6-K - FFMGF - Equibles

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

UNDER THE SECURITIES EXCHANGE ACT OF 1934

For the month of December 2025

Commission File Number: 000-55607

First Mining Gold Corp.

| (Translation of registrant’s name into English) |

Suite 2070, 1188 West Georgia Street, Vancouver, B.C., V6E 4A2

(Address of principal executive office)

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 ☒

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): ☐

DOCUMENTS INCORPORATED BY REFERENCE

Exhibits 99.1 and 99.2 to this Report on Form 6-K are hereby incorporated by reference as Exhibits to the Registration Statement on Form F-10 of First Mining Gold Corp. (File No. 333-231801).

DOCUMENTS FILED AS PART OF THIS FORM 6-K

Exhibits	Description

| 99.1 | Technical Report |

2

SIGNATURES

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.

	First Mining Gold Corp.

| | Richard Huang |

| | VP, Corporate Development & Corporate Secretary |

3

EXHIBIT INDEX

Exhibits	Description

| 99.1 | Technical Report |

4

firstmining_ex991.htm EXHIBIT 99

Springpole Gold Project NI 43-101 Technical Report and Pre-Feasibility Study Ontario, Canada Effective Date: December 1, 2025 Report Date: December 19, 2025 Prepared for: First Mining Gold Corp. Suite 2070 - 1188 West Georgia Street Vancouver, British Columbia, Canada V6E 4A2 Prepared by: Ausenco Engineering Canada ULC 1050 West Pender Street, Suite 1200 Vancouver, British Columbia, Canada V6E 3S7 List of Qualified Persons: Tommaso Roberto Raponi, P. Eng., Ausenco Engineering Canada ULC Gordon Zurowski, P. Eng., AGP Mining Consultants Inc. Gilles Arseneau, P.Geo., SRK Consulting (Canada) Inc. David Bleiker, P. Eng., WSP Canada Inc. Daniel Russell, P.Geo., WSP Canada Inc.

CERTIFICATE OF QUALIFIED PERSON

Tommaso Roberto Raponi, P. Eng.

I, Tommaso Roberto Raponi, P. Eng., certify that:

1.	I am employed as a Senior Mineral Processing Specialist with Ausenco Engineering Canada ULC, (Ausenco), with an office address of 15th Floor, 11 King Street West, Toronto, Ontario, M5H 4C7.
2.	This certificate applies to the technical report titled “Springpole Gold Project NI 43-101 Technical Report and Pre-Feasibility Study, Ontario, Canada” that has an effective date of December 1, 2025 and a report date of December 19, 2025 (the “Technical Report).
3.	I graduated from University of Toronto with a Bachelor of Applied Science in Geological Engineering with specialization in Mineral Processing in 1984.
4.	I am a professional engineer registered with the Professional Engineers Ontario (No. 90225970), Engineers and Geoscientists British Columbia (No. 23536) and NWT and Nunavut Association of Professional Engineers and Geoscientists (No. L4508).
5.	I have practiced my profession continuously for 40 years with experience in development, design, operation and commissioning of mineral processing plants, focusing on gold projects, both domestic and internationally. My project design and development experience includes the generation of capital and operating costs for mineral processing plants and associated infrastructure and financial modelling of project economics.
6.	I have read the definition of “Qualified Person” set out in the National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections of the Technical Report that I am responsible for preparing.
7.	I have visited the project site on 29 July 2025.
8.	I am responsible for 1.1, 1.2, 1.9, 1.13, 1.14.1, 1.16, 1.17, 1.18, 2.1, 2.2, 2.3, 2.4.1, 2.5, 2.7, 12.3, 13, 17, 18.1, 18.2, 18.3, 18.4, 18.8, 18.9, 19, 21.1, 21.2.1, 21.2.2, 21.2.4, 21.2.5, 21.2.6, 21.2.7, 21.2.8, 21.2.10, 21.3.1, 21.3.3, 21.3.4, 21.3.5, 21.3.6, 22, 24, 25.1, 25.3, 25.7, 25.8.1, 25.9, 25.10, 25.11, 25.12.1.1, 25.12.1.3, 25.12.2.1, 25.12.2.3, 26.1, 26.4, 27 of the Technical Report.
9.	I am independent of First Mining Gold Corp.as independence is defined in Section 1.5 of NI 43-101.
10.	I have not been previously involved with the Springpole Gold Project. I was involved with reviewing sample selections on an earlier phase but for the purposes of this technical report, I have no prior involvement.

Page 1 of 2

11.

I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared in compliance with that Instrument. As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the Technical Report not misleading.

Dated: December 19, 2025

/signed/

Tommaso Roberto Raponi, P. Eng.

PEO Certificate of Authorization #: 100101607

Page 2 of 2

CERTIFICATE OF QUALIFIED PERSON

Gordon Zurowski, P. Eng.

I, Gordon Zurowski, P. Eng., certify that:

1.	I am employed as a Principal Mining Engineer with AGP Mining Consultants Inc., (AGP), with an office address of 132 Commerce Park Drive, Unit K #246, Barrie, Ontario, L4N 0Z7.
2.	This certificate applies to the technical report titled “Springpole Gold Project NI 43-101 Technical Report and Pre-Feasibility Study, Ontario, Canada” that has an effective date of December 1, 2025 and a report date of December 19, 2025 (the “Technical Report).
3.	I graduated from the University of Saskatchewan with a degree in B.Sc. Geological Engineering, 1989.
4.	I have practiced my profession continuously for over 30 years with experience in mineral reserve estimations and PEA, Pre-Feasibility and Feasibility studies in Canada, the United States, Central and South America, Europe, Asia, Africa, and Australia.
5.	I have practiced my profession continuously for over 30 years with experience in mineral reserve estimations and PEA, Pre-Feasibility and Feasibility studies in Canada, the United States, Central and South America, Europe, Asia, Africa, and Australia. As a result of my experience and qualifications, I am a Qualified Person as defined in NI 43–101.
6.	I have read the definition of “Qualified Person” set out in the National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections of the Technical Report that I am responsible for preparing.
7.	I have visited the project site with my last visit from August 30 to September 1, 2020.
8.	I am responsible for 1.11, 1.12, 15, 16, 21.2.3, 21.2.9, 21.3.2, 25.6, 25.12.1.2, 25.12.2.2, 26.5 and 27 of the Technical Report.
9.	I am independent of First Mining Gold Corp.as independence is defined in Section 1.5 of NI 43-101.
10.	I have been previously involved with the Springpole Gold Project as an author on the previous Pre-Feasibility Study title “NI 43-101 Technical Report and Pre-Feasibility Study on the Springpole Gold Project, Ontario, Canada” dated February 26, 2021.
11.	I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared in compliance with that Instrument. As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the Technical Report not misleading.

Dated: December 19, 2025

/signed/

Gordon Zurowski, P. Eng.

Page 1 of 1

CERTIFICATE OF QUALIFIED PERSON

Gilles Arseneau, Ph.D., P.Geo.

I, Gilles Arseneau, Ph.D., P.Geo., certify that:

1.	I am employed as an Associate Consultant with SRK Consulting (Canada) Inc., (SRK), with an office address of Suite 2600-320 Granville Street, Vancouver, British Columbia, V6C 1S9.
2.	This certificate applies to the technical report titled “Springpole Gold Project NI 43-101 Technical Report and Pre-Feasibility Study, Ontario, Canada” that has an effective date of December 1, 2025 and a report date of December 19, 2025 (the “Technical Report).
3.	I graduated from University of New Brunswick with a Bachelor of Science in Geology in 1979, from the University of Western Ontario with a Master of Science in Geology in 1984 and from the Colorado School of Mines with a Doctor of Philosophy in Geology in 1995.
4.	I am a professional geoscientist registered with the Engineers and Geoscientists British Columbia (No. 23474).
5.	I have practiced my profession continuously for 30 years with experience in exploration in North and South America and have extensive experience with Archean Gold deposits and porphyry hosted precious metal mineralization such as the Springpole Gold Project.
6.	I have read the definition of “Qualified Person” set out in the National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections of the Technical Report that I am responsible for preparing.
7.	I have visited the project site from February 10 to 12, 2012 and from August 8 to 9, 2022.
8.	I am responsible for Sections 1.3, 1.4, 1.5, 1.6, 1.7, 1.8, 1.10, 2.4.3, 2.6, 3, 4-11, 12.1, 14, 23, 25.2, 25.4, 25.5, 26.2, 26.3, 27 of the Technical Report.
9.	I am independent of First Mining Gold Corp. as independence is defined in Section 1.5 of NI 43-101.
10.	I have been previously involved with the Springpole Gold Project as a co-author of the previous technical report titled “NI 43-101 Technical Report and Pre-Feasibility Study on the Springpole Gold Project, Ontario, Canada” dated February 26, 2021, with an effective date of January 20, 2021.
11.	I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared in compliance with that Instrument. As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the Technical Report not misleading.

Dated: December 19, 2025

/signed/

Dr. Gilles Arseneau, Ph.D., P. Geo.

PEO Certificate of Authorization #: 11361849

Page 1 of 1

CERTIFICATE OF QUALIFIED PERSON

David Ernest Bleiker, P. Eng.

I, David Ernest Bleiker, P. Eng., certify that:

1.	I am employed as a Fellow Engineer with WSP Canada Inc., (WSP), with an office address of 6925 Century Ave, Mississauga, Ontario, L5N 7K2.
2.	This certificate applies to the technical report titled “Springpole Gold Project NI 43-101 Technical Report and Pre-Feasibility Study, Ontario, Canada” that has an effective date of December 1, 2025 and a report date of December 19, 2025 (the “Technical Report).
3.	I graduated from the University of Waterloo with a Bachelor of Applied Science in Civil Engineering in 1990 and a Masters of Applied Science in 1993.
4.	I am a professional engineer registered with the Professional Engineers Ontario (No. 90401936).
5.	I have practiced my profession continuously for 32 years with experience in mine water and waste management. I have designed and reviewed dry cover, water covers, completed water balances and undertaken geotechnical designs for mine water and mine waste facilities and lead multi-disciplinary teams for all of these.
6.	I have read the definition of “Qualified Person” set out in the National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections of the Technical Report that I am responsible for preparing.
7.	I have visited the project site on September 4, 2025.
8.	I am responsible for 1.14.2, 1.14.3, 2.4.2, 12.2, 18.5, 18.6, 18.7, 18.10, 25.8.2, 25.8.3, 25.12.1.4, 25.12.2.4, 26.6.1, 27 of the Technical Report.
9.	I am independent of First Mining Gold Corp. as independence is defined in Section 1.5 of NI 43-101.
10.	I have been previously involved with the Springpole Gold Project with respect to preparing the Environmental Impact Statement issued in October 2024.
11.	I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared in compliance with that Instrument.
12.	As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the Technical Report not misleading.

Dated: December 19, 2025

/signed/

David E. Bleiker, P. Eng.

Page 1 of 1

CERTIFICATE OF QUALIFIED PERSON

Daniel Russell, P.Geo.

I, Daniel Russel, P.Geo., certify that:

1.	I am employed as a Senior Manager, Geoscience with WSP Canada Inc., (WSP), with an office address of 6925 Century Ave, Mississauga, Ontario, L5N 7K2, Canada.
2.	This certificate applies to the technical report titled “Springpole Gold Project NI 43-101 Technical Report and Pre-Feasibility Study, Ontario, Canada” that has an effective date of December 1, 2025 and a report date of December 19, 2025 (the “Technical Report).
3.	I graduated from McMaster University with a Bachelor of Science (Honours) in Geology in 2000.
4.	I am a professional geoscientist registered with the Professional Geoscientists of Ontario (P.Geo. No. 1139).
5.	I have practiced my profession continuously for 25 years with experience obtained through roles as an exploration geoscientist in Ontario and Nunavut as well as 15 years’ experience as an environmental consultant specializing in permitting and approvals for mine development projects in Ontario.
6.	I have read the definition of “Qualified Person” set out in the National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections of the Technical Report that I am responsible for preparing.
7.	I have not visited the Springpole Gold Project site
8.	I am responsible for 1.15, 20, 25.12.1.5, 26.7, and 27 of this Technical Report.
9.	I am independent of First Mining Gold Corp. as independence is defined in Section 1.5 of NI 43-101.
10.	I have been previously involved with the Springpole Gold Project as author of the Assessment of Alternatives for Mine Waste Disposal, submitted with the Environmental Impact Statement.
11.	I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared in compliance with that Instrument.
12.	As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the Technical Report not misleading.

Dated: December 19, 2025

/signed/

Daniel F Russell, P.Geo.

PEO Certificate of Authorization #90054

Page 1 of 1

Important Notice

This report was prepared as National Instrument 43-101 Technical Report for First Mining Gold Corp. (First Mining*)* by Ausenco Engineering Canada ULC (Ausenco), AGP Mining Consultants Inc. (“AGP”), SRK Consulting (Canada) Inc. (“SRK”) and WSP Global Inc. (“WSP”), collectively the Report Authors. The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in the Report Authors’ services, based on i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by First Mining subject to terms and conditions of its contracts with each of the Report Authors. Except for the purposed legislated under Canadian provincial and territorial securities law, any other uses of this report by any third party are at that party’s sole risk.

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| NI 43-101 Technical Report and Pre-Feasibility Study | December 1, 2025 |

Table of Contents

1	Summary.			1

| | 1.1 | Introduction. | | 1 |

| | 1.2 | Terms of Reference. | | 1 |

| | 1.3 | Property Description and Location. | | 1 |

| | 1.4 | History. | | 2 |

| | 1.5 | Geology and Mineralization. | | 2 |

| | 1.6 | Exploration and Drilling. | | 3 |

| | 1.7 | Sampling Preparation, Analyses and Security. | | 3 |

| | 1.8 | Data Verification. | | 4 |

| | 1.9 | Mineral Processing and Metallurgical Testwork. | | 4 |

| | 1.10 | Mineral Resource Estimate. | | 6 |

| | 1.11 | Mineral Reserve Estimate. | | 7 |

| | 1.12 | Mining Methods. | | 8 |

| | 1.13 | Recovery Methods. | | 9 |

| | 1.14 | Project Infrastructure. | | 10 |

| | | 1.14.1 | Infrastructure Summary. | 10 |

| | | 1.14.2 | Dikes. | 12 |

| | | 1.14.3 | Co-Disposal Facility. | 12 |

| | 1.15 | Environmental and Social Setting. | | 13 |

| | | 1.15.1 | Environmental and Social Impacts. | 14 |

| | | 1.15.2 | Waste and Water Management. | 15 |

| | | 1.15.3 | Environmental Assessment Process. | 15 |

| | | 1.15.4 | Engagement and Consultation. | 15 |

| | | 1.15.5 | Mine Closure and Rehabilitation. | 15 |

| | 1.16 | Capital and Operating Cost. | | 16 |

| | | 1.16.1 | Capital Cost Estimate. | 16 |

| | | 1.16.2 | Operating Cost Estimate. | 17 |

| | 1.17 | Economic Analysis. | | 17 |

| | | 1.17.1 | Economic Summary. | 17 |

| | | 1.17.2 | Sensitivity Analysis. | 20 |

| | 1.18 | Conclusions and Recommendations. | | 20 |

| 2 | Introduction. | | | 21 |

| | 2.1 | Introduction. | | 21 |

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	2.2	Terms of Reference.		21

| | 2.3 | Qualified Persons. | | 21 |

| | 2.4 | Site Visits and Scope of Personal Inspection. | | 22 |

| | | 2.4.1 | Site Inspection for Tommaso Roberto Raponi. | 22 |

| | | 2.4.2 | Site Inspection for David Bleiker. | 22 |

| | | 2.4.3 | Site Inspection for Gilles Arseneau. | 22 |

| | 2.5 | Effective Dates. | | 22 |

| | 2.6 | Information Sources and References. | | 23 |

| | | 2.6.1 | Previous Technical Reports. | 23 |

| | 2.7 | Currency, Units, Abbreviations and Definitions. | | 23 |

| 3 | Reliance on Other Experts. | | | 28 |

| | 3.1 | Introduction. | | 28 |

| | 3.2 | Ownership, Mineral Tenure and Surface Rights. | | 28 |

| 4 | Property Description and Location. | | | 29 |

| | 4.1 | Summary. | | 29 |

| | 4.2 | Land Area. | | 30 |

| | 4.3 | Mineral Tenure. | | 33 |

| | | 4.3.1 | Jubilee Gold Claims and Royalty. | 33 |

| | | 4.3.2 | Leased Claims from R&S Legacy Inc. and Royalty. | 34 |

| | | 4.3.3 | Leased Claims from Springpole Group and Royalty. | 36 |

| | | 4.3.4 | Claims Leased from the Crown (Mining Claims). | 36 |

| | | 4.3.5 | Mining Leases. | 37 |

| | | 4.3.6 | Claim Maintenance. | 37 |

| | | 4.3.7 | Silver Stream with First Majestic Silver Corp. | 38 |

| | | 4.3.8 | Royalties Assumptions for Mine Planning and Economic Evaluation. | 39 |

| | 4.4 | Environmental Considerations. | | 39 |

| | 4.5 | Permitting Considerations. | | 39 |

| | 4.6 | Social License Considerations. | | 39 |

| | 4.7 | Project Risks and Uncertainties. | | 39 |

| 5 | Accessibility, Climate, Local Resources, Infrastructure and Physiography. | | | 40 |

| | 5.1 | Accessibility. | | 40 |

| | 5.2 | Local Resources and Infrastructure. | | 40 |

| | 5.3 | Climate and Physiography. | | 40 |

| 6 | History. | | | 41 |

| | 6.1 | Regional History of the Area. | | 41 |

| | 6.2 | History of Exploration on the Springpole Property. | | 42 |

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| NI 43-101 Technical Report and Pre-Feasibility Study | December 1, 2025 |

	6.3	Acquisition by First Mining.		44

| 7 | Geological Setting and Mineralization. | | | 45 |

| | 7.1 | Regional Geology. | | 45 |

| | 7.2 | Project Geology. | | 47 |

| | | 7.2.1 | Springpole Project Geology. | 47 |

| | | 7.2.2 | Birch-Uchi Regional Targets Geology. | 48 |

| | 7.3 | Structure. | | 50 |

| | 7.4 | Alteration. | | 50 |

| | 7.5 | Mineralization. | | 51 |

| | | 7.5.1 | Porphyry-style Mineralization. | 51 |

| | | 7.5.2 | Lode Gold Mineralization. | 52 |

| | | 7.5.3 | Gold Remobilization During Metamorphism. | 52 |

| 8 | Deposit Types. | | | 53 |

| | 8.1 | Deposit Model. | | 53 |

| | 8.2 | Depositional Environment. | | 53 |

| | | 8.2.1 | Springpole Genetic Model. | 54 |

| 9 | Exploration. | | | 56 |

| | 9.1 | Introduction. | | 56 |

| | 9.2 | Mapping and Sampling Programs, 2021 to 2025. | | 56 |

| | | 9.2.1 | 2021 Program. | 56 |

| | | 9.2.2 | 2022 Program. | 57 |

| | | 9.2.3 | 2023 Program. | 57 |

| | | 9.2.4 | 2024 Program. | 58 |

| | | 9.2.5 | 2025 Program. | 62 |

| | | 9.2.6 | Methodologies. | 67 |

| | 9.3 | Airborne Survey. | | 67 |

| 10 | Drilling. | | | 70 |

| | 10.1 | Introduction. | | 70 |

| | 10.2 | Historical Programs. | | 70 |

| | | 10.2.1 | Historical Drilling by Gold Canyon. | 70 |

| | | 10.2.2 | 2013 Geotechnical and Structural Program. | 75 |

| | 10.3 | First Mining Drilling, Springpole Project. | | 76 |

| | | 10.3.1 | Introduction. | 76 |

| | | 10.3.2 | 2016 Drill Program. | 76 |

| | | 10.3.3 | 2018 Drill Program. | 78 |

| | | 10.3.4 | 2020 Drill Program. | 79 |

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		10.3.5	2021 Drill Program.	80

| | | 10.3.6 | 2022 Drill Program. | 86 |

| | | 10.3.7 | 2024 Drill Program. | 92 |

| | 10.4 | First Mining Drilling, Regional Programs. | | 97 |

| | 10.5 | Core Resampling Program 2019 – 2020. | | 103 |

| | 10.6 | Drill Collar Surveying. | | 103 |

| | 10.7 | Oriented Core Surveying. | | 103 |

| | 10.8 | Down Hole Surveying. | | 104 |

| | | 10.8.1 | Gold Canyon Programs (Historical). | 104 |

| | | 10.8.2 | First Mining Programs. | 104 |

| | 10.9 | Drilling Pattern and Density. | | 107 |

| 11 | Sample Preparation, Analyses, and Security. | | | 108 |

| | 11.1 | Introduction. | | 108 |

| | 11.2 | Core Drilling and Sampling. | | 108 |

| | | 11.2.1 | Historical Programs (Gold Canyon). | 108 |

| | | 11.2.2 | First Mining Programs. | 109 |

| | 11.3 | Core Sampling, Handling and Chain-of-Custody. | | 110 |

| | | 11.3.1 | Historical Programs (Gold Canyon). | 110 |

| | | 11.3.2 | First Mining Programs. | 111 |

| | 11.4 | Sample Security. | | 111 |

| | 11.5 | Sample Preparation and Analytical Procedures. | | 112 |

| | | 11.5.1 | Historical Programs (Gold Canyon). | 112 |

| | | 11.5.2 | First Mining 2016, 2018 and 2020 (Metallurgical) Programs. | 113 |

