, taking into account the needs of each family with respect to restructuring their working activities and adapting to relocation. The purpose of the plan was to cover all aspects a family needs to live a sustainable life. To this end, an approach was developed where the compensation plan established general criteria that should be adjusted in the individual negotiation process with each family, to reach an agreement that satisfied both parties. As the impact on each family was different, the Resettlement Action Plan (RAP) sought to ensure that each affected family was compensated fairly. This criterion was prepared in compliance with national laws and international provisions and standards in resettlement matters and in particular the IFC performance standards. As a result of the RAP implementation, the resettlement process was successfully completed in 2016.

A Community Relations Plan has been developed taking into account the following requirements:

Final determination of the Direct and Indirect Areas of Influence of the mine, based on the outcomes of the Social Impact Analysis as mentioned above.
The needs in the construction and operation phases, as determined in the Social Impact Analysis.
The State's requirements, as expressed in the Community Relations Guidelines of the MINEM and the Prior Commitment Act (DS-042-2003-EM).
The requirements of international financial institutions, taking into account as main references the Equator Principles, IFC Performance Standards, United Nations Environment Programme's APELL (Awareness and Preparedness for Emergencies at Local Level) for Mining, and supplementary social management standards.

Specific social programs were designed for the mitigation and prevention of identified impacts. Generally, these programs addressed the following key topics:

Communication and consultation

Participatory monitoring
Infrastructure
Social development
Economic development
Strategic development
Claims and dispute resolution.

Stakeholder mapping has been undertaken and is periodically updated, identifying the two principal stakeholder groups in the direct area of influence as the communities of Uchucarco and Chilloroya. There are three well-defined groups in each of the communities, namely:

Community assembly: people from the community.
Artisanal miners: from the community and who also possess land in the community (but do not possess title to the land).

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Youth Groups: (in the case of Chilloroya and Uchucarco, represented through the Youth Association) have higher education than the rest of the local residents and the experience of having lived in other cities; however, they own no land and possess no economic capital.

Benefits from the artisanal mining activities in Chilloroya are not distributed evenly across the population, but are primarily collected by those members of the population who possess the land where the illegal mining activities occur.

Uchucarco has a greater number and variety of organisations inside its community structure; some of which have duties that are independent from the Governing Board, these include: JAAS, the Committee of Water Users ("Comité de Regantes"), among others. In addition, Uchucarco is the most active community in the District of Chamaca, and has more relationships with district and private institutions.

In Chilloroya and Uchucarco, the political strength of the Governing Board is solid and the different associations within the community are not well organised. Although there is an array of organisations, many are focused on specific tasks and have little political presence, as in the case of women's and parents' organizations.

PAMPACANCHA

Since 2015 Hudbay has obtained the ESIA MOD II and the mine closure plan, which are the required environmental certifications to support mining activities in Pampacancha. In parallel, Hudbay started a negotiation process with the Chilloroya community to acquire rights over the Pampacancha surface. .

The Ministry of Energy and Mines authorized the start of the exploitation activities for the Pampacancha pit in Decembre 2020 (Directorial Resolution No. 790-2020-MINEM-DGM) This permit included the Prior Consultation process with the community of Chilloroya. Hudbay has also received approval for one of the three stages of Pampacancha pit drainage permits; the next stage will be evaluated by the National Water Authority in March 2021. Currently, the explosives permit has been under evaluation since January 2021.

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21. CAPITAL AND OPERATING COSTS

21.1. CAPITAL COSTS

The LOM sustaining capital is estimated to be $898 million (excluding capitalized stripping) at Constancia and Pampacancha while the project capital is estimated to be $51 million. All capital items are reported in real 2021 $USD.

The total includes capital required for major mining equipment acquisition, rebuilds, and major repair. The cost also includes site infrastructure expansion (tailings management facility, waste rock facility and others) and process plant infrastructure however they exclude all the cost related to mine closure. Project capital includes Pampacancha project development expenditures but excludes additional costs related to the remaining individual land user agreements.

The capital costs for Constancia are developed and revised on an annual basis as part of the budget cycle. The LOM capital plan is shown in Table 21-1.