| | | 11.5.3 | First Mining 2020 (Condemnation), 2021 and 2022 Programs. | 115 |

| | | 11.5.4 | First Mining 2023 and 2024 Programs. | 116 |

| | 11.6 | Bulk Density Data. | | 117 |

| | 11.7 | Quality Assurance and Quality Control Programs. | | 118 |

| | | 11.7.1 | Historical Programs (Gold Canyon). | 118 |

| | | 11.7.2 | First Mining Programs. | 120 |

| | 11.8 | QP Comments. | | 125 |

| 12 | Data Verification. | | | 126 |

| | 12.1 | Data Verification by Dr. Arseneau (QP). | | 126 |

| | | 12.1.1 | Verification performed by the QP. | 126 |

| | 12.2 | Data Verification by David Ernest Bleiker (QP). | | 128 |

| | | 12.2.1 | Geochemistry. | 128 |

| | | 12.2.2 | Hydrology. | 128 |

| | 12.3 | Verification by Tommaso Roberto Raponi. | | 128 |

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		12.3.1	QP’s Comments.	128

| 13 | Mineral Processing and Metallurgical Testing. | | | 129 |

| | 13.1 | Introduction. | | 129 |

| | 13.2 | Metallurgical Testwork. | | 129 |

| | | 13.2.1 | Historical Metallurgical Testwork Programs. | 129 |

| | | 13.2.2 | Current Metallurgical Testing Programs. | 131 |

| | 13.3 | Metallurgical Testing Programs. | | 135 |

| | | 13.3.1 | Sample Head Analyses. | 135 |

| | | 13.3.2 | Sample Mineralogy. | 137 |

| | | 13.3.3 | Comminution Testing. | 141 |

| | | 13.3.4 | Flotation. | 146 |

| | | 13.3.5 | Leaching. | 149 |

| | | 13.3.6 | Merrill Crowe Feed Solution Composition. | 157 |

| | | 13.3.7 | Cyanide Detoxification. | 158 |

| | | 13.3.8 | Tailings Dewatering. | 161 |

| | | 13.3.9 | Concentrate Dewatering. | 163 |

| | 13.4 | Metallurgical Variability. | | 164 |

| | 13.5 | Deleterious Elements. | | 166 |

| | 13.6 | Recovery Estimates. | | 166 |

| 14 | Mineral Resource Estimates. | | | 169 |

| | 14.1 | Introduction. | | 169 |

| | 14.2 | Resource Estimation Procedures. | | 169 |

| | 14.3 | Drill Hole Database. | | 170 |

| | 14.4 | Core Recovery. | | 170 |

| | 14.5 | Geological Domains. | | 171 |

| | 14.6 | Surface Topography. | | 173 |

| | 14.7 | Compositing. | | 174 |

| | 14.8 | Grade Capping. | | 175 |

| | 14.9 | Statistical Analysis and Variography. | | 176 |

| | 14.10 | Block Model and Grade Estimation. | | 177 |

| | | 14.10.1 | Gold and Silver Grade Models. | 178 |

| | | 14.10.2 | Bulk Density Model. | 178 |

| | | 14.10.3 | Total Sulphur and Arsenic Models. | 179 |

| | | 14.10.4 | Acid Base Accounting Model. | 182 |

| | 14.11 | Model Validation. | | 183 |

| | 14.12 | Mineral Resource Classification. | | 186 |

| | 14.13 | Mineral Resource Statement. | | 189 |

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		14.13.1	Note on Inferred Resources.	191

| | 14.14 | Grade Sensitivity Analysis. | | 192 |

| | 14.15 | Previous Mineral Resource Estimates. | | 194 |

| 15 | Mineral Reserve Estimates. | | | 196 |

| | 15.1 | Summary. | | 196 |

| | 15.2 | Introduction. | | 197 |

| | 15.3 | Mineral Reserves Statement. | | 197 |

| | 15.4 | Estimation Procedure. | | 198 |

| | 15.5 | Factors that May Affect the Mineral Reserve Estimates. | | 199 |

| 16 | Mining Methods. | | | 200 |

| | 16.1 | Summary. | | 200 |

| | 16.2 | Geotechnical and Hydrogeological Considerations used in Mine Planning. | | 201 |

| | | 16.2.1 | Geotechnical Considerations. | 201 |

| | | 16.2.2 | Hydrogeological Considerations. | 208 |

| | 16.3 | Geologic Model Importation. | | 209 |

| | 16.4 | Mining Loss and Dilution. | | 211 |

| | 16.5 | Pit Limit Analysis. | | 212 |

| | | 16.5.1 | Methodology. | 212 |

| | | 16.5.2 | Pit Limit Analysis Inputs and Parameters. | 213 |

| | | 16.5.3 | Pit Limit Analysis Results. | 216 |

| | 16.6 | Pit Design Parameters. | | 218 |

| | | 16.6.1 | Bench Design. | 218 |

| | | 16.6.2 | Haul Ramp Design. | 219 |

| | | 16.6.3 | Pit and Phase Selection. | 221 |

| | | 16.6.4 | Pit Design. | 222 |

| | | 16.6.5 | Phases. | 225 |

| | 16.7 | Waste Rock Storage Facility. | | 231 |

| | 16.8 | Life-of-Mine (LOM) Schedule. | | 232 |

| | 16.9 | LOM Plan Sequence. | | 238 |

| | | 16.9.1 | Blasting and Explosives. | 247 |

| | 16.10 | Mining Equipment. | | 247 |

| | 16.11 | Grade Control. | | 249 |

| | 16.12 | Quarry Area. | | 250 |

| 17 | Recovery Methods. | | | 251 |

| | 17.1 | Overview. | | 251 |

| | 17.2 | Process Flowsheet. | | 251 |

| | 17.3 | Plant Design. | | 254 |

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		17.3.1	Primary Crushing & Stockpiling.	255

| | | 17.3.2 | Grinding Circuit. | 255 |

| | | 17.3.3 | Flotation Circuit. | 256 |

| | | 17.3.4 | Flotation Concentrate Regrind Circuit. | 256 |

| | | 17.3.5 | Flotation Concentrate Dewatering. | 256 |

| | | 17.3.6 | Concentrate Leach. | 256 |

| | | 17.3.7 | Counter Current Decantation. | 257 |

| | | 17.3.8 | Flotation Tailings Dewatering. | 257 |

| | | 17.3.9 | Flotation Tailings Leach & Adsorption Circuit. | 257 |

| | | 17.3.10 | Cyanide Detoxification. | 258 |

| | | 17.3.11 | Carbon Acid Wash. | 258 |

| | | 17.3.12 | Carbon Stripping (Elution). | 259 |

| | | 17.3.13 | Carbon Reactivation. | 259 |

| | | 17.3.14 | Concentrate Merrill Crowe Circuit. | 260 |

| | | 17.3.15 | Tailings Merrill Crowe Circuit. | 260 |

| | | 17.3.16 | Gold Room. | 260 |

| | 17.4 | Product/Materials Handling. | | 261 |

| | 17.5 | Energy, Water and Process Materials Requirements. | | 261 |

| | | 17.5.1 | Process Materials Requirements. | 261 |

| | | 17.5.2 | Process Water Requirements. | 265 |

| | | 17.5.3 | Power Requirements. | 265 |

| 18 | Project Infrastructure. | | | 266 |

| | 18.1 | Introduction. | | 266 |

| | 18.2 | Access. | | 268 |

| | | 18.2.1 | Site Access. | 268 |

| | | 18.2.2 | On-site Roads. | 269 |

| | 18.3 | Built Infrastructure. | | 269 |

| | | 18.3.1 | Built Infrastructure – Surface Mine Buildings. | 270 |

| | | 18.3.2 | Built Infrastructure – Process Buildings. | 270 |

| | | 18.3.3 | Built Infrastructure – Infrastructure Buildings. | 271 |

| | | 18.3.4 | Accommodation. | 272 |

| | 18.4 | Stockpiles. | | 272 |

| | 18.5 | Tailings and Mine Rock Co-Disposal Facility. | | 272 |

| | | 18.5.1 | Background Information. | 273 |

| | | 18.5.2 | Design Basis, Requirements and Criteria. | 278 |

| | | 18.5.3 | CDF Ultimate and Starter Dams Sizing. | 283 |

| | | 18.5.4 | Tailings and Mine Rock Management. | 283 |

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		18.5.5	Water Management.	283

| | | 18.5.6 | Closure Concept. | 284 |

| | | 18.5.7 | CDF Dam Design and Analyses. | 285 |

| | | 18.5.8 | Instrumentation and Monitoring. | 286 |

| | | 18.5.9 | Construction Rock Mass Balance. | 289 |

| | 18.6 | Dikes. | | 289 |

| | | 18.6.1 | Design Basis. | 289 |

| | | 18.6.2 | Dike Design. | 290 |

| | | 18.6.3 | Hazard Classification. | 292 |

| | | 18.6.4 | Monitoring. | 292 |

| | | 18.6.5 | Environmental. | 293 |

| | 18.7 | Water Management. | | 293 |

| | | 18.7.1 | Water Management Plan. | 293 |

| | | 18.7.2 | Construction Water Management. | 294 |

| | | 18.7.3 | Operations Water Management. | 294 |

| | | 18.7.4 | Fresh Water Facilities – Operations. | 296 |

| | | 18.7.5 | Effluent Treatment Plant and Discharge. | 297 |

| | | 18.7.6 | Water Balance. | 298 |

| | 18.8 | Power and Electrical. | | 302 |

| | 18.9 | Fuel. | | 302 |

| | 18.10 | Water Supply and Management. | | 303 |

| 19 | Market Studies and Contracts. | | | 304 |

| | 19.1 | Market Studies. | | 304 |

| | 19.2 | Gold and Silver Price. | | 304 |

| | 19.3 | Contracts. | | 304 |

| | 19.4 | Comments on Market Studies and Contracts. | | 304 |

| 20 | Environmental Studies, Permitting, and Social or Community Impact. | | | 305 |

| | 20.1 | Environmental Considerations. | | 305 |

| | | 20.1.1 | Baseline and Supporting Studies. | 305 |

| | | 20.1.2 | Environmental Monitoring. | 311 |

| | | 20.1.3 | Waste Management. | 312 |

| | | 20.1.4 | Water Management. | 312 |

| | 20.2 | Permitting Considerations. | | 313 |

| | | 20.2.1 | Environmental Permits. | 313 |

| | 20.3 | Social Considerations. | | 317 |

| | | 20.3.1 | Social and Community Setting. | 317 |

| | | 20.3.2 | Engagement and Consultation. | 318 |

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	20.4	Closure and Reclamation Planning.		320

| | | 20.4.1 | Closure and Reclamation Plans. | 321 |

| | | 20.4.2 | Closure Cost Estimates. | 322 |

| 21 | Capital and Operating Costs. | | | 323 |

| | 21.1 | Introduction. | | 323 |

| | 21.2 | Capital Costs. | | 323 |

| | | 21.2.1 | Overview. | 323 |

| | | 21.2.2 | Basis of Estimate. | 325 |

| | | 21.2.3 | Mine Capital Costs. | 326 |

| | | 21.2.4 | Process Plant Capital Costs. | 332 |

| | | 21.2.5 | Infrastructure Capital Costs. | 340 |

| | | 21.2.6 | Indirect Capital Costs (WBS 6000). | 341 |

| | | 21.2.7 | Owner (Corporate) Capital Costs (WBS 8000). | 344 |

| | | 21.2.8 | Estimate Contingency (WBS 9000). | 344 |

| | | 21.2.9 | Sustaining Capital. | 345 |

| | | 21.2.10 | Closure Costs. | 345 |

| | 21.3 | Operating Costs. | | 346 |

| | | 21.3.1 | Overview. | 346 |

| | | 21.3.2 | Mine Operating Costs. | 346 |

| | | 21.3.3 | Process Operating Costs. | 356 |

| | | 21.3.4 | Basis of Estimate. | 357 |

| | | 21.3.5 | Infrastructure Operating Costs. | 358 |

| | | 21.3.6 | General and Administrative Operating Costs. | 358 |

| 22 | Economic Analysis. | | | 360 |

| | 22.1 | Forward-Looking Information. | | 360 |

| | 22.2 | Methodologies Used. | | 361 |

| | 22.3 | Financial Model Parameters. | | 361 |

| | | 22.3.1 | Assumptions. | 361 |

| | | 22.3.2 | Taxes. | 362 |

| | | 22.3.3 | Working Capital. | 362 |

| | | 22.3.4 | Closure Costs and Salvage Value. | 363 |

| | | 22.3.5 | Royalties. | 363 |

| | | 22.3.6 | Off-Site Costs. | 363 |

| | 22.4 | Economic Analysis. | | 363 |

| | 22.5 | Sensitivity Analysis. | | 369 |

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23	Adjacent Properties.			373

| 24 | Other Relevant Data and Information. | | | 375 |

| 25 | Interpretation and Conclusions. | | | 376 |

| | 25.1 | Introduction. | | 376 |

| | 25.2 | Exploration. | | 376 |

| | 25.3 | Metallurgical Testwork. | | 376 |

| | 25.4 | Quality Assurance and Quality Control. | | 378 |

| | 25.5 | Mineral Resource Estimate. | | 379 |

| | 25.6 | Mining Methods. | | 380 |

| | 25.7 | Recovery Methods. | | 381 |

| | 25.8 | Infrastructure. | | 382 |

| | | 25.8.1 | Summary. | 382 |

| | | 25.8.2 | Dikes. | 382 |

| | | 25.8.3 | Co-Disposal Facility. | 382 |

| | 25.9 | Capital Cost Estimate. | | 383 |

| | 25.10 | Operating Cost Estimate. | | 383 |

| | 25.11 | Economic Analysis. | | 383 |

| | 25.12 | Risks and Opportunities. | | 384 |

| | | 25.12.1 | Risks. | 384 |

| | | 25.12.2 | Opportunities. | 387 |

| 26 | Recommendations. | | | 389 |

| | 26.1 | Recommended Work Program. | | 389 |

| | 26.2 | Geology. | | 389 |

| | | 26.2.1 | Inferred Resource Upgrade. | 389 |

| | | 26.2.2 | Density Measurements. | 389 |

| | | 26.2.3 | Geology Estimated Budget. | 389 |

| | 26.3 | Exploration. | | 390 |

| | 26.4 | Metallurgical Testing. | | 390 |

| | 26.5 | Mining Methods. | | 391 |

| | | 26.5.1 | Mine Geotechnical. | 391 |

| | | 26.5.2 | In-Situ Characterization. | 391 |

| | | 26.5.3 | Laboratory Testing for Rock. | 392 |

| | | 26.5.4 | Laboratory Testing for the Sand-zone Material. | 392 |

| | | 26.5.5 | Mining Methods. | 392 |

| | 26.6 | Infrastructure. | | 393 |

| | | 26.6.1 | Dikes and CDF. | 393 |

| | 26.7 | Environmental. | | 395 |

| 27 | References. | | | 396 |

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| NI 43-101 Technical Report and Pre-Feasibility Study | December 1, 2025 |

List of Tables

Table 1‑1:	Mineral Resource Statement^1^ Inclusive of Mineral Reserves (effective September 30, 2025).	7

| Table 1‑2: | Springpole Mineral Reserve Estimate – November 13, 2025. | 7 |

| Table 1‑3: | Estimate Summary Level 1 Major Facility. | 16 |

| Table 1‑4: | Operating Cost Summary. | 17 |

| Table 1‑5: | Economic Analysis Summary. | 18 |

| Table 1‑6: | Summarized Recommended Work Program. | 20 |

| Table 2‑1: | Report Contributors. | 21 |

| Table 2‑2: | Previous Technical Reports. | 23 |

| Table 2‑3: | Abbreviations and Acronyms. | 24 |

| Table 2‑4: | Units of Measurement. | 26 |

| Table 6‑1: | Summary of Drilling at Springpole from 1986-2024. | 44 |

| Table 9‑1: | Selected Channel Sample Assays from the Challenger Target. | 60 |

| Table 9‑2: | Grab Sampling Highlights from 2022 to 2024 Field Programs. | 66 |

| Table 10‑1: | Gold Canyon Diamond Drilling Summary 2007 - 2013. | 70 |

| Table 10‑2: | Drill Hole Intercepts, Gold Canyon 2007 – 2013. | 73 |

| Table 10‑3: | 2013 Oriented-Core Drilling Program. | 76 |

| Table 10‑4: | Significant Intercepts from 2016 Metallurgical Drilling Program. | 78 |

| Table 10‑5: | Significant Intercepts from the 2020 Metallurgical Drilling Program. | 79 |

| Table 10‑6: | Summary of 2021 Drill Program. | 80 |

| Table 10‑7: | Significant Intercepts, 2021 Metallurgical Drilling Program. | 81 |

| Table 10‑8: | Significant Intercepts from 2021 Exploration Drilling (SW Pit Area). | 83 |

| Table 10‑9: | Intercepts from 2021 ARD Drill Program. | 86 |

| Table 10‑10: | Summary of 2022 Drill Program. | 86 |

| Table 10‑11: | Select Assay Highlights, 2022 Geotechnical Drilling Program. | 90 |

| Table 10‑12: | Select Assay Results, 2022 ARD Drilling Program. | 91 |

| Table 10‑13: | Summary of 2024 Drill Program. | 92 |

| Table 10‑14: | Assay Results, 2024 Exploration Drilling Program. | 93 |

| Table 10‑15: | Assay Results, 2022 Swain Drilling Program. | 97 |

| Table 10‑16: | Assay Results, 2023 Regional Drilling Program. | 99 |

| Table 10‑17: | Downhole Survey Summary, First Mining Drill Programs. | 106 |

| Table 11‑1: | SGS Multi-Element Analysis Method ICM14B – Detection Limits. | 113 |

| Table 11‑2: | Analytical Methods, First Mining Drill Programs. | 114 |

| Table 11‑3: | Summary of Assays with Possible Bias from 2022 Drilling. | 125 |

| Table 12‑1: | Assays from Duplicated Samples Collected During Site Visit. | 127 |

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Table 13‑1:	Metallurgical Testwork Summary.	130

| Table 13‑2: | Current Metallurgical Testwork Summary. | 131 |

| Table 13‑3: | Head Assays for Samples Tested in Program BL758. | 136 |

| Table 13‑4: | Head Assays for Samples Tested in Program BL1073. | 137 |

| Table 13‑5: | Mineral Proportions. | 137 |

| Table 13‑6: | Summary of Gold Deportment. | 138 |

| Table 13‑7: | Summary of Gold Carrier Minerals. | 139 |

| Table 13‑8: | Bulk Mineral Composition of Zone Composite Samples (Program BL758). | 139 |

| Table 13‑9: | Gold Deportment and Association in Zone Composite Samples (Program BL758). | 140 |

| Table 13‑10: | Breakage Data. | 143 |

| Table 13‑11: | Overall Flotation and Leach Extraction at Different Grind Sizes. | 146 |

| Table 13‑12: | Effect of Dispersant Addition on Mass Recovery to Concentrate. | 147 |

| Table 13‑13: | Flotation and Leach Extraction Test Results at Different Regrind Sizes (Program BL758). | 149 |

| Table 13‑14: | Flotation and Leach Extraction for Composites and Variability Samples (Program BL758). | 153 |

| Table 13‑15: | Flotation and Leach Extraction for Composites and Variability Samples (Program BL1073). | 156 |

| Table 13‑16: | Rougher Concentrate Leach Test Results at Varying Solids Content. | 156 |

| Table 13‑17: | Rougher Tailings Leach Test Results at Varying Conditions. | 157 |

| Table 13‑18: | Leach Solution Analysis. | 158 |

| Table 13‑19: | Production Composite Concentrate Cyanide Detoxification Test Results. | 158 |

| Table 13‑20: | Production Composite Tailings Cyanide Detoxification Test Results. | 159 |

| Table 13‑21: | MC3 Concentrate Cyanide Detoxification Test Results. | 159 |

| Table 13‑22: | MC4 Concentrate Cyanide Detoxification Test Results. | 160 |

| Table 13‑23: | MC3 Tailings Cyanide Detoxification Test Results. | 160 |

| Table 13‑24: | MC4 Tailings Cyanide Detoxification Test Results. | 160 |

| Table 13‑25: | Summary of Tailings Pressure Filtration Test Results. | 161 |

| Table 13‑26: | Summary of Static Settling Test Results. | 162 |

| Table 13‑27: | Summary of Pressure Filtration Test Results. | 162 |

| Table 13‑28: | Dynamic Settling Test Results for Rougher Concentrate. | 163 |

| Table 13‑29: | Parameters Used for Estimation of Plant Losses. | 168 |

| Table 14‑1: | Capping Levels for Springpole. | 175 |

| Table 14‑2: | Basic Univariate Statistical Information for Raw Uncapped Assay Data. | 176 |

| Table 14‑3: | Basic Univariate Statistical Information for 3 m Composites. | 176 |

| Table 14‑4: | Gold and Silver Spherical Correlogram Parameters by Domain. | 177 |

| Table 14‑5: | Block Model Setup Parameters. | 177 |

| Table 14‑6: | Search Parameters by Zone and Metal. | 178 |

| Table 14‑7: | Bulk Density of Un-estimated Blocks in the Model. | 179 |

| Table 14‑8: | Comparison of Lithology and Average Sulphur Contents. | 182 |

| Table 14‑9: | Average ABA data by Rock types. | 182 |

| Table 14‑10: | Search Parameters for ABA Parameters. | 183 |

| Table 14‑11: | Assumptions Considered for Conceptual Open Pit Resource Optimization. | 190 |

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Table 14‑12:	Mineral Resource Statement Inclusive of Mineral Reserves (effective September 30, 2025).	191

| Table 14‑13: | Indicated Block Model Quantities and Grade Estimates at Cut-off Grades. | 192 |

| Table 14‑14: | Inferred Block Model Quantities and Grade Estimates at Cut-off Grades. | 192 |

| Table 14‑15: | Previous Mineral Resource Statement of July 30, 2020 (at 0.30 g/t Cut-off). | 195 |

| Table 14‑16: | Current (2025) Mineral Resource at 0.30 g/t Cut-off. | 195 |

| Table 15‑1: | Springpole Mineral Reserve Estimate – November 13, 2025. | 196 |

| Table 15‑2: | Proven and Probable Mineral Reserves – Springpole Project (November 13, 2025). | 198 |

| Table 16‑1: | Overall Slope Angle Estimation. | 207 |

| Table 16‑2: | Open Pit Model Framework. | 209 |

| Table 16‑3: | Resource Model Item Descriptions. | 210 |

| Table 16‑4: | Open Pit Model Item Descriptions. | 210 |

| Table 16‑5: | Pit Limit Analysis Parameters. | 213 |

| Table 16‑6: | Parameters for Au Equivalent Grade. | 215 |

| Table 16‑7: | Nested Pit Shell Results. | 217 |

| Table 16‑8: | Haul Ramp Width Calculation. | 221 |

| Table 16‑9: | General Pit Statistics. | 224 |

| Table 16‑10: | Pit Inventory. | 224 |

| Table 16‑11: | Pit Inventory, Phase 1. | 225 |

| Table 16‑12: | Pit Inventory, Phase 2. | 226 |

| Table 16‑13: | Pit Inventory, Phase 3. | 227 |

| Table 16‑14: | Waste Material Storage Requirements. | 232 |

| Table 16‑15: | Mine Plan Ramp-up Schedule. | 233 |

| Table 16‑16: | LOM Production Schedule. | 235 |

| Table 17‑1: | Process Design Criteria. | 254 |

| Table 17‑2: | Average Annual Reagents Consumption. | 261 |

| Table 17‑3: | Average Annual Processing Consumables. | 262 |

| Table 18‑1: | Buildings List. | 269 |

| Table 18‑2: | Peak Ground Accelerations for Springpole Project Site. | 274 |

| Table 18‑3: | Annual Mine Waste Production Schedule. | 275 |

| Table 18‑4: | Physical and Mechanical Tailings Properties. | 275 |

| Table 18‑5: | CDF Waste Storage Volume Requirements. | 281 |

| Table 18‑6: | Summary of Environmental Design Flood Storage and Pumping Requirements. | 299 |

| Table 18‑7: | Average Annual Water Balance. | 299 |

| Table 20‑1: | Anticipated Federal Environment-related Approvals. | 314 |

| Table 20‑2: | Anticipated Provincial Environment-related Approvals. | 315 |

| Table 21‑1: | Estimate Summary Level 1 Major Facility. | 323 |

| Table 21‑2: | Initial Estimate by Major Discipline. | 324 |

| Table 21‑3: | CAPEX Contributor Definition. | 324 |

| Table 21‑4: | Mining Capital Costs Estimate Summary. | 327 |

| Table 21‑5: | Major Mine Equipment – Capital Cost, Full Finance Cost and Down Payment. | 329 |

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Table 21‑6:	Mine Equipment on Site.	330

| Table 21‑7: | Process Plant Initial Direct Costs. | 332 |

| Table 21‑8: | Onsite and Offsite infrastructure Initial Direct Costs. | 340 |

| Table 21‑9: | Third-Party (OHPL) Estimate Summary. | 341 |

| Table 21‑10: | Construction Camp Costs. | 343 |

| Table 21‑11: | Spares Costs. | 343 |

| Table 21‑12: | Contingency Summary. | 344 |

| Table 21‑13: | Estimate Summary Level 1 Sustaining Capital. | 345 |

| Table 21‑14: | Operating Cost Summary. | 346 |

| Table 21‑15: | Mine Staffing Requirements and Annual Employee Salaries. | 346 |

| Table 21‑16: | Hourly Manpower Requirements and Annual Salaries (Year 4). | 348 |

| Table 21‑17: | Major Equipment Operating Costs – No Labour (US$/h). | 350 |

| Table 21‑18: | Drill Pattern Specifications. | 351 |

| Table 21‑19: | Drill Productivity Criteria. | 351 |

| Table 21‑20: | Design Powder Factors. | 352 |

| Table 21‑21: | Loading Equipment Parameters. | 352 |

| Table 21‑22: | Support Equipment Operating Factors. | 353 |

| Table 21‑23: | Total Open Pit Mine Operating Cost Estimate - With Leasing. | 355 |

| Table 21‑24: | Adjusted Open Pit Mine Operating Cost Estimate - With Leasing. | 356 |

| Table 21‑25: | Process Plant Operating Cost Summary. | 356 |

| Table 21‑26: | G&A Cost Summary. | 359 |

| Table 22‑1: | Estimate of Working Capital. | 363 |

| Table 22‑2: | Smelter Term Summary. | 363 |

| Table 22‑3: | Economic Analysis Summary. | 364 |

| Table 22‑4: | Cashflow Statement on an Annualized Basis. | 366 |

| Table 22‑5: | Post-tax Sensitivity. | 370 |

| Table 22‑6: | Pre-tax Sensitivity. | 371 |

| Table 22‑7: | Sensitivity to Gold Price. | 372 |

| Table 26‑1: | Estimated Budget of Proposed Work. | 389 |

| Table 26‑2: | Exploration Program. | 390 |

| Table 26‑3: | Recommended Work Program. | 394 |

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| NI 43-101 Technical Report and Pre-Feasibility Study | December 1, 2025 |

List of Figures

Figure 1‑1:	Overall Site Layout Plan.	11

| Figure 1‑2: | Process Infrastructure Layout Plan. | 12 |

| Figure 1‑3: | Undiscounted, Unlevered, Free Cash Flow – Post Tax Basis. | 18 |

| Figure 4‑1: | Springpole Gold Project Location Map. | 30 |

| Figure 4‑2: | Mineral Tenure Plan. | 32 |

| Figure 6‑1: | Historical Exploration Summary Map, 1928 - 2001. | 42 |

| Figure 7‑1: | Regional Geology Plan. | 46 |

| Figure 7‑2: | Bedrock Geology and Target Locations, Gold Canyon Property. | 49 |

| Figure 9‑1: | Location Map of Main Targets Saddle and Challenger, including Charger Trend. | 59 |

| Figure 9‑2: | Rock Grab Sample Highlights from the 2023-2024 Programs in the Saddle-Challenger Target Areas. | 61 |

| Figure 9‑3: | Rock Sample Locations from the 2021 to 2025 Exploration Programs. | 63 |

| Figure 9‑4: | Soil Sample Locations from the 2021 to 2025 Exploration Programs. | 64 |

| Figure 9‑5: | Bedrock Mapping Locations from the 2025 Exploration Program. | 65 |

| Figure 9‑6: | Airborne Geophysics over Gold Canyon Property Area showing 2022 Survey Area. | 69 |

| Figure 10‑1: | Springpole Gold Project Historical Drill Hole Collar Location Map. | 72 |

| Figure 10‑2: | Location of first Mining 2016 to 2020 Drill Holes. | 77 |

| Figure 10‑3: | Location of 2021 and 2022 Drill Holes. | 87 |

| Figure 10‑4: | Location of First Mining 2024 Drill Holes. | 96 |

| Figure 10‑5: | Location of 2022 Drilling, Swain Target. | 98 |

| Figure 10‑6: | Location of 2023 Drilling at Saddle Target. | 101 |

| Figure 10‑7: | Location of 2023 Drilling at Horseshoe Target. | 102 |

| Figure 11‑1: | Shewhart Control Chart for 2022 Silver Standards. | 123 |

| Figure 11‑2: | XY Plot of Original ActLabs Silver Assays against AGAT Silver Re-assay Values. | 124 |

| Figure 13‑1: | Drill Holes Used for 2020 SGS Metallurgical Testing Programs. | 131 |

| Figure 13‑2: | Spatial Zones for 2022 and 2023 BaseMet Metallurgical Testing Programs. | 133 |

| Figure 13‑3: | Drill Hole and Interval Locations for Samples in the 2021-2023 BaseMet Program. | 134 |

| Figure 13‑4: | IsaMill Signature Plot. | 142 |

| Figure 13‑5: | SPI and DWi Relationship (Adapted from Bailey et al. 2008). | 143 |

| Figure 13‑6: | Example of Crumbly and Competent Drill Core. | 145 |

| Figure 13‑7: | Distribution to Rougher Flotation Concentrate - Variability Samples (Program BL758). | 148 |

| Figure 13‑8: | Leach Extraction for Flotation Concentrate and Tailings Composites and Variability Samples (Program BL758). | 150 |

| Figure 13‑9: | Concentrate Leach Gold Extraction vs Cumulative Time (Program BL758). | 151 |

| Figure 13‑10: | Tailings Leach Gold Extraction vs Cumulative Time (Program BL758). | 152 |

| Figure 13‑11: | Drill Hole and Interval Locations for Samples in the 2020 SGS and 2021-2023 BaseMet Programs. | 165 |

| Figure 13‑12: | Plant Recovery for Gold (includes 1.7% deduction for plant losses). | 167 |

| Figure 13‑13: | Plant Recovery for Silver (includes 2.2% deduction for plant losses). | 167 |

| Figure 14‑1: | Gold Grade versus Core Recovery Relationship. | 171 |

| Figure 14‑2: | Geological Domains for Springpole Gold Project. | 172 |

| Figure 14‑3: | Cross Section 1100NW Looking NW Showing Portage and East Extension Domains. | 173 |

| Figure 14‑4: | Histogram of Sample Lengths within Mineralized Domains. | 174 |

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Figure 14‑5:	Comparison of Sample Length and Average Gold Grade.	175

| Figure 14‑6: | Comparison of Historical S% by ICP with Re-assayed Total S%. | 180 |

| Figure 14‑7: | Contact Profile for Sulphur for the Portage Mineralized Zone. | 181 |

| Figure 14‑8: | Comparison of Gold Grades for Well-Informed Blocks. | 184 |

| Figure 14‑9: | Comparison of Silver Grades for Well-Informed Blocks. | 184 |

| Figure 14‑10: | Swath Plots for Gold re (a) the East Extension, (b) the Camp, and (c) the Portage Zone. | 185 |

| Figure 14‑11: | Swath Plot for Silver within the Portage Zone. | 186 |

| Figure 14‑12: | Average Distance of Drill Holes by Resource Class. | 187 |

| Figure 14‑13: | Number of Drill Holes by Resource Class. | 188 |

| Figure 14‑14: | Distance of the Nearest Drill Hole by Resource Class. | 189 |

| Figure 14‑15: | Grade-Tonnage Curves for the Springpole Indicated Mineral Resource. | 193 |

| Figure 14‑16: | Grade-Tonnage Curves for the Springpole Inferred Mineral Resource. | 194 |

| Figure 16‑1: | Oblique View Looking Eastward showing PFS Pit and Modelled Sand-zone. | 202 |

| Figure 16‑2: | Material Sampled for 2022 Characterization of Sand-zone. | 203 |

| Figure 16‑3: | (SW-SM) Well Graded Sand with Silt and Gravel. | 204 |

| Figure 16‑4: | Vertical Section of Drill Hole Locations and Zone of Increased Geotechnical Risk. | 206 |

| Figure 16‑5: | Geotechnical Sectors. | 208 |

| Figure 16‑6: | Cut-off Grade Analysis. | 215 |

| Figure 16‑7: | Grade-Tonnage Curve, Mining Block Model. | 216 |

| Figure 16‑8: | Pit-by-Pit Graph. | 218 |

| Figure 16‑9: | Pit Slope Design Terminology. | 219 |

| Figure 16‑10: | Haul Ramp Design Section, Double Lane. | 220 |

| Figure 16‑11: | Haul Ramp Design Section, Single Lane. | 220 |

| Figure 16‑12: | Ultimate Pit Design, Plan View. | 222 |

| Figure 16‑13: | Ultimate Pit Design, Section Views. | 223 |

| Figure 16‑14: | Phase Design Section Views. | 228 |

| Figure 16‑15: | Phase 1 Design, Plan View. | 229 |

| Figure 16‑16: | Phase 2 Design, Plan View. | 230 |

| Figure 16‑17: | Mill Feed Tonnage and Au Grade by Year Graph. | 236 |

| Figure 16‑18: | Total Material Mined and Mining Production Rate by Year Graph. | 236 |

| Figure 16‑19: | Material Mined Production Graph. | 237 |

| Figure 16‑20: | Material Mined by Phase by Year Graph. | 237 |

| Figure 16‑21: | End-of-Year Progression Plan, Year -1. | 238 |

| Figure 16‑22: | End-of-Year Progression Plan, Year 1. | 239 |

| Figure 16‑23: | End-of-Year Progression Plan, Year 2. | 240 |

| Figure 16‑24: | End-of-Year Progression Plan, Year 3. | 241 |

| Figure 16‑25: | End-of-Year Progression Plan, Year 4. | 242 |

| Figure 16‑26: | End-of-Year Progression Plan, Year 5. | 243 |

| Figure 16‑27: | End-of-Year Progression Plan, Year 6. | 244 |

| Figure 16‑28: | End-of-Year Progression Plan, Year 7. | 245 |

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Figure 16‑29:	End-of-Year Progression Plan, Year 8 (End of Mining).	246

| Figure 16‑30: | Major Production Equipment, by type, by year. | 248 |

| Figure 16‑31: | Support Equipment, by type, by year. | 249 |

| Figure 17‑1: | Process Flowsheet. | 253 |

| Figure 18‑1: | Overall Site Layout Plan. | 267 |

| Figure 18‑2: | Process Infrastructure Layout Plan. | 268 |

| Figure 18‑3: | Springpole Site Layout with Co-Disposal Facility. | 273 |

| Figure 18‑4: | Boreholes, Monitoring Wells and Test Pits Locations. | 278 |

| Figure 18‑5: | CDF Layout with Offset Constraints. | 280 |

| Figure 18‑6: | Proposed Ultimate CDF Configuration. | 287 |

| Figure 18‑7: | Proposed CDF – Profiles and Sections. | 288 |

| Figure 18‑8: | Location of West and East Dikes, and Limits of Bay Dewatering. | 290 |

| Figure 18‑9: | Dike Typical Section. | 292 |

| Figure 18‑10: | Watersheds and Flow Concept. | 300 |

| Figure 18‑11: | Water Management During Mining. | 301 |

| Figure 22‑1: | Undiscounted, Unlevered, Free Cash Flow – Post Tax Basis. | 364 |

| Figure 22‑2: | Sensitivity Analysis. | 369 |

| Figure 23‑1: | Adjacent Properties Map. | 374 |

List of Appendices

Appendix A	Springpole Gold Project Mineral Tenure Map.	400

| Appendix B | List of the Patents, Mining Leases and Mining Claims. | 401 |

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1 SUMMARY

1.1 Introduction

This report was prepared by Ausenco Engineering Canada ULC. (“Ausenco”), AGP Mining Consultants Inc. (“AGP”), SRK Consulting (Canada) Inc. (“SRK”) and WSP Global Inc. (“WSP”) for First Mining Gold Corp. (“First Mining” or the “Company”) to summarize the results of the Pre-Feasibility Study (“PFS”) of the Springpole Gold project (the “Project” or “Springpole”).

1.2 Terms of Reference

The technical report was prepared in accordance with the Canadian disclosure requirements of National Instrument 43-101 (“NI 43-101”) and Form 43-101 F1, and is prepared using the Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) Definition Standards for mineral resources and mineral reserves (CIM Definition Standards, 2014) and the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines (CIM Best Practice Guidelines, 2019).

The report supports disclosures by First Mining in a news release dated November 18, 2025 entitled “First Mining Announces Updated Pre-Feasibility Study for the Springpole Gold Project, Ontario.”

All measurement units used in this report are SI units unless otherwise noted. Currency is expressed in Canadian dollars (C$) or United States Dollars (US$) as noted.

1.3 Property Description and Location

The Springpole Gold project is wholly-owned and controlled by First Mining through its subsidiary company Gold Canyon Resources (“Gold Canyon”) and comprises 30 patented mining claims, 280 contiguous mining claims and 13 mining leases totalling an area of 41,952 hectares (“ha”). Additional claims adjacent to the project area within the Birch-Uchi Greenstone Belt have been acquired by Gold Canyon between 2021 and 2025 (the “Birch-Uchi Tenure”). This additional tenure comprises a further 562 mining claims totalling 15,895 ha which are 100% owned by Gold Canyon, plus a further 82 mining claims (1,656 ha) which are 70% owned by Gold Canyon and 30% owned by Whitefish Exploration Inc.

During late spring, summer, and early fall, the Springpole Gold project is accessible by floatplane direct to Springpole Lake or Birch Lake from the communities of Red Lake, Ear Falls and Sioux Lookout, Ontario. During breakup in spring and freeze-up in fall, access to the Project is by helicopter. Additional winter access may be available via temporary airstrips cleared on nearby frozen lakes. The closest road access at present is approximately 18 kilometres (“km”) away at the extension of the Wenesaga forestry access road.

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1.4 History

Gold exploration was carried out during two main periods, one during the 1920s to 1940s, and a second period from 1985 to the present.

Between 1933 and 1936, the Windigokan Sturgeon Mining Syndicate conducted extensive trenching and prospecting, including 10 short holes totalling 458.5 metres (“m”). The claims were then transferred to Springpole Mines Ltd. who carried out limited trenching and prospecting in 1945.

From 1986 through 1989, GFCM completed 118 diamond drill holes in seven drill phases totalling 38,349 m. In addition, during 1986 and 1987, approximately 116,119 square metres (“m^2^”) of mechanical stripping was carried out.

Late in 1989, GFCM entered into a 50/50 joint venture with the combined interests of Noranda and Akiko-Lori Resources Ltd. (Akiko-Lori).

From 1989 through 1992, Noranda conducted an induced polarization (IP) survey over the central portion of the Portage zone and tested the property with eighteen core holes totalling 5,993 m.

During 1992 to 1994, Akiko-Lori/Akiko Gold completed an additional 15 diamond drill holes at the Project totalling 5,154 m.

During 1995 and 1996, Santa Fe drilled an additional 69 core holes totalling 15,085 m on the Springpole Gold project. After Santa Fe’s departure, Gold Canyon Resources Inc. (Gold Canyon) continued exploration at the Springpole Gold project in 1997 and 1998 with another 52 core holes totalling 5,643 m.

Paso Rico Resources Ltd. (Paso Rico) had an option to earn an interest in the Project and, in 1998 and 1999, conducted with Gold Canyon a lake bottom sediment sampling program in several areas of Springpole Lake, as well as 12 diamond drill holes totalling 2,779 m.

Between 2004 and 2013, diamond drilling programs were conducted on the property by Gold Canyon under its previous management. A total of 371 holes were completed during this period, over 108,932 m.

On November 13, 2015, First Mining (which was called First Mining Finance Corp. at the time) completed the acquisition of Gold Canyon, and as a result, acquired the Springpole Gold project.

1.5 Geology and Mineralization

A polyphase alkali, trachyte intrusion displaying autolithic breccia textures lies at the heart of the Springpole Gold project. The intrusion is comprised of a system of multiple phases of trachyte believed to be part of the roof zone of a larger syenite intrusion, as fragments displaying phaneritic textures were observed from deeper drill cores in the southeast portion of the Portage zone. Early intrusive phases consist of megacrystic feldspar phenocrysts of albite and orthoclase feldspar in an aphanitic groundmass. Successive phases show progressively finer-grained porphyritic texture while the final intrusive phases are aphanitic. Pervasive alteration and metamorphism have reduced the original porphyry intrusion to a complex alteration assemblage dominated by sericite, biotite, pyrite, calcite/dolomite, and quartz. Within the country rocks to the north and east are trachyte and lamprophyre dikes and sills that source from the trachyte or syenite-porphyry intrusive system.

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The main intrusive complex appears to contain many of the characteristics of alkaline, porphyry-style mineralization associated with diatreme breccias (e.g., Cripple Creek, Colorado). This style of mineralization is characterized by the Portage zone and portions of the East Extension zone where mineralization is hosted by diatreme breccia in aphanitic trachyte. Ductile shearing and brittle faulting may have played a role in redistributing structurally controlled blocks of the mineralized rock. Diamond drilling has revealed a more complex alteration with broader, intense zones of potassic alteration replacing the original rock mass with biotite and pyrite. In the core area of the deposit where fine-grained, disseminated gold mineralization occurs with biotite, the primary potassic alteration mineral, gold, displays a good correlation with potassium/rubidium.

1.6 Exploration and Drilling

First Mining completed diamond drilling programs at the Project in 2016, 2018, 2020, 2021, 2022 and 2024, totaling 212 holes over 28,086 m. Of this drilling, 18 holes (6,658 m) were exploration holes, 17 holes (5,527 m) were for metallurgical sample collection, and the remainder was for other site investigation purposes such as geotechnical, hydrogeological, condemnation, and sample collection for metal leaching/acid rock drainage testwork programs.

Exploration drilling in 2022 in the southwest area of the deposit identified additional mineralization (the “SW Extension Zone”) highlighting extension of the deposit continuity outside the mineral resource area as defined in the 2021 PFS.

The most recent exploration drilling at the Project area was in 2024, where five holes totalling 2,293 m were completed in a Phase 1 drilling campaign which focused on a 150 m strike area at the Southeast Extension target located at the southeastern boundary of the current mineral resource and proposed open pit wall design. The Southeast Extension target remains open along strike towards the south and southeast of the main Portage zone and has the potential to add meaningful mineralization extension or additional zones within or near the current PFS proposed open pit shell.

First Mining has also completed several other exploration programs between 2021 and 2025 over the Springpole Project area and its adjacent mineral tenure in the Birch-Uchi Greenstone Belt. These activities consisted of diamond drilling and prospecting campaigns on regional targets, an airborne electromagnetic and magnetic survey, and detailed geological mapping and sampling over select targets on the Project area.

1.7 Sampling Preparation, Analyses and Security

For Gold Canyon’s 2010 and 2011 drill programs, all drill core intervals were sampled using sample intervals of 1 m. During the 2013 drilling program, Gold Canyon changed its standard sample length from 1 m to 2 m lengths. However, in zones of poor recovery, 1.5 m or 3 m samples were sometimes collected. Samples over the standard sample length were typically half core samples and whole core was generally only taken in intervals of poor core recovery across the sampled interval. For First Mining’s drill programs, the minimum sampling length for core samples was generally 0.3 m and maximum length was 1.5 m. Samples were selected taking lithological boundaries into account as well as reducing sample widths over areas with increased mineralization. For all programs, sampling marks were made on the core and sample tickets were stapled into the core boxes at the beginning of each sample interval. Quality control samples were inserted into the sample stream.