TABLE 21-1: LIFE OF MINE CAPITAL EXPENDITURES (SUSTAINING AND GROWTH)

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21.2. OPERATING COSTS

The operating costs at Constancia are developed annually as part of the site budget process and based on past performance and planned operational improvements. All operating costs are reported in real 2021$ USD. The operating costs are divided in three categories: mining, milling and G&A. The LOM operating costs are shown in Table 21-2.

TABLE 21-2: ON SITE OPERATING COSTS

The mining operating costs are estimated for each activity, i.e. drilling, blasting, loading, haulage, auxiliary and indirect cost with the haulage activity having the highest impact on the total mining operation cost. For each activity, the volume of consumables and manpower requirements are calculated from the details of the life of mine plan. A breakdown of the mining operating costs by activity and component are shown in Table 21-3 and Table 21-4.

TABLE 21-3: OPERATING COSTS - OPEX - MINING/ACTIVITY

TABLE 21-4: OPERATING COSTS - OPEX - MINING/COMPONENTS

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The operating costs in the process plant are divided into seven main components, with power having the greatest impact on the process plant cost. The G&A cost includes all administrative areas at site as well the main office in Lima. A breakdown of milling and G&A costs are show in Table 21-5.

TABLE 21-5: OPERATING COSTS - OPEX - MILLING AND G&A

Waste stripping operating costs that provide future economic benefits are capitalized when strip ratios in a given year are above the average strip ratio for a total pit phase (capitalized stripping). Operating costs that are expected to be capitalized as stripping are summarized in Table 21-6.

TABLE 21-6: CAPITALIZED STRIPPING (TOTAL ESTIMATE AND UNIT COSTS)

Cash costs and sustaining cash costs per pound of copper are summarized in Table 21-7. Cash costs include mining, milling, G&A, off-site costs and TCRCs. Sustaining cash costs also include sustaining capital and royalties (but exclude Pampacancha project capex). Both cash costs and sustaining costs include the impact of capitalized stripping are reported net of by-product credits (calculated at reserves prices) including the impact of the precious metal streams.

TABLE 21-7: OPERATING COSTS AND SUSTAINING COSTS PER POUND

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22. ECONOMIC ANALYSIS

As Hudbay is a producing issuer, it has excluded information required by Item 22 of Form 43-101F1 as the Pampacancha expansion does not represent a material expansion of current production at the Constancia mine.

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23. ADJACENT PROPERTIES

There are no relevant adjacent properties to the Constancia and Pampacancha deposits.

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24. OTHER RELEVANT DATA AND INFORMATION

There is no other relevant data or information material to the Constancia mine that is necessary to make this Technical Report not misleading.

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25. CONCLUSIONS

The Constancia mine operation has been mining ore and waste since March 2014 and began commercial production in April 2015. Since that time the mine has run relatively uninterrupted and consistently achieved improved performance with regards to metallurgical recovery and throughput. The deposit represents what is considered by industry standards a low grade copper porphyry with the principal mineralized ore type being hypogene in nature. The plant has proven to produce a marketable clean concentrates for both molybdenum and copper.

As of December 31st, 2020 the mineral resource estimate is distributed in two open pit shells, Constancia and Pampacancha. The Hudbay internal peer reviews and validation processes including a thorough reconciliation between the reserve model and the past 2.5 years of operations have confirmed that the models are constructed utilizing industry best practices and sound QAQC principles.

The economic and design parametres modelled, assumed and stated provide for economic extraction of reserves from the stated resource.

The resource and reserve estimate meet the criteria outlined in industry best practices such as Canadian institute of Mining, Metallurgy and Petroleum Definition Standards for mineral Resources and Mineral Reserves (May 2014).

The production and compilation of this Technical Report was performed by the capable and professional management and staff of the corporate office of Hudbay and at the Constancia mine. The supervision, revision and approval of the assembly of this report is by the QP, Olivier Tavchandjian, P.Geo,, Vice-President Exploration and Geology of Hudbay.