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In the opinion of the QP, the sampling preparation, security and analytical procedures used in the drill programs conducted by Gold Canyon for gold analyses are acceptable but not fully consistent with generally accepted industry best practices because of the lack of standard reference material for silver for the earlier drill campaigns. However, because of the relatively low economic value of silver, the QP concludes that the assay data is adequate for use in resource estimation. First Mining has an established QA/QC protocol for the acceptance of assay batches with respect to the performance of standard reference material, duplicates, and blanks. The sampling preparation, security and analytical procedures used by First Mining in all their drilling programs are acceptable for inclusion in a Pre-Feasibility Study and are consistent with generally accepted industry best practices.

1.8 Data Verification

As part of the mineral resource estimation process, Dr. Arseneau reviewed the QA/QC data collected by Gold Canyon, reviewed the procedures in place to assure assay data quality, and verified the assay database against original assay certificates provided directly to SRK by SGS Red Lake, the assay laboratory. A total of 53,431 gold assays, 46% of the assay data, were checked against original assay certificates. No significant database errors were identified. About 143 minor rounding errors were observed. None of the rounding errors are deemed material or of any significance to the mineral resource estimate presented in this report.

1.9 Mineral Processing and Metallurgical Testwork

The Springpole deposit has been the subject of several metallurgical testwork programs and previous studies.

Earlier programs from 1989 to 2019 investigated whole ore leach and flotation followed by concentrate regrind and leach as different processing options. The 2021 PFS testwork focused on investigating flotation and leach processing. The testwork indicated that a whole ore leaching flowsheet would result in poor leach extractions. Instead, separate leaching of flotation concentrate and flotation tailings were selected as the recommended flowsheet.

Gold occurs as fine-grained and occurs primarily as telluride minerals as well as in silicates. The telluride associated gold recovers well to flotation concentrate and further recovery improvement is seen when regrinding concentrate for improved liberation of fine-grained gold. Most of the remaining gold is associated with silicates and is recoverable by cyanidation.

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For the 2025 PFS, the flowsheet featured split tailings streams for flotation concentrate and flotation tailings to manage sulphur and potential acid generating (PAG) rock. A Merrill Crowe circuit was selected for precious metal recovery from the concentrate leach solution because of the mineralized material’s high silver to gold ratio and variable flotation and leach extraction.

The 2021-2023 testwork program focused on optimizing the flotation and leach process flowsheet.

Head assays confirm the presence of deleterious elements mercury and arsenic. Mercury is present at concentrations that warrant mercury control. Some areas of elevated copper and zinc concentrations are seen which may result in higher than average cyanide consumption.

Ore is considered highly variable in terms of breakage characteristics, confirming observations made in the previous 2021 PFS. Mass and metal recovery to flotation concentrate was highly variable, as was leach extraction for reground flotation concentrate and tailings.

Optimal rougher concentrate leach conditions were determined to be a target regrind P80 size of 17 micrometres (“µm”), a slurry density of 45% solids with a pre-aeration period of eight hours (“h”). The optimal rougher tailings leach conditions were set at 53% solids. Leaching at a higher solids content for both the concentrate and tailings streams showed no adverse impact on leach extraction.

Concentrate leach is essentially complete after eight hours and tailings leach after 10 h. Average sodium cyanide consumption from the production and master composite samples was 0.12 kilogram/tonne (“kg/t”) of tailings for the tailings leach circuit and 2.87 kg/t of concentrate for the concentrate leach circuit. Average lime consumption was 1.33 kg/t of tailings and 3.98 kg/t of concentrate.

The overall flotation and leach extraction for the production, master and variability composites ranged between 79.2 and 94.5% for gold over a range of calculated head grades from 0.47 to 6.02 gram/tonne (“g/t”). The overall flotation and leach extraction for the production, master and variability composites ranged between 72.9 and 95.8% for silver over a range of calculated head grades from 0.92 to 29.86 g/t.

Analysis of the pregnant leach solution produced from rougher concentrates shows no species of concern that would negatively impact precipitation efficiency.

Cyanide detoxification conditions for the master composites 3 and 4 at 45% solids were optimized to be a 5:1 ratio of SO2 to CNWAD, 25 milligrams per liter (“mg/L”) copper and 30 minutes retention for rougher concentrates. For rougher tailings at 53% solids, conditions included a 5:1 ratio of SO2 to CNWAD and 60 minutes retention. Under these conditions, the process met the targets of <5ppm CNWAD for cyanide detoxification.

Dynamic settling tests and slurry viscosity tests on the finely ground concentrates showed good settling characteristics and viscosities within an acceptable range for mixing and pumping applications.

Material flow property testing was carried out for crushed ore and identified there are minimal design risks for the major bulk solids handling systems for the Project.

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A flowsheet maintaining separate high sulphide and low sulphide leach tailings streams reduces the acid generation potential in the co-disposal facility.

1.10 Mineral Resource Estimate

The current mineral resource model prepared by the QP utilizes results from 499 core boreholes. The resource estimation work was completed by Dr. Gilles Arseneau, P.Geo., an independent Qualified Person as this term is defined in NI 43-101. The effective date of the resource statement is September 30, 2025.

The mineral resource estimate was based on a gold price of US$2,450/oz and a silver price of US$27.50/oz, both considered reasonable economic assumptions by the QP. To establish a reasonable prospect of economic extraction in an open pit context, the resources were defined within an optimized pit shell with pit walls set at 35 to 45° based on domain. Assumed metallurgical recoveries of 87.2% for gold and 85.5% for silver were used. Mining costs were estimated at C$2.30/t of total material, processing costs estimated at C$14.50/t, and general and administrative (G&A) costs estimated at C$0.90/t. A cut-off grade (COG) of 0.2 g/t Au was calculated and is considered to be an economically reasonable value corresponding to breakeven mining costs. Approximately 90% of the revenue for the proposed Project is derived from gold, with 10% derived from silver.

Mineral resources were estimated by Ordinary Kriging using Gemcom block modelling software in 10 x 10 x 6 m blocks. Grade estimates were based on capped, 3 m composited assay data, and used a minimum of four and a maximum of 15 composites with no more than three composites permitted from a single drill hole. Grade interpolations were carried out in three passes with each successive pass using a larger search radius than the preceding pass and only estimating the blocks that had not been interpolated by the previous pass. Capping levels were set at 25 g/t for Au and 200 g/t for Ag. Any blocks estimated during Pass 1 or Pass 2 with at least two drill holes and six composites were classified as Indicated mineral resources. All other interpolated blocks were classified as Inferred mineral resource. Mineral resources were then validated using Gemcom GEMS (6.7) software.

The resource model includes mineralized material in the camp, East Extension and Portage zones spanning 1,860 m in the southeast direction along the axis of the Portage zone and 900 m in the northeast direction perpendicular to the long axis of the Portage zone. Resource modelling includes mineralized material generally ranging from 340 m to 440 m below surface.

Mineral resources that are not mineral reserves do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues. There has been insufficient exploration to define these Inferred mineral resources as an Indicated or Measured mineral resource, but the QP is of the opinion that with additional drilling that the majority of the Inferred mineral resources could be upgraded to Indicated mineral resources.

The mineral resources in this Report were estimated in accordance with current Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) standards definitions, and guidelines, and reported using the CIM Definition Standards for Mineral Resources and Mineral Reserves 2014 (CIM Definition Standards, 2014). The updated mineral resource estimate for the Springpole Gold Project is summarized in Table 1‑1.

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Table 1‑1: Mineral Resource Statement^1^ Inclusive of Mineral Reserves (effective September 30, 2025)

Category	Quantity	Grade		Metal

| | | Au | Ag | Au | Ag |

| | (Mt) | (g/t) | (g/t) | (Moz) | (Moz) |

| Indicated | 191 | 0.78 | 4.6 | 4.8 | 28 |

| Inferred | 64 | 0.38 | 3.1 | 0.8 | 6.5 |

Note:

1.	Mineral resources are reported in relation to a conceptual pit shell. Mineral resources are not mineral reserves and do not have demonstrated economic viability. All figures are rounded to reflect the relative accuracy of the estimate. All composites have been capped where appropriate

| 2. | Open pit mineral resources are reported at a COG of 0.20 g/t Au. COGs are based on a gold price of US$2,450/oz and a gold processing recovery of 87.2% and a silver price of US$27.50/oz and a silver processing recovery of 85.5% |

| 3. | Preliminary mining cost assumptions of C$2.60/tonne mined of waste, C$2.30/tonne mined of ore, and C$2.00/tonne mined of overburden, with an incremental mining cost of C$0.02/tonne/6m mined |

| 4. | Preliminary processing cost assumptions of C$14.50/tonne processed, general & administration assumption of C$0.90/tonne processed, stockpile cost assumption of C$0.75/tonne processed, and incremental ore mining cost of C$0.56/tonne processed |

| 5. | Overall pit shell slope angles ranged from 20 - 45° |

1.11 Mineral Reserve Estimate

The mineral reserve estimate for the Springpole Gold Project is based on the conversion of the Indicated mineral resources within the current technical report mine plan. Indicated mineral resources in the mine plan were converted directly to Probable mineral reserves. There are currently no Measured mineral resource estimates for the Project and therefore there are no Proven mineral reserves. The estimated mineral reserves for the Project are shown in Table 1‑2.

Table 1‑2: Springpole Mineral Reserve Estimate – November 13, 2025

Category	Cut-off Grade<br> <br>(g/t Au)	Tonnes<br> <br>(Mt)	Grade Au<br> <br>(g/t)	Grade Ag<br> <br>(g/t)	Contained Ounces Au<br> <br>(M oz)	Contained Ounces Ag<br> <br>(M oz)

| Proven | 0.27 | - | - | - | - | - |

| Probable | 0.27 | 102.0 | 0.94 | 4.90 | 3.1 | 16.1 |

| Total | 0.27 | 102.0 | 0.94 | 4.90 | 3.1 | 16.1 |

Note:

1.	This mineral reserve estimate is as of November 13, 2025, and is based on the new mineral resource estimate dated September 30,2025

| 2. | The mineral reserve estimation was completed under the supervision of Gordon Zurowski, P.Eng. of AGP Mining Consultants Inc., who is a Qualified Person as defined under NI 43-101 |

| 3. | Mineral reserves are stated within the ultimate design pit based on: |

o	US$2100/ounce gold price, US$24/ounce silver price

| o | Pit Limit corresponds to a pit shell with a revenue factor of 0.60, corresponding to a US$1,260 /ounce gold price and US$14.40/oz silver |

| o | A cut-off grade of 0.27 g/t Au for all pit phases |

| o | Preliminary mining cost assumptions of CDN$2.60/tonne mined of waste, CDN$2.30/tonne mined of ore, and CDN$2.00/tonne mined of overburden, with an incremental mining cost of CDN$0.02/tonne/6 m mined |

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o	Preliminary processing cost assumptions of CDN$14.50/tonne processed, general and administration assumption of CDN$0.90/tonne processed, stockpile cost assumption of CDN$0.75/tonne processed, and incremental ore mining cost of CDN$0.56/tonne processed

| o | Preliminary process recovery assumptions of 87.2% for gold and 85.5% for silver |

| o | An exchange rate of CDN$1.35 equal to US$1.00 |

| o | The preliminary economic, cost and recovery assumptions used at the time of mine planning and reserve estimation may not necessarily conform to those stated in the economic model |

4.	Pit slope inter-ramp slope angle assumptions ranged from 22 - 54°

The mineral reserves are based solely on the Springpole open pit. The reference point at which mineral reserves are defined is the point where the ore is delivered to the process plant complex, which includes the ore stockpiles.

At the time of preparing the mineral reserve estimate, the QP has not identified any known legal, political, environmental, or other risks that would materially affect the potential development of the mineral reserves. The risk of not being able to secure the necessary permits from the government for development and operation of the Project exists but the QP is not aware of any issues that would prevent those permits from being withheld per the normal permitting process.

1.12 Mining Methods

The PFS mine plan is based on open pit mining. A single pit will provide the open pit feed material necessary to maintain the process plant feed rate at 30,000 t/d while operational. The Springpole pit is proposed as a three-phase design using 12 m benches which provides 102.0 million tonne (“Mt”) of mill feed grading 0.94 g/t Au, and 4.90 g/t Ag. Waste from this pit will total 309.5 Mt for a strip ratio of 3.0 (waste:mill feed).

The mill feed cut-off used is 0.25 g/t AuEq, accounting for the contribution of both gold and silver. This is equivalent to a gold only cut-off of 0.27 g/t Au. The factors contributing to the calculation of Au equivalent grade AuEQ are discussed in Section 16.5.2.1 Cut-off Grade. During the mine operation, material would be stockpiled to optimize the plant feed grade and defer lower-grade material until later in the mine schedule. The calculation for the Au Equivalent grade is provided in Section 16.

The phases are scheduled to provide 30,000 t/d of feed to the mill over a 9.4-year mining life after one year of pre-production stripping. The last 1.4 years of mine life are stockpile reclaim. The pits are sequenced to minimize initial stripping and provide higher feed grades in the early years of the mine life which the stockpiling strategy accomplishes.

The main fleet will consist of up to four 251 mm rotary drills and two 140 mm drills, two 37 m^3^ electric hydraulic shovels and three 23 m^3^ front end loaders. The truck fleet will total 25 – 240-tonne trucks at the peak of mining. This is due to the long hauls from the pit to the co-disposal facility (“CDF”). The usual assortment of dozers, graders, small backhoes, and other support equipment is considered in the equipment cost estimate. A smaller front-end loader (13 m^3^) will be stationed at the primary crusher.

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Year -1 is the start of major mining activity using the larger equipment when the controlled dewatering of the open pit basin has advanced sufficiently for mining and the site infrastructure (power lines, roads, etc.) are in place. The early phases will provide the highest grade to the mill early in the schedule. The open pit will be in operation until Year 8 followed by 1.4 years of stockpile reclaim to feed the plant.

Waste material from the pit will be stored in the CDF. Non-acid generating (NAG)/non-metal leaching material will be used for the outer embankments while potentially acid generating (PAG) material will be co-disposed with process plant tailings. The majority of the NAG waste rock material from the open pit will be contained within the CDF (134.6 Mt), but a small portion of NAG/non-metal leaching material will be backfilled into Phase 2 near the end of the mine life. This reduces the overall haul length and helps in reclamation of the pit. An estimated total of 15.4 Mt will be backfilled into the pit.

In addition to the open pit, two quarries will be established during the pre-production period and at closure. These quarries will be used to provide rock material for various mine infrastructure including haul roads, dikes, CDF embankments, and to meet site fill requirements for other infrastructure. Approximately 7.4 Mt of material is planned to be excavated from the CDF quarry in the preproduction period, while 26.2 Mt of material is planned from the fish habitat development area at the end of the mine life.

When the open pit is completed, the larger mining fleet will move to complete the fish habitat area. Material will be used as cover for the CDF cells and will be dumped into the open pit. This serves to cover the slopes in the pit for reclamation purposes and upon closure the quarry will become additional lake area which will be contoured to provide suitable fish habitats.

1.13 Recovery Methods

The process plant will process 30 kilotonnes/day (“kt/d”) over the life of mine (“LOM”).

The selected flowsheet includes primary crushing with the crushed product stored in a covered crushed ore stockpile. Crushed ore is reclaimed to a grinding circuit consisting of a SAG mill and a ball mill operating in closed circuit with a classification cyclone cluster. Cyclone overflow feeds a single stage rougher flotation circuit, with a concentrate regrind stage following flotation.

Reground concentrate reports to a concentrate thickener, whilst tailings from the flotation circuit reports to a tailings thickener. The concentrate circuit utilizes cyanide leach and counter current decantation (CCD) circuit followed by a Merrill Crowe precipitation circuit and smelting to produce doré to market. The tailings circuit utilizes a cyanide leach circuit and carbon in pulp (CIP) circuit followed by elution, Merrill Crowe, and smelting to produce doré. The leach residue from the concentrate leach circuit feeds the concentrate cyanide detoxification circuit, whilst the CIP residue from the tailings CIP circuit feeds the tailings cyanide detoxification circuit. Material from the concentrate cyanide detoxification circuit is sent to the PAG tailings deposition cell and material from the tailings detoxification is sent to the CDF.

Key design operating availabilities are 75% for primary crushing and 92% for the grinding, flotation, leach, CCD and CIP circuits.

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1.14 Project Infrastructure

1.14.1 Infrastructure Summary

Key project infrastructure includes the open pit mine area with associated haul roads and ramps, the dikes required to hydraulically isolate the pit after bay dewatering, site access roads, and the administrative and dry facilities. The Project will also include construction and permanent camp accommodations, the process plant electrical room, the crushing area electrical room, the central control room, and the reagent storage building. Additional infrastructure consists of the gold room, the assay laboratory and sample preparation facilities, the plant workshop and warehouse, the truck shop and warehouse, the tire-change facility, the truck wash building, and the fuel storage and dispensing facilities. Water and power infrastructure includes the freshwater intake system, the 230 kilovolt (“kV”) overland and 25 kV underground power distribution lines, the freshwater intake pumping and treatment systems, the waste storage and CDF, contact water collection ponds, the wastewater treatment plant, and the explosives magazine.

The main access road will be a private extension of the existing Wenesaga Road, which currently supports forestry operations and has been constructed to within approximately 15 km of the Project site.

Project electrical demand, estimated at approximately 60 megawatt (“MW”), will be supplied through a new 230 kV overhead transmission line that will connect to the provincial grid roughly 90 km southeast of the site. Power will be stepped down through a 230 kV/25 kV transformer before being distributed to six electrical rooms. Variable-frequency drives will be used where required, and all medium-voltage motors and drives will operate at 4.16 kV. To support early works and construction activities, a temporary tie-in to a 115 kV line located approximately 30 km to the south will be utilized until the permanent transmission infrastructure is commissioned.

The layout of the site is presented below in Figure 1‑1 and layout of the process plant is presented below in Figure 1‑2.

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Figure 1‑1: Overall Site Layout Plan

Source: Ausenco, 2025

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Figure 1‑2: Process Infrastructure Layout Plan

Source: Ausenco, 2025

1.14.2 Dikes

Two dikes will be constructed to isolate the area of the proposed open pit from Springpole Lake and facilitate mining following dewatering. The dikes will have total length of about 1,000 m, average height of about 7 m and a maximum height of about 15 m. The dikes will be constructed using NAG/non-metal leaching rockfill from a quarry located within CDF and/or from pre-production phase open pit stripping (sourced from an area outside the existing lake). Seepage cut-off wall and foundation grout curtain will be incorporated within the dikes to establish a hydraulic barrier. At closure, once the water quality meets all regulatory requirements, the dewatered open pit area will be allowed to flood back by lowering of the dikes in a controlled manner and reconnection of the reclaimed basin to Springpole Lake.

1.14.3 Co-Disposal Facility

The CDF will be constructed to manage and store 101 Mt of tailings and 146 Mt of potentially acid generating (PAG) mine rock. The CDF will be located immediately west of the open pit on an area of generally thin overburden and bedrock outcrops. The perimeter containment dams will generally be founded on bedrock which is andesite. The andesite bedrock provides good foundation characteristics (high integrity and low permeability) supporting a structurally safe facility with effective seepage capture and management.

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The CDF will consist of two cells: south cell and north cell. The south cell perimeter dams will be constructed with quarried rock and NAG/non-metal leaching (“NML”) mine rock as a downstream raise and will incorporate a low-permeability liner to maintain saturated conditions during operations. The north cell perimeter dams will be constructed as a center line raise using quarried rock and NAG/NML mine rock. The south and north cells will be separated by an internal dam. The north and south cell perimeter dams will reach an average ultimate height of about 75 m. PAG mine rock (146 Mt) will be co-disposed with NAG tailings (80.8 Mt) in the north cell. The NAG tailings will encapsulate the mine rock providing a low permeability zone to limit oxygen ingress. PAG tailings (20.2 Mt) will be subaqueously deposited in the south cell. CDF contact water will be collected and managed in the south cell and pumped to the central water management pond located in proximity to the plant. The water stored in the contact water management pond (CWMP) will be used to supplement mineral processing and/or will be treated at the on-site effluent treatment plant prior to release.

At closure, the north cell will be graded and provided with erosion protection cover to continue directing runoff to south cell. NAG tailings and/or suitable soil and granular cover will be provided over the south cell PAG tailings to manage water and safely manage extreme runoff events.

1.15 Environmental and Social Setting

First Mining has been actively collecting environmental baseline data necessary to support the Project’s environmental assessment and permitting since 2011.

The Red Lake area has been a historic mining district in northwestern Ontario since the gold rush of the 1920s, and there are currently several active mining projects and other decommissioned mines situated within the area. The Project site is located to the east of this district in a remote area and there are no nearby active industrial/commercial developments. Temperatures at the Project site annually average 1.3 degrees Celsius (°C) and range from a low of -18.3°C (January) to a high of 18.1°C (July). Owing to the remote location of the project, anthropogenic sources of industrial air emissions are limited in the area. The property is overlain by glaciated terrain common throughout the Canadian Shield. Land areas are generally of low relief with less than 30 m of local elevation change. Tree cover consists of mature trembling aspen, black spruce, white birch, balsam fir, and white spruce and jack pine. Black spruce and muskeg swamps occupy low-lying areas. Glacial till is generally thin across the site (less than 4 m in thickness). Outcrops are limited and small and are generally covered by a layer of moss or muskeg. Land areas are permeated by a series of interconnected shallow ponds and lakes. There is generally low to moderate relief in the vicinity of the Project, with generally dry uplands and poorly drained lowland valleys with thick accumulations of organic soils. Groundwater flow generally follows surface water flow directions. From the monitoring data it was observed that the greatest surface water flows were measured in the spring following freshet, and the lowest flows in the late winter.

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Springpole Lake as well as the nearby Birch Lake support diverse fish communities including sport species common to both cold-water lakes (lake trout and whitefish) and cool-water lakes (walleye, northern pike, and yellow perch), and several other non-game and forage fish. Among the small lakes surveyed, six host fish communities that include sport species including yellow perch and northern pike, four host only forage species and two are considered devoid of fish.

Forest composition at the Project is typical of the Lac Seul Upland. Dominant tree species include trembling aspen, black spruce, white birch, balsam fir, and white spruce and jack pine. Understory ground cover species composition and abundance is typical of mesic mixed wood boreal sites and lacks microhabitats likely to harbour rare vascular plant species. The Project is within the Trout Lake Forest Management Area and forestry activities are ongoing in the region, in accordance with the Crown Forest Sustainability Act. Several large forest fires have occurred in the Trout Lake Forest Management Area since 1961 and these areas are currently in various stages of regeneration.

The Project area is within the Churchill Caribou Range, and numerous wildlife species are present within the region, including caribou, moose, marten, lynx, snowshoe hare, and fisher, as well as many species of birds.

The Project is located approximately 40 km from Cat Lake First Nation, 45 km from Slate Falls Nation, and 120 km from Lac Seul First Nation. First Mining is engaging with eight Indigenous communities (including seven First Nations and the Northwestern Ontario Métis Community) and will continue this engagement during the permitting process.

1.15.1 Environmental and Social Impacts

A variety of environmental protection, mitigation and management measures have been incorporated into planning, design, construction, operation and ultimate closure of the Project. The engineering design of the Project incorporates climate change considerations, and Project components and infrastructure are being designed to manage variable weather events.

The environmental assessment (EA) process is a planning tool to ensure the Project is considered in a careful and precautionary manner that avoids or mitigates potential environmental effects and considers the benefits and opportunities from the Project. First Mining published the final EIS/EA in November 2024 which assessed the potential effects of the Project on all relevant components of the environment including air quality, noise, water, fish, vegetation, wildlife, land and resource use, the economy, local infrastructure and services, archaeology, cultural heritage, traditional land and resource use, and human and ecological health. The final EIS/EA concluded that, considering the Project designs and mitigation measures, no significant adverse environmental effects were predicted for the Project or for the Project in combination with other projects (WSP, 2024).

The follow-up and monitoring program will be implemented to verify predicted effects, evaluate the effectiveness of mitigation, and measure compliance with future permit and authorization conditions. Modifications to follow-up and monitoring programs may occur as a result of applying an adaptive management approach over all phases of the Project (construction, operations, and decommissioning and closure).

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1.15.2 Waste and Water Management

Approximately 101 Mt of tailings, 66 Mt of NAG mine rock and 146 Mt of PAG mine rock will be stored within the CDF. The construction of perimeter embankment dams using NAG mine rock will improve stability and provide freeboard for runoff collection as well as storage for storm events.

The surface of the tailings and PAG mine rock within the CDF will be graded to encourage flow to defined sumps/pumping points. The collected surface water will be directed to a central water storage pond located east of the CDF. The water stored in the central water storage pond will be used to supplement mineral processing and/or treated and discharged to the environment in accordance with applicable regulatory requirements.

1.15.3 Environmental Assessment Process

On February 23, 2018, First Mining submitted a project description to the Impact Assessment Agency of Canada (IAAC). IAAC determined an environmental assessment is required for the Springpole Gold Project under the Canadian Environmental Assessment Act (2012). First Mining has also entered into a voluntary agreement with the Ministry of the Environment, Conservation and Parks (MECP) to undertake a comprehensive Environmental Assessment (EA) under the provincial Environmental Assessment Act.

A final environmental impact statement/environmental assessment document (EIS/EA) was submitted to IAAC and MECP in November 2024. The document was prepared in accordance with the provincially approved Amended Terms of Reference (“ToR”) and the requirements of the Ontario Environmental Assessment Act, as well as the Canadian Environmental Assessment Act, 2021. The information presented in the EIS/EA is currently under review by both IAAC and MECP.

1.15.4 Engagement and Consultation

The federal government identified Cat Lake First Nation, Slate Falls Nation, Lac Seul First Nation, Wabauskang First Nation, Mishkeegogoamang Ojibway Nation, and Métis Nation of Ontario in 2018 (updated in 2020), while in 2018 the provincial government identified Cat Lake First Nation, Slate Falls Nation, Lac Seul First Nation, Wabauskang First Nation, Mishkeegogoamang Ojibway Nation, Ojibway Nation of Saugeen, Pikangikum First Nation, and Métis Nation of Ontario, as potentially impacted by the Project or having an interest in the Project. The Métis Nation of Ontario, Region 1, as represented by the Northwestern Ontario Métis Community (NWOMC), office is located in the City of Dryden, 185 km from the Project site.

First Mining has continued to share information and engage extensively with the participating communities throughout the EA process.

1.15.5 Mine Closure and Rehabilitation

The conceptual closure strategy has been presented in the final EIS/EA to support the assessment of potential effects. The general rehabilitation and closure approach for the Project will meet the objectives of the Ontario Mining Act and Regulation 35/24. The overall objective for closure is to return the project site to a productive condition after mining is complete that is capable of supporting plant, wildlife and fish communities, and other applicable land uses. A Certified Closure Plan prepared in accordance with O.Reg. 35/24 will be filed under the Mining Act prior to construction.

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1.16 Capital and Operating Cost

1.16.1 Capital Cost Estimate

The following basic information pertains to the estimate:

·	Cost estimate base date is Q4, 2025.
·	Expressed in Canadian dollars (C$) unless otherwise noted.
·	Currency exchange rate 0.74 US$ - 1 C$.

The cost estimate is based on an engineering, procurement, and construction management (EPCM) implementation approach. The capital cost estimate has an accuracy of -20%/+30% (AACE Class 4). The estimate includes the cost to complete the design, procurement, construction, and commissioning of all the identified facilities.

The estimate was based on the 2021 PFS and updated with revised assumptions as indicated in process flow diagrams, general arrangements, mechanical equipment list, electrical equipment list, material take-offs (MTOs), electrical layouts, scope definition and a work breakdown structure. The estimate included all associated infrastructure as defined by the scope of work.

The study was prepared in Canadian dollars (CAD) and is reported in United States dollars (USD) to enable consistent comparison with other projects denominated in USD. Mining Operating costs, provided in Section 21.3.2, have been expressed in CAD.

The initial and LOM capital cost estimate is summarized in Table 1‑3. The operating cost estimate is shown in Table 1‑4.

Table 1‑3 : Estimate Summary Level 1 Major Facility

WBS Level 1	Description	Initial CAPEX<br> <br>(US$M)	Sustaining CAPEX<br> <br>(US$M)	Closure Costs<br> <br>(US$M)	Total CAPEX<br> <br>(US$M)

| 1000 | Mining | 302.5 | 304.1 | 40.5 | 647.1 |

| 2000 | Site Development | 39.6 | - | - | 39.6 |

| 3000 | Process Plant | 348.6 | - | - | 348.6 |

| 4000 | On-Site Infrastructure | 75.5 | - | - | 75.5 |

| 5000 | Off-Site Infrastructure | 47.0 | - | - | 47.0 |

| Total Direct Costs | | 813.2 | 304.1 | 40.5 | 1157.9 |

| 6000 | Indirects | 68.6 | - | - | 68.6 |

| 7000 | EPCM Services | 70.2 | 18.5 | - | 88.8 |

| 8000 | Owners Costs | 24.4 | - | - | 24.4 |

| Total Indirect Costs | | 163.3 | 18.5 | - | 181.8 |

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WBS Level 1	Description	Initial CAPEX<br> <br>(US$M)	Sustaining CAPEX<br> <br>(US$M)	Closure Costs<br> <br>(US$M)	Total CAPEX<br> <br>(US$M)

| 9000 | Provisions (Contingency) | 127.5 | - | - | 127.5 |

| Project Total | | 1,104.1 | 322.6 | 40.5 | 1,467.2 |

1.16.2 Operating Cost Estimate

The costs considered on-site operating costs are those related to mining, processing, maintenance, power, water treatment and general and administrative activities. The operating cost estimate was developed in Q4 2025 using data from vendor quotations, projects, studies and previous operations from internal databases. The operating cost estimate accuracy is approximately -25 to +15%.

Mining operating costs are US$12.44/t of ore processed. The average process operating cost is US$10.72/t processed, and the annual G&A costs are US$27M.

A summary of the average operating costs is shown below in Table 1‑4.

Table 1‑4 : Operating Cost Summary

Cost Area	Total (US$M/a)	US$/t

| Mining | 106.8 | 12.44 |

| Process | 113.1 | 10.72 |

| G&A | 27.0 | 2.56 |

| Total | 246.9 | 23.15 |

1.17 Economic Analysis

1.17.1 Economic Summary

The economic analysis was performed assuming a 5% discount rate for gold projects in Canada. Cash flows have been discounted to the start of construction, assuming that the project execution decision will be made, and major project financing will be carried out at this time.

The pre-tax NPV discounted at 5% (NPV5%) is US$3,228 million; the internal rate of return (IRR) is 53.8%, and payback period is 1.4 years. On a post-tax basis, the NPV5% is US$2,150 million, the IRR is 40.9%, and the payback period is 1.8 years. Cumulative post-tax unlevered free cash flow totals US$3,137 million. Tax calculations are based on the applicable tax law in place as of the date of this report which includes both federal and provincial taxes.

A summary of the project economics is presented in Table 1‑5, and post-tax-free cash flow is shown graphically in Figure 1‑3.

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Figure 1‑3: Undiscounted, Unlevered, Free Cash Flow – Post Tax Basis

Source: Ausenco, 2025

Table 1‑5 : Economic Analysis Summary

Item	Units	LOM	Years 1 to 5

| Gold Price | US$/oz | 3,100 | 3,100 |

| Silver Price | US$/oz | 35.50 | 35.50 |

| Foreign Exchange Rate | CAD:USD | 0.74 | 0.74 |

| Total Tonnes Processed | Mt | 102.0 | 53.6 |

| Total Tonnes Waste Mined | Mt | 309.5 | 217.0 |

| Mill Feed Grade – Au | g/t | 0.94 | 1.09 |

| Mill Feed Grade – Ag | g/t | 4.9 | 5.7 |

| Mine Life | Years | 9.4 | 5.0 |

| Mill Throughput | t/d | 30,000 | 30,000 |

| Average Strip Ratio | waste:ore | 3:1 | 3.2:1 |

| Average Recovery Rate – Au | % | 86.0 | 86.7 |

| Average Recovery Rate – Ag | % | 86.2 | 87.1 |

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Item	Units	LOM	Years 1 to 5

| Total Production – Au | koz | 2,648 | 1,626 |

| Total Production – Ag | koz | 13,842 | 8,518 |

| Average Annual Production – Au | koz/a | 281 | 325 |

| Average Annual Production – Ag | koz/a | 1,468 | 1,704 |

| Total Payable Metal – Au | koz | 2,646 | 1,624 |

| Total Payable Metal – Ag | koz | 13,150 | 8,092 |

| Total Revenue – LOM | U$M | 8,668 | 5,323 |

| Average Annual Revenue | US$M/a | 920 | 1,065 |

| Total EBITDA – LOM | US$M | 6,079 | 3,830 |

| Average Annual EBITDA | US$M/a | 645 | 766 |

| Total Operating Costs – LOM | US$M | 2,361 | 1,352 |

| Average Annual Operating Cost | US$M/a | 251 | 270 |

| Mining Cost | US$/t mined | 2.57 | 2.41 |

| Mining Cost | US$/t milled | 9.87 | 11.93 |

| Processing Cost | US$/t milled | 10.72 | 10.74 |

| G&A and Site Services Cost | US$/t milled | 2.56 | 2.58 |

| Total Operating Cost | US$/t milled | 23.15 | 25.26 |

| Total Cash Cost^1^ | US$/oz Au | 802.4 | 742.4 |

| All-in Sustaining Cost^2^ | US$/oz Au | 937.9 | 876.2 |

| Initial Capital Cost | US$M | 1,104 | - |

| Sustaining Capital Cost | US$M | 322 | 217 |

| Closure Cost | US$M | 40 | - |

| Salvage Value | US$M | 4 | - |

| NPV5% | US$M | 3,228 | 2,150 |

| IRR | % | 53.8 | 40.9 |

| Payback Period | Years | 1.4 | 1.8 |

| Undiscounted Cash Flow | US$M | 4,588 | 3,137 |

| NPV5%: Initial Capital Cost | NPV:Capex | 2.9 | 1.9 |

Note:

1.	Cash costs consist of mining costs, processing costs, mine-level G&A and refining charges, royalties, and streaming costs

| 2. | AISC includes cash costs plus sustaining capital, closure cost and salvage value |

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1.17.2 Sensitivity Analysis

A sensitivity analysis was conducted on the base case pre-tax and post-tax NPV and IRR of the Project, using the following variables: metal prices, discount rate, head grade, total operating cost, and total capital cost.

The sensitivity analysis revealed that the Project’s NPV is most sensitive to changes in gold price and operating cost, whereas IRR is sensitive to metal price and initial capital cost.

1.18 Conclusions and Recommendations ****

The Mineral Resource estimate establishes a reasonable basis for the Project and demonstrates sufficient technical merit and economic potential to justify advancement to the next stage of project development.

The objective of the recommended work program is to advance the Project by improving confidence in the Mineral Resource estimate and supporting future technical studies. The recommended scope of work and associated estimated costs are summarized below in Table 1‑6 with further details provided in Section 25 and 26.

Table 1‑6 : Summarized Recommended Work Program

Proposed Work	Approximate Cost (US$M)

| Inferred Material Upgrade (including contingency) | $3.3 |

| Exploration (including contingency) | $4.3 |

| Metallurgical testing | $0.6 |

| Mine Geotechnical Program | $0.6 |

| Mining Methods | $0.4 |

| Infrastructure | $11.3 |

| Total | $20.5 |

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2 INTRODUCTION

2.1 Introduction

This report was prepared by Ausenco Engineering Canada ULC. (“Ausenco”), AGP Mining Consultants Inc. (“AGP”), SRK Consulting (Canada) Inc. (“SRK”) and WSP Global Inc. (“WSP”) for First Mining Gold Corp. (“First Mining” or the “Company”) to summarize the results of the Pre-Feasibility Study (PFS) of the Springpole Gold Project (the “Project” or “Springpole”).

2.2 Terms of Reference

The technical report was prepared in accordance with the Canadian disclosure requirements of National Instrument 43-101 (NI 43-101) and Form 43-101 F1, and is prepared using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for mineral resources and mineral reserves (CIM Definition Standards, 2014) and the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines (CIM Best Practice Guidelines, 2019).

The report supports disclosures by First Mining in a news release dated November 18, 2025 entitled “First Mining Announces Updated Pre-Feasibility Study for the Springpole Gold Project, Ontario.”

All measurement units used in this report are SI units unless otherwise noted. Currency is expressed in Canadian dollars (C$).

2.3 Qualified Persons

The Qualified Persons for the report are listed in Table 2‑1. By virtue of their education, experience and professional association membership, they are considered Qualified Person as defined by NI 43-101.

Table 2‑1 : Report Contributors

Qualified Person	Professional<br> <br>Designation	Position	Employer	Independent of Client

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2.4 Site Visits and Scope of Personal Inspection

2.4.1 Site Inspection for Tommaso Roberto Raponi

Mr. Raponi visited the Project site on July, 29 2025 in preparation for this report. The following areas were inspected:

·	The proposed location of the open pit, tailings and waste rock deposition areas
·	The proposed location of the process plant

In addition, Mr. Raponi discussed roads and infrastructure components (power, roads, access points, etc.) with site personnel and mineralized core from the different ore types that would be processed was inspected at the off-site core shack.

2.4.2 Site Inspection for David Bleiker

Mr. Bleiker visited the Project site on September 4, 2025 in preparation for this report. The following areas were inspected:

·	The CDF area (by air) and the dike abutments.

2.4.3 Site Inspection for Gilles Arseneau

Mr. Arseneau visited the project site from February 10 and 11, 2012, August 8 and 9, 2012 and June 20 to 22, 2022.

During the site visits, core logging procedures were reviewed. Several sections of core from the Portage, Camp, and East Extension zones were examined. Sampling procedures and handling were observed. The deposit geology, alteration, and core recovery data were observed for the Portage zone.

Dr Arseneau re-logged mineralized sections of drill core from the Springpole deposit and checked geological units against the recorded written logs. Down-hole survey data entered in the digital database was checked against data entered on paper logs at the site and no errors were noted. Drill site locations could not be verified as most drill sites are situated under Springpole Lake, but two drill platforms drilling on Springpole Lake were observed during one visit.

2.5 Effective Dates

Effective date of Mineral Resource Estimate: September 30, 2025

Effective date of Mineral Reserve Estimate: November 13, 2025

Effective date of Financial Model: December 1, 2025

The effective date of this report is December 19, 2025

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There were no material changes to the scientific and technical information on the Project between the effective date and the signature date of the Report.

2.6 Information Sources and References

2.6.1 Previous Technical Reports

The main sources of information in preparing this report are based on information located within internal reports and memoranda obtained from First Mining. Information, conclusions and recommendations contained herein are based on a field examination, including a study of relevant and available technical data, including, and not limited to the numerous reports listed in the references section of this technical report. This technical report is prepared with the most recent information available at the time of study.

The Project has been the subject of several technical reports. The previous NI 43-101 technical reports are listed in the References section of this report and are summarized in Table 2‑2.

Table 2‑2 : Previous Technical Reports

Reference	Date	Company	Name

| Armstrong T, et al., 2006 | 2006 | P & E Mining Consultants | Technical Report and Resource Estimate on the Springpole Lake Gold Property, Red Lake Mining Division, NW Ontario for Gold Canyon Resources |

2.7 Currency, Units, Abbreviations and Definitions

All grid references are based on the NAD83 Datum (NAD83) UTM coordinate system unless otherwise stated. All units of measurement in this report are metric and all currencies are expressed in Canadian dollars (symbol: C$ or currency: CAD) unless otherwise stated. Contained gold metal is expressed as troy ounces (oz), where 1 oz = 31.1035 g. All material tonnes are expressed as dry tonnes (t) unless stated otherwise. A list of abbreviations and acronyms is provided in Table 2‑3, and units of measurement are listed in Table 2‑4.

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Table 2‑3: Abbreviations and Acronyms

Abbreviation	Description

| AA | atomic absorption spectroscopy |

| AACI | Accredited Appraiser Canadian Institute |

| AGP | AGP Mining Consultants Inc. |

| Ag | silver |

| Au | gold |

| Ausenco | Ausenco Engineering Canada ULC |

| Az | azimuth |

| BIF | banded iron formation |

| BWi | bond ball mill work index |

| CAD:USD | Canadian-American exchange rate |

| CCD | counter current decantation |

| CDF | co-disposal facility |

| CIM | Canadian Institute of Mining, Metallurgy and Petroleum |

| CIP | carbon in pulp |

| CIM Best Practice Guidelines | CIM Estimation of Mineral Resources & Mineral Reserves Best Practice Guidelines |

| CIM Definition Standards | CIM Definition Standards for Mineral Resources and Mineral Reserves |

| CIP | carbon in pulp |

| CoG | cut-off grade |

| CRM | certified reference material |

| CWi | Bond crusher work index |

| CWMP | Contact Water Management Pond |

| DCIP | direct current resistivity and induced polarization |

| DDH | diamond drill hole |

| DFO | Fisheries and Oceans Canada |

| EA | environmental assessment |

| E-GRG | extended gravity recoverable gold |

| EIS | environmental impact statement |

| EM | electromagnetic |

| FA | fire assay |

| FET | federal excise tax |

| FS | feasibility study |

| G&A | general and administration |

| GPR | gross production royalty |

| GQCV | greenstone-hosted quartz-carbonate vein deposits |

| GRAV | gravimetric finish method |

| ICP | inductively coupled plasma |

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Abbreviation	Description

| IAAC | Impact Assessment Agency of Canada |

| ICP-OES | inductively coupled plasma - optical emission spectrometry |

| ID2 | inverse distance squared |

| ID3 | inverse distance cubed |

| IDF | intensity duration frequency |

| IOCG | iron oxide copper gold |

| IP | induced polarization |

| IRGS | intrusion-related gold system |

| IRR | internal rate of return |

| ISO | International Organization for Standardization |

| LIDAR | light detection and ranging |

| LUP | land use permit |

| MCF | mechanized cut and fill |

| MECP | Ministry of the Environment, Conservation and Parks |

| MLAS | Mining Lands Administration System |

| MRE | mineral resource estimate |

| MTOs | material take-offs |

| NAD 83 | North American Datum of 1983 |

| NAG | Non-Acid Generating |

| NI 43‑101 | National Instrument 43-101 (Regulation 43-101 in Quebec) |

| NN | nearest neighbour |

| NSR | net smelter return |

| NTS | national topographic system |

| OK | ordinary kriging |

| PAG | potentially acid generating |

| PEA | preliminary economic assessment |

| PFS | Pre-Feasibility Study |

| PGE | platinum group elements |

| QA/QC | quality assurance/quality control |

| QP | qualified person (as defined in National Instrument 43-101) |

| ROM | run of mine |

| RQD | rock quality designation |

| SAG | semi-autogenous grinding |

| SCC | Standards Council of Canada |

| SD | standard deviation |

| Sd-BWI | micro hardness or bond ball mill work index on SAG ground material |

| SEDEX | sedimentary exhalative deposits |

| SG | specific gravity |

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Abbreviation	Description

| Springpole Gold Project | The Project or Springpole |

| SRK Consulting (Canada) Inc. | SRK |

| TMF | tailings management facility |

| UG | underground |

| UTM | Universal Transverse Mercator coordinate system |

| UV | ultraviolet |

| VLF-EM | very low frequency electromagnetic |

| VMS | volcanogenic massive sulphide |

| WSP Global Inc. | WSP |

Table 2‑4 : Units of Measurement

Abbreviation	Description

| % | percent |

| % solids | percent solids by weight |

| CAD | Canadian dollar (currency) |

| C$ | Canadian dollar (as symbol) |

| $/t | dollars per metric ton |

| ° | angular degree |

| °C | degree Celsius |

| μm | micron (micrometre) |

| cm | centimetre |

| cm^3^ | cubic centimetre |

| ft | foot (12 inches) |

| g | gram |

| g/t | gram/tonne |

| g/cm^3^ | gram per cubic centimetre |

| g/L | gram per litre |

| g/t | gram per metric ton (tonne) |

| h | hour (60 minutes) |

| h/day | hours per day |

| ha | hectare |

| kg | kilogram |

| kg/t | kilogram per tonne |

| km | kilometre |

| km^2^ | square kilometre |

| kt/d | kilotonne/day |

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Abbreviation	Description

| kV | kilovolt |

| kW | kilowatt |

| kWh/t | kilowatt-hour per tonne |

| L | litre |

| lb | pound |

| m, m^2^, m^3^ | metre, square metre, cubic metre |

| mg/L | milligram per litre |

| M | million |

| Ma | million years (annum) |

| masl | metres above mean sea level |

| mm | millimetre |

| Moz | million (troy) ounces |

| Mt | million tonnes |

| MW | megawatt |

| oz | troy ounce |

| oz/t | ounce (troy) per tonne |

| oz/ton | ounce (troy) per short ton (2,000 lbs) |

| ppb | parts per billion |

| ppm | parts per million |

| t | metric tonne (1,000 kg) |

| ton | short ton (2,000 lbs) |

| t/d | tonnes per day |

| USD | US dollars (currency) |

| US$ | US dollar (as symbol) |

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3 RELIANCE ON OTHER EXPERTS

3.1 Introduction

The QPs have relied upon the following other expert reports, which provided information regarding mineral rights, surface rights, property agreements, royalties and taxation as noted below.