The resulting mine plans account for feed restrictions as stipulated by the processing plant with regards to feed grade and ratios of zinc, iron, oxide and mineral type.
The current performance of the operation supports the position that the plant is capable of 90,000 tonnes per day average production, including consideration for maintenance.
Ongoing sustaining capital reported, sufficiently outlines the material capital requirements for mine fleet, plant maintenance and tailings expansion.
Separate project capital outlined adequately accounts for development of the Pampacancha satellite deposit.
Relevant mining permits and land title are in good standing excluding the required surface rights for the Pampacancha deposit exploitation.
The Pampacancha pit provides for higher grade than what remains available in the Constancia open pit.
Other than the risks described in this report, the general political and social risks associated with operating in Peru and the other risk factors described in Hudbay's most recent annual information form, there are no known significant risks and uncertainties that could reasonably be expected to materially affect the potential development of the mineral reserve and resource estimates in this report.

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26. RECOMMENDATIONS

Since 2017, Hudbay has consolidated a very prospective regional land package within trucking distance of the Constancia mining and milling infrastructures. Considerable progress has been made in concluding exploration agreements with the communities of Uchucarco, Anahuichi, Quehuincha and Collana-Velille resulting in 2 exploration agreements being signed and the obtention of a drill permit to test a skarn target at Quehuincha Norte. Hudbay will continue to very actively pursue this significant upside exploration potential that could prove that Constancia was only the first step in the creation of a large scale modern mining district in the region.

In 2020, the reserve pit has been extended by approximately 300m to the North to include the mineralization identified through the Constancia Norte drill program conducted in 2019-2020. A significant portion of the mineralization delineated during this program is still classified as inferred resources and would be amenable to reserve conversion with infill drilling. In addition, some of the mimeralization not included in the open pitable resource may be amenable to production through a small satellite underground operations that will be further investigated through trade-off studies in 2021.

Hudbay has a regional annual exploration operating budget of approximately $10 millions to cover the cost of these drilling program planned for Constancia Norte and other regional targets.

Thorough reconciliations from mineral reserve estimates to mill credited production will continue to be closely monitored in order to continue to validate the performance of the new reserve model.

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27. REFERENCES

Andes, Actualización del Estudio Geotécnico del Tajo Comnstancia (Spanish), Technical Report, Enero 2016
AMEC, Amec QC Evaluation, NI 43 101 Technical Report, Constancia Project, August 2011.
AMEC, Checking Samples, NI 43 101 Technical Report, Constancia Project, From 2006 to 2007.
AMEC, Constancia Mineral Resources Estimate, NI 43 101 Technical Report, Constancia Project, August 2011.
Amphos 21, Estudio Hidrogeológio Tajo Pampacancha Proyecto Expansion Constancia Estudio de Factyibilidad de Pampacancha (Spanish), (Spanish), Technical Report, September 2013
Amphos 21, Actualización del Manejo de Aguas. Proyecto Expansion Constancia Estudio de Factyibilidad de Pampacancha (Spanish), Technical Report, September 2013
Amphos 21, Estudio Hidrogeológico de la Ampliación del WRF. Proyecto Expansion Constancia Estudio de Factyibilidad de Pampacancha (Spanish), Technical Report, September 2013
Ausenco PSI, Constancia Tailings Transport System Detail Engineering, Lab Report Samples AdoB420 and AdoB421 (PSP5605-2000-INF-HI-006), April 2012.
C.H. Plenge & Cia, Metallurgical Investigation No.6447-6449 Norsemont Peru S.A.C. Mina Constancia Composites HY, SG, SK, March 2007.
C.H. Plenge &Cia, Metallurgical Investigation No.6637-6639 Norsemont Peru S.A.C. Mina Constancia Composites PSG, PHY, PSK, July 2008.
Golder Associates. Prefeasibility Hydrogeological Assessment Pampacancha Pit, Final Report, Constancia Project, August 2012.
Golder Associates. Pre-feasibility geotechnical study for Constancia Pit Expantion. Tehnical Report. March 2014.
Golder Associates. Pre -feasibility study for expanded capacity of the Tailings management Facility - Constancia Projetc. Technnical Report, March 2014.
Golder Associates. Pre-Feasibility study for expande capacity of waste dump (WRF) - Constancia Projet. Technical report, March 2014
JKTech Pty Ltd, SMC Test Report on Fifty Samples from Constancia Project Tested at A.R. MacPherson Consultants Ltd., SGS Lakefield Research Chile S.A., Santiago, Chile for Norsemont (JKTech Job No. 08004), May-June 2008
JKTech Pty Ltd, SMC Test Report on Two Samples from Constancia Project Tested at A.R. MacPherson Consultants Ltd., SGS Lakefield Research Chile S.A., Santiago, Chile for Norsemont (JKTech Job No. 08004) , June 2008
Knight Piesold, "Constancia Project Detailed Engineering Climatological Data Analysis Report", December 8, 2011, Prepared for Hudbay Perú S.A.C. by Knight Piésold Consultores S.A.
Knight Piésold, Open Pit Slope Design, Technical Report, Constancia Project, January 2013.
Knight Piésold, Prefeasibility Pit Slope Design for the Pampacancha Pit, Technical Report, Constancia Project, July 2012.
Knight Piésold, Pampacancha Preliminary Geochemical Assessment, Technical Report, Constancia Project, May 2012.
Itasca, Estudio Geotécnico del Tajo Pampacancha (spanish), Technical Report, september 2013.
Norsemont, Data Verification (Norsemont QC Protocol), NI 43 101 Technical Report, Constancia Project, From 2006 to 2012.
Pampacancha: SGS, Sample Preparation and Assaying, NI 43 101 Technical Report, Constancia Project, For Pampacancha 2011