3.2 Ownership, Mineral Tenure and Surface Rights

The QPs have not conducted detailed land status evaluations and have relied upon previous qualified reports and public documents and have reviewed an opinion provided by Bennett Jones LLP to First Mining dated August 26, 2020 as to the apparent registered owner with respect to certain real property interests and mining claims comprising the Project as of such date, and which is subject to certain assumptions and qualifications as referred therein.

The above referenced document is:

·	Title Opinion Letter from Bennett Jones LLP addressed to First Mining, August 26, 2020

This information is used in Section 4 of the report and in support of the Mineral Resource estimate in Section 14 and the financial analysis in Section 22.

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4 PROPERTY DESCRIPTION AND LOCATION

4.1 Summary

The Project is located 110 km northeast of the Municipality of Red Lake in northwest Ontario, Canada (Figure 4‑1). The Project is centered on a temporary exploration camp situated on a small land bridge between Springpole Lake and Birch Lake. The latitude and longitude coordinates are:

·	Latitude	N51° 23ʹ 44.3ʺ
·	Longitude	W92° 17ʹ 37.4ʺ

The Universal Transverse Mercator (UTM) map projection based on the World Geodetic System 1984 (WGS84) zone 15N is:

·	Easting	549,183
·	Northing	5,693,578
·	Average Elevation	395 m

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Figure 4‑1: Springpole Gold Project Location Map

Source: First Mining, 2025

4.2 Land Area

The Project is wholly-owned and controlled by First Mining through its subsidiary company Gold Canyon Resources (“Gold Canyon”) and comprises 30 patented mining claims, 280 contiguous mining claims and 13 mining leases totalling an area of 41,952 ha.

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Additional claims adjacent to the Project area within the Birch-Uchi Greenstone Belt have been acquired by Gold Canyon between 2021 and 2025 (the ‘Birch-Uchi Tenure”). This additional tenure comprises a further 562 mining claims totalling 15,895 ha which are 100% owned by Gold Canyon, plus a further 82 mining claims (1,656 ha) which are 70% owned by Gold Canyon and 30% owned by Whitefish Exploration Inc.

The overall Springpole Gold Project land area is represented in Figure 4‑2 and a more detailed tenure map is shown in Appendix A. A full list of the patents, mining leases and mining claims that comprise the Project and adjacent Birch-Uchi tenure is provided in Appendix B.

The proposed Springpole mine development area, as defined in the published EIS/EA for the Project, totals 1,528 ha (not including access or transmission routes) and this area is highlighted on Figure 4‑2.

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Figure 4‑2: Mineral Tenure Plan

Source: First Mining, 2025

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4.3 Mineral Tenure

First Mining acquired 100% of the Springpole Gold Project on November 13, 2015 when it completed the acquisition of Gold Canyon. On acquisition, the Gold Canyon mineral tenure consisted of 30 patented mining claims and 300 unpatented, contiguous mining claims and six Crown mining leases, totalling an area of approximately 32,448 ha. Additional mining claims were subsequently acquired by First Mining in the Satterly Lake area, and the original unpatented ‘legacy’ claims were converted into the new Ontario cell claim system in April 2018 (see Section 4.3.4). A further seven mining leases were acquired by Gold Canyon in 2019 by conversion of existing mining claims covering 1,531 ha to mining leases. As of the date of this report, the Project area comprises 30 patented mining claims, 280 contiguous mining claims and 13 mining leases totalling an area of 41,952 ha. With the exception of the 25 patented claims discussed in Sections 4.3.2 and 4.3.3, all mining claims, leases and patents at the Project are registered 100% to Gold Canyon, a wholly-owned subsidiary of First Mining.

Between 2021 and 2025, an additional 562 mining claims adjacent to the Project within the Birch-Uchi Greenstone Belt were acquired by Gold Canyon, totalling 15,895 ha. Gold Canyon also hold a 70% interest in a further 82 mining claims in the Birch-Uchi Greenstone Belt, following the 29 April 2024 completion of an Option Agreement with Whitefish Exploration Inc. (“Whitefish”) on the Swain property, with Whitefish holding the remaining 30% interest.

The QPs have not conducted detailed land status evaluations and have relied upon previous qualified reports and public documents and have reviewed a title opinion provided by Bennett Jones LLP to First Mining dated August 26, 2020 as to apparent registered owner with respect to certain real property interests and mining claims comprising the Project as of the date of such opinion, and which is subject to certain assumptions and qualifications as referred to therein.

4.3.1 Jubilee Gold Claims and Royalty

Gold Canyon had acquired ownership of five patented mining claims in 1993 (11229, 11230, 11231, 12868, and 12869) covering a total area of 96.54 ha from Milestone Exploration Limited, a predecessor entity by way of amalgamation with Jubilee Gold Inc. (Jubilee). These claims are subject to a 3% net smelter returns (NSR) royalty on all minerals mined, produced, and sold from these patented claims, provided the monthly average gold price is US$700 or more. The NSR was increased to 5%, together with an NSR of 1 - 2.5% on other adjoining properties in which Gold Canyon conducted any mining operations.

In 2010, Gold Canyon renegotiated the applicable NSR on these patented claims with Jubilee. This agreement terminated any applicable royalty on adjoining claims and set the applicable NSR rate payable upon commencement of commercial production at 3% with advance royalty payments of CDN$70,000/a, adjusted using the yearly Consumer Price Index.

Gold Canyon retained an option to acquire 1% of the NSR for CDN$1,000,000 at any time. As consideration for the renegotiated NSR, it was agreed that previously paid advanced royalties would be forfeited and not credited to any NSR subsequently payable to Jubilee. Gold Canyon paid Jubilee CDN$50,000 and issued 100,000 common shares of Gold Canyon to Jubilee on the date the agreement for the renegotiated NSR was accepted by the TSX Venture Exchange (the Effective Date). In addition, Gold Canyon issued 100,000 common shares to Jubilee on each anniversary date of the Effective Date for the five years that followed. As a result of its acquisition of Gold Canyon, First Mining is subject to this royalty agreement with Jubilee.

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First Mining may terminate all royalty obligations by transferring the patented mining claims back to Jubilee. First Mining retains a right of first refusal on any sale of the remaining royalty interest on certain terms and conditions. The five patented claims identified above are fee simple parcels with mining and surface rights attached to all five claims registered with the Land Registry Office, Kenora, Ontario. First Mining has confirmed via independent legal counsel that the five claims have been surveyed, are in good standing, and the property taxes are paid to date.

4.3.2 Leased Claims from R&S Legacy Inc. and Royalty

First Mining, through its wholly-owned subsidiary Gold Canyon, leases 10 patented mining claims (11233-11235, 12896-12901, and 13043) covering a total area of 182.25 ha. These claims were originally leased to Gold Canyon under the terms of a mineral claims agreement between Gold Canyon and Shirley V. Frahm (Frahm) dated September 22, 2010 (the Original Frahm Agreement). On September 24, 2020, Frahm entered into a purchase and sale agreement with R&S Legacy Inc. (R&S) in furtherance of transferring beneficial ownership to these 10 patented mining claims to her children and grandchildren, and on the same date, Frahm, R&S and Gold Canyon entered into an assignment and assumption of mineral claims agreement (the Assignment Agreement) pursuant to which all of Frahm’s rights and obligations under the Original Frahm Agreement were assigned to R&S, and Gold Canyon consented to such assignment. As a result of the Assignment Agreement, the parties to the Original Frahm Agreement are now Gold Canyon and R&S only.

These 10 patented claims are fee simple parcels with all mining and surface rights attached, and registered, together with the notices of lease, with the Land Registry Office in Kenora, Ontario. Under the Original Frahm Agreement, the lease is for a term of 21 years less one day and terminates on April 14, 2031.

On December 11, 2020, First Mining, Gold Canyon and R&S amended certain provisions of the Original Frahm Agreement by entering into a letter agreement (the Amending Agreement). Pursuant to the Amending Agreement, Gold Canyon paid US$350,000 to R&S as consideration for the amendments, and as a result of the amendments:

Gold Canyon has the irrevocable option to purchase the 10 patented mining claims from R&S from December 11, 2020 until April 15, 2021 (Purchase Option 1) for US$7,000,000 (provided that R&S would still retain a 3% NSR on the claims, unless the NSR buy-back right had been exercised by this time), of which US$1,000,000 may, at Gold Canyon’s option, be satisfied by the issuance of common shares of First Mining (First Mining Shares) to R&S.

Gold Canyon has the irrevocable option to purchase the 10 patented mining claims from R&S from April 16, 2021 until April 15, 2025 (Purchase Option 2) for US$8,000,000 (provided that R&S would still retain a 3% NSR on the claims, unless the NSR buy-back right had been exercised by this time), of which US$2,000,000 may, at Gold Canyon’s option, be satisfied by the issuance of First Mining Shares to R&S.

If, on or before April 15, 2025, First Mining provides R&S with written notice, pays US$250,000 in cash to R&S, and issues 1,000,000 First Mining Shares to R&S, Gold Canyon shall immediately acquire a further irrevocable option to purchase the 10 patented mining claims from April 16, 2025 until April 14, 2031 (Purchase Option 3) for US$10,000,000, less US$250,000 (provided that R&S would still retain a 3% NSR on the claims, unless the NSR buy-back right had been exercised by this time). Of the total purchase price, US$3,000,000 may be satisfied by the issuance of First Mining Shares to R&S.

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·	If, on or before April 14, 2031, First Mining provides R&S with written notice and pays US$2,000,000 in cash to R&S, the 21-year term of the lease under the Original Frahm Agreement shall automatically be extended by five additional years and the new expiry date of the lease will be April 14, 2036. In addition, Gold Canyon would immediately acquire a further irrevocable option to purchase the 10 patented mining claims from R&S from April 15, 2031 until April 14, 2036 (Purchase Option 4) for US$12,000,000 (provided that R&S would still retain a 3% NSR on the claims, unless the NSR buy-back right had been exercised by this time), less US$2,250,000. Of the total purchase price, US$4,000,000 may be satisfied by the issuance of First Mining Shares to R&S.
·	If, on or before April 14, 2036, First Mining provides R&S with written notice and pays a further US$2,000,000 in cash to R&S, then the term of the lease shall automatically be further extended by five additional years and the new expiry date of the lease will be April 14, 2041. In addition, Gold Canyon would immediately acquire a final irrevocable option to purchase the 10 patented mining claims from April 15, 2036 until April 14, 2041 (Purchase Option 5) for US$12,000,000 (provided that R&S would still retain a 3% NSR on the claims, unless the NSR buy-back right had been exercised by this time), less US$4,250,000. Of the total purchase price, US$4,000,000 may be satisfied by the issuance of First Mining Shares to R&S.
·	If at any time during the term of the lease, First Mining commences commercial production, R&S can, by written notice, require Gold Canyon to purchase the 10 patented mining claims for US$12,000,000 (provided that R&S would still retain a 3% NSR on the claims, unless the NSR buy-back right had been exercised by this time), less any cash payments made by Gold Canyon to R&S in connection with Purchase Option 3, Purchase Option 4, and Purchase Option 5. Of the total purchase price, US$4,000,000 may be satisfied by the issuance of First Mining Shares to R&S.
·	If Gold Canyon purchases the 10 patented mining claims from R&S prior to the commencement of commercial production, upon achieving commercial production, Gold Canyon must make a top-up payment to R&S such that R&S would have received an aggregate of US$12,000,000 from Gold Canyon for the claims (after taking into account any amounts previously paid by Gold Canyon to R&S in connection with the various purchase options). This top-up payment can be satisfied through any combination of cash payments and First Mining Shares.
·	Gold Canyon must pay R&S advance royalty payments on a sliding scale of US$33,000/year (2010 – 2011), US$50,000/year (2011 – 2016), US$60,000/year (2016 – 2021), US$100,000/year (2021-2031), and US$120,000/year (2031 – 2041), and all such advance royalty payments shall be deducted from any future NSR payments made by Gold Canyon to R&S.

A 3% NSR with respect to these 10 patented mining claims is payable to R&S upon commencement of commercial production on such claims. During the term of the lease (including if the lease is extended to April 14, 2041), Gold Canyon may, at any time, acquire up to 2% of the NSR for US$1,000,000 per 1%. In addition, during the term of the lease (including if the lease is extended to April 14, 2041), Gold Canyon has the right to access the 10 patented mining claims to conduct mining operations and produce all ores, minerals, and metals that are or may be found therein or thereon, subject to the small portion of the aggregate surface area that has been reserved for recreational use by R&S.

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The remainder of the terms of the Original Frahm Agreement remain unchanged and in full force and effect, including the requirement that Gold Canyon must pay all applicable property taxes related to the 10 patented mining claims during the term of the lease (including if the lease is extended to April 14, 2041), and Gold Canyon maintains a right of first refusal on any sale by R&S of its interest in these 10 patented mining claims on certain terms and conditions set out in the Original Frahm Agreement.

4.3.3 Leased Claims from Springpole Group and Royalty

First Mining has an option and lease to a further 15 patented mining claims (11236, 12867, 12871-12874, 12902-12909) covering a total area of 310.19 ha from a group of individuals and/or companies collectively referred to as the “Springpole Group”. These 15 patented claims are fee simple parcels with mining and surface rights attached to all 15 patented claims registered, together with the notice of option and lease, with the Land Registry Office, Kenora, Ontario. The purchase option was originally granted on September 9, 2004 for a term of five years, and the option can be renewed for successive five-year terms by delivering a written renewal notice to the Springpole Group, along with a renewal fee of US$50,000 and confirmation that at least CDN$300,000 was spent on mining operations in the prior option period. First Mining last renewed the purchase option in September 2024, and the current five-year term of the purchase option expires on September 9, 2029. The current term of the purchase option may be extended by First Mining for further five-year terms by delivering the aforementioned written renewal notice, cash payment and confirmation of expenditures incurred in the prior option period to the Springpole Group.

First Mining is required to make option payments in the aggregate amount of US$35,000/a. and to expend an aggregate of CDN$300,000 on mining operations in each option term as a condition of any renewal and to pay all property taxes related to these patented claims. During the option term, First Mining has been granted the exclusive lease, the right and interest to enter upon the 15 patented claims, the right to conduct mining operations, and the right to have quiet possession thereof. First Mining also has the right, at its discretion, to make any use or uses of the 15 patented claims consistent with the foregoing including the construction of roads, railways, conveyors, plants, buildings, and aircraft landing areas, as well as the alteration of the surface of the Project subject to all applicable laws. First Mining has reserved a small portion of the aggregate surface area for the recreational use of a cabin by the members of the Springpole Group.

First Mining holds an option to acquire the 15 claims and would be required to do so upon the commencement of commercial production at any time during the option period by payment of an aggregate of US$2,000,000. Upon exercise of the purchase option, First Mining must also acquire the cabin on the property for the lesser of fair market value or US$20,000. A 3% NSR is applicable during the option term upon commencement of commercial production or a 1% NSR if the purchase option is exercised prior to commercial production. First Mining can acquire the remaining 1% NSR by a payment of US$500,000.

4.3.4 Claims Leased from the Crown (Mining Claims)

In Ontario, Crown Lands are available to licensed prospectors for the purposes of mineral exploration. A licensed prospector must first register a mining claim to gain the exclusive right to prospect on Crown Land. Claims can also be registered in areas where surface rights are not owned by the Crown if the ground is open for staking and mineral rights can be obtained.

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Traditional claim staking in Ontario came to an end on January 8, 2018, and on April 10, 2018 the Ontario Ministry converted all existing ground or map-staked mining claims (now referred to as ‘legacy’ claims) into one or more cell claims or boundary claims as part of their new provincial grid system. The provincial grid is latitude- and longitude-based and is made up of more than 5.2 M cells ranging in size from 17.7 ha in the north to 24 ha in the south.

Under the new provincial system, the 300 legacy claims at the Springpole Gold Project were converted into single cell claims and boundary cell claims on April 10, 2018. Many of these cell claims were subsequently amalgamated into larger multi-cell claims by First Mining, for ease of administration. The current mining claim fabric held by Gold Canyon, including the Springpole Project Area, consists of 112 multi-cell claims, 26 boundary cell claims and 786 single cell claims, covering an area of 58,157 ha. A full list of these mining claims including townships/area, claim number, claim type, and claim due date is included in Appendix B, and their locations are shown on the land tenure map in Appendix A. Dispositions (such as mining leases and patents) were not converted and remain as they were. Claim registration is governed by the Ontario Mining Act and is administered through the Ontario Mining Lands Administration System (MLAS) which is the electronic system established by the Minister for this purpose.

4.3.5 Mining Leases

Six mining leases (Leases 108953 to 108958, comprising mining claims KRL562895 to KRL562900) plus one related Crown lease for surface rights were acquired by Gold Canyon from an individual in July 2011 for an aggregate payment of US$300,000. Under the original purchase agreement, these claims were subject to a 3% NSR payable upon the commencement of commercial production, with advance royalty payments of US$50,000/a. First Mining retained an option to acquire all or a portion of the applicable NSR at a rate of $500,000 per 1% of the NSR at any time. First Mining permitted the vendor to use a small portion of the property subject to the Crown surface lease, including a vacation home, for recreational purposes provided First Mining was granted a 20-year option to purchase the vacation home for the price determined by an Accredited Appraiser Canadian Institute (AACI) valuator.

Subsequent to the acquisition, these mining leases were set to expire. Following the acquisition, Gold Canyon, as the new lessee, renewed all six leases for a further 21 years, effective September 1, 2011.

In February 2021, First Mining entered into a Purchase and Sale agreement with the owner of the six leases, in which it exercised its option to acquire the vacation home and buy back the 3% NSR.

A further seven mining leases (109846 to 109852) were acquired by Gold Canyon in 2019, by conversion of existing mining claims to mining leases. These leases cover an area of 1,531 ha and have been granted for a 21-year term, effective July 1, 2019.

4.3.6 Claim Maintenance

All mining claims are liable for inspection at any time by the Ministry of Energy and Mines of Ontario (MEM) and may be cancelled for irregularities or fraud in the staking process. Disputes of mining claims by third parties are accepted after one year of the recording date or after the first unit of assessment work is filed and approved. A claim remains valid as long as the claim holder properly completes and files the assessment work as required by the Mining Act and the Minister approves the assessment work.

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To keep a mining claim current, the mining claim holder must perform C$400 per single cell mining claim unit worth of approved assessment work per year, or C$200 per boundary cell mining claim unit, immediately following the initial registration date. The claim holder has two years to file one-year worth of assessment work.

Surface rights are separate from mining rights. Should any method of mining be appropriate, other than those claims for which Crown leases were issued, the surface rights would need to be secured.

All claims registered to Gold Canyon are in good standing as of the effective date of this report.

4.3.7 Silver Stream with First Majestic Silver Corp.

First Mining, Gold Canyon, and First Majestic Silver Corp. (First Majestic) entered into a silver purchase agreement dated June 10, 2020 (the Silver Purchase Agreement) pursuant to which First Majestic agreed to pay US$22,500,000 to First Mining over three tranches for the right to purchase 50% of the payable silver produced from Springpole for the life of the Project (the silver stream). First Majestic made the first payment of US$10,000,000 to First Mining when the transaction closed on July 2, 2020, consisting of US$2,500,000 in cash and US$7,500,000 in common shares of First Majestic (First Majestic Shares). A further US$7,500,000 was paid by First Majestic upon the announcement by First Mining of the completion of the Springpole PFS, with US$3,750,000 paid in cash and US$3,750,000 paid in First Majestic Shares. First Majestic will make a final payment of US$5,000,000 to First Mining upon the earlier receipt by First Mining of approval of a federal or provincial EA for Springpole, with US$2,500,000 of this amount payable in cash, and the remaining US$2,500,000 payable in First Majestic Shares.

In addition, in return for the silver stream, following the commencement of production at Springpole, First Majestic will make ongoing cash payments to First Mining equal to 33% of the lesser of the average spot price of silver for the applicable calendar quarter, and the spot price of silver at the time of delivery, subject to a price cap of US$US$7.50 per ounce of silver (the price cap). The price cap is subject to annual inflation escalation of 2%, starting at the third anniversary of the commencement of production at Springpole.

Under the terms of the Silver Purchase Agreement, First Mining has the right to repurchase 50% of the Silver Stream for US$22,500,000 at any time prior to the commencement of production at Springpole (following such repurchase, First Majestic would be left with a right to purchase 25% of payable silver produced from Springpole). First Mining also granted First Majestic a right of first refusal with respect to any future silver stream financings related to Springpole, and First Mining issued 30,000,0000 common share purchase warrants (warrants) to First Majestic on closing of the transaction, with each warrant entitling First Majestic to purchase one First Mining Share at an exercise price of C$0.40 for a period of five years.