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Pocock Industrial, Sample Characterization & PSA, Flocculant, Screening, Gravity Sedimentation and Pulp Rheology Studies, April 2012
SGS Canada Vancouver, An investigation into Flotation Behavior and QEMScan Characteristics of Constancia Samples, April 2005.
SGS Lakefield Research Chile, An investigation by QEMScan into the Mineralogical Characteristics of the Composite Samples from the Constancia Project (Project 4169-Q77), April 2009
SGS Lakefield Research Chile, Final Report: Third Report (Third Flotation Stage) Metallurgical Test Program (Project Code 4169) , September 2009.
SGS Lakefield Research Chile, June 2008. An investigation by QEMScan into the Mineralogical Characteristics of the Heads Products from the Constancia Project (Project 4169-Q33)
SGS Lakefield Research Chile, Pruebas de Molienda y Medición de Densidad, (Proyecto OL- 4081/Informe final), August 2007
SGS Lakefield Research Chile, Report: Molybdenum Separation Flotation Test Program (Project Code 4169) , September 2009
SGS Mineral Services Chile, Programa de Ensayos Piloto de Flotación con Mineral de Constancia (Proyecto 4169 - Informe #2), March 2009
"Norsemont Perú S.A.C. Proyecto Constancia Estudio de Impacto Social y Ambiental" (Spanish), March 2010, Prepared for Norsemont Perú S.A.C. by Knight Piésold Consultores S.A.
"Cuarta Modificación Del Estudio De Impacto Ambiental Semi-Detallado (Categoria II) Proyecto De Exploración Constancia" (Spanish), April 2012, Prepared for Hudbay Perú S.A.C. by Insideo S.A.C.
"Norsemont Perú S.A.C. Proyecto Constancia Plan de Cierre" (Closure Plan in Spanish), November 2011, Prepared for Norsemont Perú S.A.C. by Knight Piésold Consultores S.A.
"Norsemont Perú S.A.C. Proyecto de Exploración Constancia Tercera Modificación del Estudio de Impacto Ambiental Semidetallado Categoría II" (Spanish), December 2010, Prepared for Norsemont Perú S.A.C. by Knight Piésold Consultores S.A.
"Hudbay Peru S.A.C. Modificación del Estudio de Impacto Social y Ambiental del Proyecto Constancia" (Spanish), October 2012, Prepared for Hudbay Perú S.A.C. by Insideo S.A.C.
"Hudbay Peru S.A.C. Informe Técnico Sustentatorio de la Ampliación de Componentes del EIA del Proyecto Constancia" (Spanish), September 2013, Prepared for Hudbay Perú S.A.C. by Golder Associates S.A.C.
"Hudbay Peru S.A.C. Segunda Modificación del Estudio de Impacto Social y Ambiental del Proyecto Constancia - Ampliación Pampacancha" (Spanish), December 2013, Prepared for Hudbay Perú S.A.C. by Golder Associates S.A.C.
"Hudbay Peru S.A.C. Informe Técnico Sustentatorio para la Inclusión de Componentes Auxiliares en la U.M. Constancia" (Spanish), August 2015, Prepared for Hudbay Perú S.A.C. by Golder Associates S.A.C.
"Hudbay Peru S.A.C. Modificación del Plan de Cierre de Minas de la U.M. Constancia" (Spanish), February 2015, Prepared for Hudbay Perú S.A.C. by Golder Associates S.A.C.
"Hudbay Peru S.A.C. Segundo Informe Técnico Sustentatorio: Modificación, Sustitución y Adición de Componentes Auxiliares" (Spanish), Julio 2016, Prepared for Hudbay Perú S.A.C. by Golder Associates S.A.C.
"Titan-24 DC/IP/MT Survey Geophysical Report", August 2011, Prepared for Norsemont Perú S.A.C by Quantec Geoscience Ltd.

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Constancia Mine, Cuzco, Peru<br><br> <br>Form 43-101F1 Technical Report

OLIVIER TAVCHANDJIAN

CERTIFICATE OF QUALIFICATION

Re: NI 43-101 Technical Report, Constancia Mine, Cuzco, Peru March 29^th^ 2021

I, Olivier Tavchandjian, P. Geo., of Toronto, Ontario, do hereby certify that:

I am currently employed as Vice President, Exploration and Geology, with Hudbay Minerals Inc., 25 York Street, Suite 800, Toronto, Ontario, Canada, M5J 2V5.
I graduated from the Université du Québec à Chicoutimi with a Ph.D. in Mineral Resources in 1996.
I am a member in good standing with the Association of Professional Geoscientists of Ontario, member 0714.
I have practiced my profession continuously for over 30 years and have been involved in mineral exploration, project evaluation, resource and reserve evaluation, and technical support to mine operations in underground and open pit mines for base metals, precious metals and bulk commodities in the Americas, Europe, Asia and Africa.
I have read the definition of "qualified person" set out in National Instrument 43-101 ("NI 43-101") and certify that by reason of my education and affiliation with a professional association and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purpose of NI 43-101.
I have reviewed and approved the Summary of the Technical Report and I am responsible for the preparation of this Technical Report titled "NI 43-101 Technical Report, Constancia Mine, Cuzco, Peru" (the "Technical Report"), dated March 29^th^ 2021 and effective as of January 1^st^ 2021.
I last visited the Constancia Mine on September 9^th^ 2019 for a duration of four (4) days. I also visited the mine several times prior to that date.
As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
I am not independent of the Issuer. Since I am an employee of the Issuer, a producing issuer, I fall under subsection 5.3(3) of NI 43-101 where "a technical report required to be filed by a producing issuer is not required to be prepared by or under supervision of an independent qualified person".
I have been involved with the Constancia Mine, which is the subject of the Technical Report, continuously since March 2017.
I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with the instrument and form.
I consent to the public filing of the Technical Report with any stock exchange, securities commission or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Constancia Mine, Cuzco, Peru<br><br> <br>Form 43-101F1 Technical Report

Dated as of March 29^th^ 2021.

Original signed by:

/s/ Olivier Tavchandjian

Olivier Tavchandjian, P. Geo.

Hudbay Minerals Inc.: Exhibit 99.6 - Filed by newsfilecorp.com

CONSENT OF QUALIFIED PERSON

March 29, 2021

Dear Sir/Madam:

Re: Technical Report titled "NI 43-101 Technical Report, Constancia Mine, Cuzco, Peru" dated March 29, 2021, with an effective date of January 1, 2021

I, Olivier Tavchandjian, consent to the public filing of the technical report titled "NI 43-101 Technical Report, Constancia Mine, Cuzco, Peru" dated March 29, 2021, with an effective date of January 1, 2021 (the "Technical Report") by Hudbay Minerals Inc. ("Hudbay").

I certify that I have read the Disclosure Documents and that they each fairly and accurately represent the information in the Technical Report for which I am responsible.

/s/ Olivier Tavchandjian

Olivier Tavchandjian P. Geo.

Vice President, Exploration and Geology, Hudbay

Hudbay Minerals Inc 6-K, filed Mar 30, 2021

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