An amending agreement to the Silver Purchase Agreement was executed on March 13, 2025, which amended the terms of the final payment due from First Majestic, such that the Tranche 3 payment would be for US$5 million in cash. As consideration for amending the terms of the Tranche 3 payment, the Company has amended the terms of the common share purchase warrants that were issued to First Majestic on July 2, 2020, under the terms of the Silver Purchase Agreement.

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4.3.8 Royalties Assumptions for Mine Planning and Economic Evaluation

For the purposes of mine planning and economic evaluation, it has been assumed that all buy-back options would be exercised prior to production in accordance with the terms of the agreements summarized in this section.

4.4 Environmental Considerations

Regularized procedure reviews and operational audits are used to ensure best practice methods are applied to mineral exploration at the Springpole Gold Project. Improvements to critical areas that affect the environment are underway at all times in an attempt to reduce and reclaim the environmental footprint of exploration activities. No material environmental liabilities or public hazards associated with the Springpole Gold Project are known to exist on the property. A temporary camp (~0.5 ha) with wood-frame tents was erected for ongoing drilling campaigns, with the tents replaced in recent years by modular, hard walled, dormitory-style accommodation. There has been occasional surface clearing related to drilling work.

4.5 Permitting Considerations

First Mining complies with permit, notice and consultation requirements as they relate to the ongoing exploration work on the Springpole Gold Project. Legislation that requires material permits and notices include the provincial Mining Act, Public Lands Act, Lakes and Rivers Improvement Act, Ontario Water Resources Act, as well as the federal Fisheries Act.

Information on permitting required for Project development is provided in Section 20.

4.6 Social License Considerations

Information on social license considerations for the Project is provided in Section 20.

4.7 Project Risks and Uncertainties

To the extent known to the QPs, there are no other significant factors and risks that may affect access, title, or the right or ability to perform work on the Project that are not discussed in this report.

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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Accessibility

During late spring, summer, and early fall, the Springpole Gold Project is accessible by floatplane direct to Springpole Lake or Birch Lake from the communities of Red Lake, Ear Falls and Sioux Lookout, Ontario. In some years it has been accessed by a seasonal (January – March) winter road constructed over a distance of 40 km, commencing from the 85 km mark of the South Bay Mine Road located east of Ear Falls, Ontario. The closest road access at present is approximately 15 km away at the extension of the Wenasaga forestry access road (Figure 4-2).

During breakup in spring and freeze-up in fall, access to the Project is by helicopter. Additional winter access may be available via temporary airstrips cleared on nearby frozen lakes.

5.2 Local Resources and Infrastructure

There is no existing infrastructure within 20 km of the Springpole Gold Project area. Businesses in Red Lake, a long-established mining community 110 km to the southwest, as well as in the town of Sioux Lookout, provide the majority of the camp’s supply needs. The nearest emergency medical facilities are at the Margaret Cochenour Hospital in Red Lake. The nearest major city is Winnipeg, Manitoba, which is approximately 370 km southwest of the Project and about a 1.3-hour flight by Cessna Caravan. The nearest power line to the Project is currently the 115 kV E1C transmission line owned by Hydro One Networks Inc. which runs from Ear Falls to Pickle Lake on a right of way located approximately 26 km to the south of Springpole, and which supplies power to the communities of Cat Lake, Slate Falls, and Pickle Lake as well as Newmont’s Musselwhite Mine.

Water and infrastructure site requirements are discussed further in Section 18.

5.3 Climate and Physiography

January temperatures at the Project range between -40°C and 0°C, and July temperatures range between 20°C and 40°C. Exploration and mining activities can be carried out year-round.

Springpole and Birch Lakes are part of an extensive watershed which flows eastward into the Cat River. The Project is underlain by glaciated terrain characteristic of a large part of the Canadian Shield. Land areas are generally of low relief with less than 30 m of local elevation and are separated by a series of interconnected, shallow lakes. The average elevation of the property above sea level is 395 m.

Tree cover consists of mature spruce, balsam, birch, and poplar. Black spruce and muskeg swamps occupy low-lying areas. Glacial till is generally less than 1 m in thickness. Outcrops are limited and small and are generally covered by a thick layer of moss or muskeg.

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6 HISTORY

6.1 Regional History of the Area

The history of the Springpole Gold Project prior to 2006 is summarized from the Technical Report and Resource Estimate on the Springpole Lake Gold Property (Armstrong et al., 2006).

Gold exploration in the area was carried out during two main periods, one during the 1920s to 1940s, and a second period from 1985 to the present. The following information in this section refers to mineralization that includes areas outside the current Gold Canyon property boundaries.

The Casey Summit Mine (later renamed the Argosy then Casummit Mine), approximately 10 km to the north of the Springpole deposit, started operation around this time. This mine ultimately produced 101,975 oz of gold and 9,788 oz of silver (Beakhouse, 1990) and is the only significant past producer of precious metals in the Birch-Springpole Lake area.

This early prospecting activity and production from the Casummit Mine region prompted a more detailed geological investigation of the vicinity by the Ontario Department of Mines. The Birch Lake area was mapped at a scale of 1:63,360 by Harding (1936).

Reconnaissance-style mapping of the Birch Springpole Lake area, which includes the Springpole Project, has since been repeated four times by the Ontario Government:

1.	To study volcanic characteristics of selected Superior Province greenstone belts (Ontario Department of Mines -Goodwin, 1967)
2.	To extend volcanic stratigraphy hosting the South Bay base metal mine into the Springpole area (Ontario Geological Survey - Thurston et al., 1981)
3.	To stimulate gold exploration in the area after closure of several mines near Red Lake (Ontario Geological Survey - Good et al., 1988)
4.	To study the stratigraphy of epiclastic and volcaniclastic facies units, northern Birch-Uchi greenstone belt (Ontario Geological Survey - Devaney, 2001a)

The historical exploration areas discussed above are shown on Figure 6‑1.

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Figure 6‑1: Historical Exploration Summary Map, 1928 - 2001

Source: First Mining, 2025

6.2 History of Exploration on the Springpole Property

In 1925, the discovery of gold at Red Lake brought prospectors into the Springpole Lake area. Gold in outcrop on the Project was first discovered in the Birch-Springpole Lake portage area and prospected by Northern Aerial Mineral Exploration Ltd. in 1928 (Harding, 1936). The showing was initially covered with eight claims around 1934 by prospector Tom Dunkin, who then completed the first stripping and shallow trenching that same year (Figure 6‑1).

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Between 1933 and 1936, the Windigokan Sturgeon Mining Syndicate conducted extensive trenching and prospecting, including 10 short holes totalling 458.5 m. The claims were then transferred to Springpole Mines Ltd. who carried out limited trenching and prospecting in 1945.

The area remained dormant until 1985 when Gold Fields Canadian Mining Ltd. (GFCM) optioned the Frahm claims and, in 1986, the Milestone claims and Maple Leaf (now Springpole Group) claims. GFCM conducted an airborne (Aerodat) geophysical survey in 1985 over the entire claim group. On the 30 patented mining claims (Frahm, Milestone, and Springpole Group), line cutting was done at both 30.5 m (100 ft) centres (Milestone claims) and 61 m (200 ft) centres (Frahm and Springpole Group claims). Subsequently, geological mapping, humus geochemistry, and ground geophysics (very low frequency (VLF), magnetic (Mag), and IP) were conducted over the grids.

From 1986 through 1989, GFCM completed 118 diamond drill holes in seven drill phases totalling 38,349 m. In addition, during 1986 and 1987, approximately 116,119 m^2^ of mechanical stripping was carried out by GFCM, and four petrographic reports were produced. As a result of this work, GFCM identified several gold-bearing zones on the Project site that included:

·	Portage zone, entirely under a portion of Springpole Lake, but the largest of the zones and therefore the main focus of the bulk of the exploration work.
·	Jasper zone, a deep narrow higher-grade zone in a banded iron formation horizon.
·	Several smaller but higher-grade zones on the land portion of the Project and close to surface, including the Main zone, Vein zone, Hillside zone, Camp zone, North Porphyry zone and East Extension zone.

Late in 1989, GFCM entered into a 50/50 joint venture with the combined interests of Noranda and Akiko-Lori.

From 1989 through 1992, Noranda conducted an IP survey over the central portion of the Portage zone and tested the property with eighteen core holes totalling 5,993 m. The majority of the drilling was conducted on the Portage zone.

At the same time, and under a separate option agreement with BP Resources Canada, Noranda completed a seven-core hole drill program around the east margins of Springpole Lake on claims then owned by BP Resources. BP Resources in turn completed lake-bottom sediment sampling of Springpole Lake east of Johnson Island.

In 1992, Noranda dropped its interest in the property leaving Akiko-Lori to carry out further exploration while carrying its 50% partnership with GFCM. During 1992 to 1994, Akiko-Lori/Akiko Gold completed an additional 15 diamond drill holes totalling 5,154 m.

By 1995, Akiko Gold was reorganized into Gold Canyon and GFCM's interest was acquired by Santa Fe Mining as part of an asset exchange with London based Hanson Plc., which controlled GFCM. During 1995, a joint venture between Gold Canyon and Santa Fe carried out an exploration program consisting of remapping of the main area, of some of the existing drill core, and a reinterpretation of the geology.

During the 1995 and 1996 programs, Santa Fe drilled an additional 69 holes totalling 15,085 m on the Springpole Gold Project proper and two drill holes on Johnson Island. By late 1996, the takeover of Santa Fe by Newmont Gold Company was nearing completion. Just prior to the merger with Newmont, Santa Fe exchanged their 50% interest in the property for a tax credit that left Gold Canyon with a 100% ownership. After Santa Fe’s departure, Gold Canyon continued exploration at the Springpole Gold Project in 1997 and 1998 with another 52 core holes totalling 5,643 m.

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Paso Rico had an option to earn an interest in the Project and, in the summer of 1998, conducted with Gold Canyon a lake bottom sediment sampling program in several areas of Springpole Lake. The results of this survey identified several follow-up targets that were tested in 1999 by Paso Rico with 12 core holes totalling 2,779 m. In 2000, Paso Rico withdrew from the Project leaving Gold Canyon with its current 100% interest.

Between 2004 and 2013, diamond drilling programs were conducted on the property by Gold Canyon under its previous management. A total of 371 holes were completed during this period, over 108,932 m.

Subsequent to the acquisition of Gold Canyon by First Mining in 2015, drilling campaigns were completed in 2016, 2018, 2020, 2021, 2022 and 2024, the majority of which was for site investigation purposes (e.g. geotechnical, hydrogeological, metallurgical).

A summary of the drilling completed on the Property between 1986 and 2024 is provided in Table 6‑1. Drilling activities from 2007 to 2024 are discussed in more detail in Section 10 of this report.

Table 6‑1 : Summary of Drilling at Springpole from 1986-2024

Company	Period	Number of Holes	Metres Drilled

| Goldfields Canadian Mining Ltd. | 1986-1989 | 116 | 38,496 |

| Noranda/Akiko JV | 1990-1991 | 17 | 5,530 |

| Akiko Gold Resources Ltd. | 1993-1994 | 6 | 3,066 |

| Santa Fe Canadian/Gold Canyon Resources Inc. JV | 1995-1996 | 68 | 14406 |

| Gold Canyon Resources Inc. | 1997-1998 | 51 | 5,376 |

| Paso Rico | 1999 | 12 | 2,779 |

| Gold Canyon Resources Inc. | 2004-2013 | 371 | 108,932 |

| First Mining Finance Corp. | 2016 | 4 | 1,712 |

| First Mining Gold Corp. | 2018-2024 | 208 | 26,373 |

| Total | | 853 | 206,671 |

6.3 Acquisition by First Mining

On November 13, 2015, First Mining (which was called First Mining Finance Corp. at the time) completed the acquisition of Gold Canyon, and as a result, acquired the Springpole Gold Project. The Company changed its name from First Mining Finance Corp. to First Mining Gold Corp. in January 2018.

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7 GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology

The following excerpt is quoted from Devaney (2001b) and provides the most concise geologic description of the regional geology of the Springpole-Birch Lake area:

“The Birch-Uchi Greenstone Belt is the portion of the Uchi Sub-province with an arcuate, concave to the southeast, (i.e., a major oroclinal bend between the Red Lake and Meen-Dempster portions of the sub-province). Studies of the southern part of the Birch-Uchi greenstone belt as a rootless greenstone belt only a few kilometres thick, have revealed a long (ca. 3.0 to 2.7 Ga), multistage history of crustal development. Based on mapping, lithogeochemistry, and radiometric dating, the supracrustal rocks of the greenstone belt were subdivided into three stratigraphic group-scale units (listed in decreasing age): the Balmer, Woman and Confederation assemblages. This three-part subdivision was applied to most of the Uchi Subprovince. The Confederation assemblage is thought to be a continental margin (Andean-type) arc succession, versus the less certain tectono-stratigraphic context of the other assemblages. Workers performing recent and ongoing studies of the southern Birch-Uchi greenstone belt and the Red Lake greenstone belt (i.e., the Western Uchi Subprovince NATMAP Project) have proposed some modifications and additions to the Balmer-Woman-Confederation stratigraphic scheme. As discussed herein, some relatively small conglomeratic units likely form a synorogenic, discontinuously distributed, post-Confederation assemblage in the Birch-Uchi greenstone belt. Radiometrically dated plutons within the Birch-Uchi greenstone belt are of post-Confederation assemblage, ca. 2725-2700 Ma age.

The northern margin of the Birch-Uchi greenstone belt forms a pattern of sub-regional scale cusps of supracrustal strata alternating with batholiths. Basaltic units are prominent around the periphery of the greenstone belt and may be part of the Woman assemblage, but the accuracy of this stratigraphic assignment is unknown. Based on a ca. 2740 Ma age of Shabumeni Lake [intermediate to felsic fragmental] volcanic rocks at a site near the northern greenstone belt margin, suggested that Confederation assemblage age rocks make up the bulk of the greenstone belt”.

The regional geological setting of the Springpole Gold Project is illustrated in Figure 7‑1.

It is noteworthy that in many of the regional geology descriptions of the Birch-Uchi Greenstone Belt, especially those in the vicinity of Springpole and Birch Lakes, the structural geology is lesser understood than the neighbouring Red Lake Greenstone Belt. Many authors make relatively brief mention of the complexities that dominate the geology and geomorphology of the low-lying areas. However, the Archean Orogenic gold deposit model developed by various authors has been applied to the mineral deposits of the Archean Superior Province. Concise summaries of these orogenic gold deposits can be found in: Groves et al., (1998), Hagemann and Cassidy (2000), Aufarb et al., (2005), and Robert et al., (2005).

Orogenic gold deposits are epigenetic, structurally-controlled gold deposits that are hosted in orogenic belts. They are generally accepted as having formed during late stages of continental collision. Most of the discovered orogenic gold deposits in the world occur in greenstone belts situated on the margins or within Archean cratons in North America, Australia, and southern Africa.

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Figure 7‑1: Regional Geology Plan

Source: First Mining, 2025

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7.2 Project Geology

The Gold Canyon tenure is located over supracrustal rocks of the Birch-Uchi Greenstone Belt, dominantly comprised of volcanic and sedimentary units with syntectonic intrusive units (Devaney, 2001). The property geology has been divided into sub-properties as described below.

7.2.1 Springpole Project Geology

The Project has been extensively studied during past programs and the findings of those studies will not be covered in detail here; however, they are adequately covered in the technical reports of Zabev (2004) and Armstrong et al., (2006).

The following subsections summarize the geology interpreted from field observations and petrographic analysis of drill core from the 2009 re-logging program and from drill core produced during the 2010 to 2024 programs.

7.2.1.1 Trachyte Porphyry Intrusion

A polyphase alkali, trachyte intrusion displaying autolithic breccia textures lies at the heart of the Springpole Gold Project. The intrusion is comprised of a system of multiple phases of trachyte believed to be part of the roof zone of a larger syenite intrusion, as fragments displaying phaneritic textures were observed from deeper drill cores in the southeast portion of the Portage zone. Early intrusive phases consist of megacrystic feldspar phenocrysts, up to 5 cm long, of albite and orthoclase feldspar in an aphanitic groundmass. Successive phases show progressively finer-grained porphyritic texture while the final intrusive phases are aphanitic.

In 2009 and 2010, Gold Canyon carried out petrographic studies (Saunders and McIntosh 2009, 2010) of historical drill core and drill core from the drill holes SP10-001 through SP11-006. The study confirmed the trachyte intrusion is the dominant lithology within the Project area and is a host to mineralization. Interpretation of the intrusive complex is complicated by a mixture of overprinted regional and local metamorphic events related to burial and tectonism.

Pervasive alteration and metamorphism have reduced the original porphyry intrusion to a complex alteration assemblage dominated by sericite, biotite, pyrite, calcite/dolomite, and quartz. Primary igneous textures are remarkably well preserved in places and give indications to the possible genesis of the initial phase of gold mineralization. Within the country rocks to the north and east are trachyte and lamprophyre dikes and sills that source from the trachyte - or syenite-porphyry intrusive system.

7.2.1.2 Confederation Age Volcanic and Siliciclastic Rocks

The country rocks pre-date the alkali intrusion and are composed of a geologic sequence of altered and metamorphosed intermediate andesitic volcanic and associated volcaniclastic rocks, siliciclastic sedimentary rocks, chemical sediments including banded iron formation (BIF), and coarse pebble conglomerates. Devaney (2001a) indicates that the sediments are likely of the Confederation assemblage dating at around 2,740 Ma, representing the proximal portions of a mixed volcanic-sedimentary basin.

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7.2.1.3 “Timiskaming-type” Conglomerates

Barron (1996) states pebble conglomerate outcrops between Springpole Lake and Birch Lake contain clasts of the trachyte porphyry, suggesting that the “Timiskaming-type” conglomerates postdate the intrusion. Devaney (2001a) suggests this arcuate form of conglomerate represent late orogenic, deformed, dextral sense strike-slip (pull-apart) basins of “Timiskaming-type”, late Archean, post Confederation assemblage age rocks.

7.2.2 Birch-Uchi Regional Targets Geology

A summary of the geology of some of First Mining’s regional targets within the Birch-Uchi Greenstone Belt is below, and target locations are shown on Figure 7‑2.

The Horseshoe area is located approximately 10 km directly west of the Springpole deposit and is underlain by dominantly intermediate volcanic rocks varying from flow to tuffaceous stratigraphy which have seen significant deformation by the Swain Deformation zone. A thin unconformable band of massive mafic volcanic flow separates the tuffaceous units from the massive intermediate volcanic stratigraphy. The volcanic units are intruded by a multiphase tonalitic intrusion which stretches across the stratigraphy and hosts the Horseshoe Island showing.

The Satterly area is located approximately 12 km southwest of the Springpole deposit and 5 km south of the Horseshoe area. The area is dominantly underlain by thinly bedded wacke (turbidite) lithology with clearly younging direction indicators and weak to moderate foliation. The turbidite country rock is intruded by a granodiorite stock (Greencamp Stock).

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Figure 7‑2: Bedrock Geology and Target Locations, Gold Canyon Property

Source: First Mining, 2025

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7.3 Structure

The Birch-Uchi Greenstone Belt shows evidence of substantial rifting and associated depositional and magmatic phases. At least three phases of regional deformation affected the area. These deformation events pertain to a D1 northwest-southeast-striking shortening event, a D2 northeast-southwest shortening event with east to west striking folds, and a D3 east-west shortening event that is mainly restricted to the northern part of the belt (Devaney, 2001).

Regional deformation has added complexity to the apparent geometry of, and the potential of, the Springpole gold deposit. Gravity, magnetic, and field surveys carried out across the Project demonstrate that several phases of deformation are evident. Banded iron formations describe north-northwest facing tight to isoclinal antiforms and synforms and are illustrated on the geological map produced during the summer 2005 mapping program (Armstrong et al., 2006) and are evident as strong magnetic anomalies on the aeromagnetic surveys conducted by Fugro.

In 2011, SRK was contracted to carry out a preliminary study of the structural controls on mineralized deposit geometry. The study found the deposit was subjected to several deformational events including, but not limited to:

·	early folding resulting in tight to isoclinal fold geometries and development of associated shear zones
·	intermediate large scale, potentially deep-rooted shear zones
·	late-stage brittle faulting

In 2018 and 2023, structural studies were further advanced by Ibex Exploration, evolving relationships of deposit paragenesis with current deposit geometry. One of the important conclusions of the studies was that the Springpole rocks have been rotated roughly 60° about a N51W (azimuth 309°) axis since the time of initial gold deposition (Price, 2024).

7.4 Alteration

All rocks on the Project exhibit pervasive alteration that consists of multiple overprinted phases. Distinguishing between the individual phases takes considerable study on a microscopic scale. The country rocks and alkali intrusive rocks exhibit pervasive green-schist facies metamorphism and alteration, probably the result of burial. This manifests as chlorite, calcite, and pyrite in the intermediate volcanic rocks, pyritization of the banded iron formation, and sericite-pyrite alteration within the alkali intrusive associated rocks.

Studies conducted as a part of the exploration work carried out from the fall of 2009 and the winter/spring of 2010 show there is evidence of early alteration phases. These probably resulted from magmatic hydrothermal fluids associated with porphyry gold mineralization and the associated epithermal/mesothermal style gold mineralization. This occurs as potassic and phyllic/sericitic alteration: K-feldspar, biotite, and muscovite (sericite), respectively, and is nearly pervasive in the alkali intrusive rocks and surrounding country rocks. Regional metamorphism has subsequently altered the primary hydrothermal mineral assemblages, but textures have been preserved with the exception of areas of high strain (e.g., northwest trending shear zones).

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Advanced argillic alteration appears throughout the trachyte intrusive and occurs in some of the late stage lamprophyre dikes though on a small scale. It is difficult to assess at what stage argillic alteration occurs, but it appears to define an envelope around the Portage zone potassic-alteration/mineralization, suggesting an origin more in keeping with zoned alteration associated with epithermal-style porphyry intrusive hosted gold deposits.

7.5 Mineralization

Three separate mineralized zones have been modelled at the Springpole Deposit, each with their own characteristics. The Portage zone, which is hosted within volcanic trachyte and heterolithic breccia dominated by potassic (biotite) alteration and pyrite; the Camp zone, largely hosted in sheared mafic volcanics with intense ankerite and sericite alteration associated with more lode gold style of mineralization; and the East Extension zone which is associated with intrusive megacrystic trachyte and heterolithic breccia, with similar potassic (biotite) alteration and pyrite mineralization style as the Portage zone.

Apart from the Camp zone, which is interpreted to be hosted in the volcanic basement rocks, the Portage and East Extension zones are interpreted to be hosted in an unconformable package of trachytic volcanics varying from aphanitic to medium-grained porphyritic with discernible volcanic sequences. The basement volcanic sequence is comprised of subaqueous mafic volcanic flows which have been intruded by gabbroic dikes/sills. Ancillary mineralization is noted within these gabbroic units and is largely characterized as ankerite-altered shear zones with sulphide replacement mineralization associated with quartz-carbonate veining. Megacrystic trachyte intrusions are largely confined to the unconformable package of trachytic volcanic rocks and represent the porphyry-style of mineralization outlined in the Portage and East Extension zones. Intermediate volcanic flows overlay the deposit and are structurally unconformable from the main package of trachytic volcanics. These intermediate units do not display the typical chaotic volcanic stratigraphy as seen in the core mineralization footprint. Timiskaming-like conglomerates overlay all stratigraphic units previous and cobbles are often comprised of pre-dated rocks mentioned above.

7.5.1 Porphyry-style Mineralization

The main intrusive complex appears to contain many of the characteristics of alkaline, porphyry-style mineralization associated with diatreme breccias (e.g., Cripple Creek, Colorado). Direct comparison with drill core from the two sites shows several consistent textures and styles of mineralization. Observations made from drilling, combined with the airborne magnetic surveys, show that gold mineralization is largely coincident with discrete geophysical low anomalies. This style of mineralization is characterized by the Portage zone and portions of the East Extension zone where mineralization is hosted by diatreme breccia in aphanitic trachyte. Ductile shearing and brittle faulting have played a role in redistributing structurally controlled blocks of the mineralized rock. Work continues to outline a consistent form of porphyry style alteration zoning consisting of an outer zone of phyllic (sericite) dominant alteration with narrow zones of advanced argillic alteration characterized by illite and kaolinite, and a core zone of intense potassic alteration characterized by biotite and K-feldspar.

Multi-element analysis conducted during the 1992 program on the Portage zone, combined with gold assays, gave the first indication of the style of mineralization at Springpole. Diamond drilling in the winter of 2010 revealed a more complex alteration with broader, intense zones of potassic alteration replacing the original rock mass with biotite and pyrite. In the core area of the deposit where fine-grained, disseminated gold mineralization occurs with biotite, the primary potassic alteration mineral, gold, displays a good correlation with potassium/rubidium.

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7.5.2 Lode Gold Mineralization

The intrusion of the trachyte complex into the volcanic pile, as well as the chemical and siliciclastic sedimentary rocks in a near surface environment, produced mesothermal to epithermal style lode vein mineralization. The difference between mesothermal and epithermal mineralization regimes is the temperature and pressure of the mineralizing fluids.

Higher temperature (mesothermal) fluids would have existed within the emplaced intrusive, associated with the diatreme breccias, and in the immediately adjacent wall rock/country rocks. In the porphyry intrusive, and at the contact between intrusive and wall rock in the East Extension zone, and localized within the Main zone, mesothermal style quartz-biotite-calcite-sulphide veins with occasional tourmaline are observed with occasional coarse, visible gold.

Further from the intrusive complex and wall rock contact zones, where meteoric fluids have a greater influence, epithermal style vein textures and mineralization styles dominate. These consist of banded to sucrosic quartz calcite veins with a lower temperature mineral assemblage including sericite, minor biotite, possible adularia, calcite, dolomite, and ankerite; here gold, silver, and tellurium alloys dominate, including electrum and gold-silver tellurides.

7.5.3 Gold Remobilization During Metamorphism

As evidenced from the high degree of deformation, both ductile and brittle in the form of isoclinal folding, ductile shear zones with protomylonite and blastomylonite textures, and brittle fault textures the Springpole deposit has been subjected to alteration and metamorphism. These processes alone have remobilized gold in epithermal quartz veins that were the principal motivation for exploration at the Springpole Gold Project in the late 1980s and early 1990s, when shear zone hosted gold deposits were the targets of choice in the Red Lake area.

Regional targets across the Gold Canyon tenure and outside of the Springpole Project area exhibit the following gold mineralization styles: low-sulphidation epithermal gold and silver; orogenic gold; and intrusion-related gold.

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8 DEPOSIT TYPES

8.1 Deposit Model

Mineralization at the Springpole Gold project is dominated by large tonnage, low grade, disseminated porphyry-style or epithermal-style gold mineralization associated with the emplacement of an alkali trachyte intrusion. Textures observed in the extensive repository of drill core appear to confirm that the disseminated gold-silver-sulphide mineralization, the mesothermal to epithermal lode vein gold mineralization, and the banded iron-formation hosted gold mineralization are all the result of the emplacement of multiple phases of trachyte porphyry and associated diatreme breccias, hydrothermal breccias, dikes and sills.

The initial exploration on the property was conducted on the assumption the mineralization was a typical example of Archean mesothermal, sulphide-hosted lode gold type. With ongoing study, a departure from Archean mesothermal systems has been replaced in favor of a high-level emplacement porphyry model. Barron’s thesis (1996) work presented strong evidence that the gold and associated fluorite mineralization at Springpole is genetically related to the high-level emplacement of a large, alkaline porphyry intrusive and breccia pipe complex.

Barron considered the Springpole Complex to be the end product of magmatic fractionation processes and of fluids that evolved from magmatic to hydrothermal in the high level, sub-volcanic porphyry environment. These processes produced a low-grade gold-porphyry-epithermal type deposit and associated high-grade veins and breccia pipes.

Santa Fe geologists felt the nature of the mineralization at Springpole had many similarities with deposits of the Cripple Creek District, Colorado, including the Cresson Mine. Detailed mapping on the land-based portions of the property by Santa Fe geologists showed that most, if not all, of the gold mineralization at the Springpole Gold Project is spatially associated with the feldspar porphyry diatreme dikes, veins, and diatreme breccia. The following is a brief description of this model in the Springpole area.

8.2 Depositional Environment

Based upon the abundance and size of epizonal trachyte porphyry intrusive masses and the widespread brecciation and alteration centered on the Portage zone, Barron (1996) considered this area to be the apex of a buried syenite stock. A high emplacement level for the Portage zone and surrounding porphyry is further supported by the lack of contact metamorphic effects in the enclosing country rocks. Trachyte clasts within the basal conglomerate overlying the intrusive complex indicate it was subjected to surface erosion.

The rarity of trachyte clasts and their restriction to the base of the conglomerate unit would seem to indicate erosion over a short time interval. The lack of voluminous trachyte flows suggests there was no markedly positive volcanic edifice. Barron (1996) concluded that collectively these features suggested that the Portage zone and surrounding Main and East Extension zones existed as a small island of maar craters of low relief in a rapidly deepening shallow basin.

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This interpretation has its closest modern analogue in the Ladolam Gold Deposit, Lihir Island, Papua New Guinea. Mineralization at Lihir is believed to be less than 500,000 years old and is telescoped upon an earlier porphyry environment (Carman, 2003). Deposition of gold is still an active process at Ladolam as the hydrothermal system remains active. Host rocks at Ladolam can be divided into three groups:

·	Mafic lavas composed of alkali basalt, porphyritic trachybasalt, trachyandesite, and rare trachyte and phonolite.
·	Alkali intrusions that are composed of multi-phase porphyry stocks with the most voluminous phase being biotite monzonite.
·	Ladolam Breccia Complex that is composed of porphyry breccias and volcanic breccias.

Porphyry breccias are dominantly monzonite composition and occur as poorly sorted, massive, matrix supported breccias with some rounding of clasts caused by magmatic milling; the clasts are supported by a cement of altered rock flour and anhydrite. The volcanic breccias are massive, moderately to poorly sorted, rock flour matrix supported breccias containing mafic clasts.

Mineralization/alteration at Ladolam can also be sub-divided into three broad phases:

·	Biotite-orthoclase-anhydrite ± magnetite with minor copper-gold-molybdenum disseminated porphyry mineralization and veinlets.
·	Refractory sulphide-gold mineralization associated with pervasive adularia-pyrite (leucoxene-illite) alteration near surface that comprises the bulk of the near surface bulk mineable mineralized material.
·	Quartz-calcite-adularia-pyrite-marcasite ±electrum stockwork veins.

If the Ladolam Gold deposit is accepted as a reasonable genetic analogue to the Springpole deposit, then the following genetic model can be applied. This model is adapted from Barron’s thesis (1996), Zabev’s genetic summary (2004), and the genetic model of Armstrong et al., (2006), as well as observations made during the 2009 through 2012 diamond drilling programs.

8.2.1 Springpole Genetic Model

The following list summarizes the genetic model of the Springpole Gold Project area:

·	Intrusion into the lower crust of parental alkaline primitive and anhydrous magma slightly enriched in incompatible elements including fluorine.
·	Fractionation at depth, precipitation of hornblende and apatite as early crystalline phases; the magma becomes increasingly anhydrous. Gold is retained in the melt.
·	Diapiric uprise from 4 to 8 km levels into hydrous wall rock with the apex of the magma chamber at <2 km depth. Continued fractionation producing an increasingly fluorine-rich melt; feldspar of extreme composition is precipitated, and the lowered solidus allows emplacement of porphyry dikes and sills to very high crustal levels.

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·	High diffusivities and convection promotes water partitioning from wall rock into magma.
·	The magma is quickly saturated, and the sudden pressure is released (possibly from venting) prompting the immiscible separation of fluorine and carbon dioxide-rich phases, which escapes to high structural levels. Breccia pipes with rock fluorite and rounded clasts indicating turbulent fluidized and erosional vertical emplacement.
·	Fluid pressures generate dike offshoots.
·	Fluorine escapes from brecciated wall-rock causing biotization or fluoritization of breccia and wall rock. Ultimately, the fluorine-water-carbon dioxide vapours condense, resulting in the precipitation of fluorite and calcite. Magmatic gold-rich fluids permeate the breccia and surrounding porphyry, depositing porphyry style, disseminated, pyritic mineralization. The fractures along the margins of breccia pipes acts as preferred sites for later deposition of quartz, electrum, and tellurides.
·	Intrusion of a series of lamprophyre and carbonatite dikes, sills, and veinlets—due to the intensity of deformation.
·	The complex is then buried by conglomerates derived from the complex and other areas (Devaney, 2001b).
·	Continued intense deformation and associated metamorphism manifesting as folding, strike-slip faulting and shearing, coupled with regional green schist metamorphism of the region obscures primary textures and likely leads to some (possibly minor) degree of precious metal remobilization.

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9 EXPLORATION

9.1 Introduction

Between 2021 and 2025, First Mining acquired an additional 15,895 ha of mineral tenure adjacent to the Springpole Project in the Birch-Uchi Greenstone Belt to advance regional scale exploration opportunities. First Mining has completed several exploration programs between 2021 and 2025 over the newly-acquired tenure, as well as the Project area. Exploration activities consisted of prospecting campaigns on regional targets, an airborne electromagnetic and magnetic survey, and detailed geological mapping and sampling over select targets on the Springpole Project. A summary of the work programs is provided in sections to follow. Regional drilling programs were completed at the Swain, Saddle and Horseshoe targets in 2022 and 2023, and this work is discussed further in Section 10.

Exploration prior to 2024 on the Swain and Swain Post claim groups was completed when these properties were under option agreements between First Mining and the registered owners. These option agreements were completed in February 2024 at Swain Post and in April 2024 at Swain, after which Gold Canyon acquired 100% ownership of Swain Post, and 70% ownership of Swain.

A summary of any historical exploration completed on the property can be found in Section 6 of this report.

9.2 Mapping and Sampling Programs, 2021 to 2025

9.2.1 2021 Program

In 2021, Equity Exploration Consultants Ltd. was contracted by First Mining to complete an initial field prospecting and reconnaissance campaign across First Mining’s mineral tenure in the Birch-Uchi Greenstone Belt. This reconnaissance work program consisted of a desktop study and GIS compilation of data comprised of government geological mapping, assessment report maps and prospect descriptions, as well as First Mining mapping and prospecting data. In the subsequent field component of the program, 45 gold occurrences were visited, evaluated, and sampled where favourable mineralization was observed, the occurrence sites were grouped into nine mineral target camps that share similar features (e.g., location, geology, and mineralization style) and formed the basis for follow-up exploration campaigns.

The mapping and sampling campaign was targeted to follow up on historical mineralization and alteration showings, and a total of 76 rock grab samples were collected and geological mapping was completed on 194 outcrops. This work program highlighted the Satterly area (including Greencamp, Rice, and Trench Grid Targets), Sol D’Or, Sandy Point, and Wagner targets as areas with a higher potential for gold prospects and deposits.

Highlights included grab samples at Rice (7.54 g/t and 2.67 g/t), Green (1.79 g/t), and Trench Grid (4.8 g/t).

Overall, the 2021 program successfully delivered a comprehensive, property-scale GIS dataset and a regional understanding of mineral potential across First Mining’s Birch-Uchi tenure. The program provided a strong foundation for future exploration and highlighted the need to advance the project with additional datasets, including airborne geophysics, to further support regional target development.

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9.2.2 2022 Program

The 2022 program focused on regional screening at a district scale in conjunction with the completion of an airborne magnetic and electromagnetic survey (see Section 9.3). This campaign consisted of district-scale prospecting transects that included both rock grab sampling and soil sampling programs. The rock grab sampling programs in 2022 focused on favourable mineralization and alteration, returning significant gold results across multiple centres of mineralization over the project area and are highlighted by higher grade samples including 42.4 g/t Au at the Sirius target, 15.4 g/t Au at the Bronco target, and 15.3 g/t Au at the Canamer target. Prospecting by First Mining in the vicinity of the historic Sol D’Or Mine returned rock grab samples with anomalous gold, including one float sample at 34.7 g/t Au, further supporting the targeted areas’ gold endowment.

A soil survey was designed across the properties to screen for gold in soil anomalies along key structural trends and to infill gaps identified in historical soil data. Two targeted soil grids were established, one targeting the Grace Deformation Zone with 300 m line spacing and 100 m station spacing, and another over the Uchi Deformation Zone with 200 m line and station spacing. A total of 230 soil samples were collected, with the majority taken from the Swain and Swain Post claim areas.

A total of 308 rock grab samples were collected during the 2022 field campaign across the properties targeting favourable mineralization and alteration, and geological mapping was completed at 305 outcrops.

Overall, the 2022 program was successful in the advancement of regional screening that highlighted key structural trends and refined priority exploration targets for 2023 follow-up campaigns. In addition, the 2022 airborne geophysics survey supported the development of key targets for in-field follow-up and the advancement of future drilling programs.

9.2.3 2023 Program

In 2023, regional screening continued in the Birch Uchi, alongside focused rock grab and soil sampling programs which were conducted across areas of interest delineated during the 2022 field campaign. A total of 373 rock grab samples and 445 soil samples were collected, identifying targets to be followed up in future exploration programs for further evolvement. Bedrock mapping was also carried out at 824 outcrops. Highlights from this campaign included the discovery of the now-established Challenger target, located approximately 12 km southwest of Springpole, where two new mineralized occurrences were identified approximately 60 m apart, and where rock grab samples returned gold values including 25.60 g/t Au, 7.10 g/t Au and 4.42 g/t Au. The bedrock mapping campaign successfully demonstrated that the Challenger target has a total identified mineralized strike length of 120 m. A drilling program was also completed at the Saddle area in 2023 (Figure 9‑2), which is discussed further in Section 10.3.

Another highlight target advanced in 2023 was the Bronco target, which was further characterized and tested along its breadth through rock grab sampling returning 9.3 g/t Au. Follow-up detailed mapping and channel sampling are recommended at Bronco in preparation for drilling.

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Rock grab samples were collected where favourable mineralization and alteration were observed. Soil sampling varied from regional grids (>200 m line spacing, 200 m sample spacing) to target-scale and infill spacing (<150 m line spacing, 50 m sample spacing) following up on results from the 2022 soil program.

Also in 2023, a lithium prospecting program across the Allison Lake North area was conducted. A total of 22 outcrops were mapped, and a total of five soil samples, 18 rock grab samples, two float samples and five whole rock samples were collected. During this program, targeting was generally focused on available outcrop conducive to sampling. The program confirmed locations of historical showings as well as the presence of tourmaline-bearing pegmatites within First Mining’s mineral tenure. Based on historical and newly acquired data, the Allison Lake North Area shows potential for mineralization with elevated lithium (Li), caesium (Cs) and tantalum (Ta) as well as some rare earth element enrichment.

9.2.4 2024 Program

Mapping and sampling programs in 2024 were focused largely on the Saddle, Challenger and Bronco target areas, as well the Southeast Extension target at the Springpole project.

The 2024 summer field campaign at the Challenger target area focused on mineralization expansion and characterization of the discovery area (Figure 9‑1), and First Mining’s exploration teams completed 8 km of mapping transects, 40 channel samples, 59 rock grab samples, and 289 soil samples (50 m line spacing, 15 m sample spacing) across this underexplored area (Figure 9‑2). Rock grab sampling focused on collecting samples which contained favourable mineralization, alteration or structure orientation similar to Challenger and the results yielded four new gold occurrences, including visible gold in rock grab and channel samples, proximal to the original 2023 discovery area. Additional mineralized outcrops were identified 75 m to the southwest of the Challenger discovery target, returning up to 26.6 g/t Au.

A new area of gold mineralization, the Charger target, was identified approximately 300 m southwest of the Challenger discovery, between the Challenger and the Saddle targets (Figure 9‑1). Rock grab samples from Charger returned up to 20.3 g/t Au and 3.76 g/t Au. Initial data suggests that Charger represents a third interpreted mineralized trend, providing a promising focus for further exploration advancement.

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Figure 9‑1: Location Map of Main Targets Saddle and Challenger, including Charger Trend

Source: First Mining, 2024

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Detailed channel sampling at Challenger in 2024 focused on areas over high-grade rock grab samples while maintaining approximately 1.5 to 2 m spacing in hand-stripped outcrops. Assay highlights include 6.53 g/t Au over 3 m, including 57.8 g/t Au over 0.3 m; see Table 9‑1 for a list of channel sample results. Rock grab sample assay highlights included 26.6 g/t Au, 20.3 g/t Au, and 7.73 g/t Au. The Challenger mineralization appears to occur within a geophysical resistivity low signature therefore warrants follow-up exploration.

Table 9‑1 : Selected Channel Sample Assays from the Challenger Target

Channel ID		From (m)	To (m)	Length (m)	Grade (Au g/t)	Target Area

| SAT24C-001 | | 0.70 | 2.40 | 1.70 | 2.17 | Challenger |

| SAT24C-001 | inc. | 0.70 | 1.10 | 0.40 | 4.65 | |

| SAT24C-002 | | 0.30 | 2.00 | 1.70 | 2.57 | Challenger |

| SAT24C-002 | inc. | 0.30 | 1.30 | 1.00 | 3.81 | |

| SAT24C-003 | | 0.40 | 3.40 | 3.00 | 6.53 | Challenger |

| SAT24C-003 | inc. | 2.10 | 2.40 | 0.30 | 57.80 | |

| SAT24C-004 | | 0.55 | 1.80 | 1.25 | 2.01 | Challenger |

| SAT24C-005 | | 0.75 | 1.70 | 0.95 | 3.53 | Challenger |

| SAT24C-008 | | 0.75 | 1.35 | 0.60 | 6.08 | Challenger |

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Figure 9‑2: Rock Grab Sample Highlights from the 2023-2024 Programs in the Saddle-Challenger Target Areas

Source: First Mining, 2024

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A higher density rock grab sampling campaign was designed and completed at the Bronco target area focusing on systematically crossing known gold occurrences in an effort to further advance the delineation of favourable mineralization trends. Rock grab samples were collected at a sample spacing of 5 m and line spacing of 30 m. Channel samples were collected along strike over a length of 26 m near the 2022 and 2023 anomalous gold samples. The results from this campaign strengthened the geological characterization of the Bronco target, identifying favourable mineralization trends along key structural contacts and guiding future exploration.

During 2024, field mapping and sampling was also completed in the Springpole Project area, at the OMJ and Southeast Extension targets. Mapping and rock grab sampling at the OMJ target further advanced the geological characterization of the target by better delineating favourable host lithologies and the structural setting, supporting future exploration of the target. Geological mapping in the Southeast Extension target area was completed to further advance the geological interpretations and to support refinement of the stratigraphic and structural interpretations of the target to better inform exploration advancement.

In total, the 2024 field program included the collection of 606 rock grab samples and 289 soil samples, as well as 14 channels cut and 734 channel samples.

9.2.5 2025 Program

During the 2025 exploration program, 89 grab samples were collected over the First Mining mineral tenure, and six channels were cut with 67 channel samples collected. First Mining continues to integrate the data for regional consolidation, analysis and target evolvement.

Locations of regional targets and the grab and soil samples collected by First Mining in the 2021 to 2025 field programs are presented on Figure 9‑3 and Figure 9‑4, and grab sample highlights from the 2022 – 2024 programs are in Table 9‑2.

A targeted geological mapping campaign was conducted around the main Springpole project proposed infrastructure areas in the summer of 2025, to enhance and update the property-scale geological mapping originally completed in 2005. This work was aimed at improving the resolution and confidence of the geological characterization in the main project area. A total of 378 outcrops and approximately 30 km of planned mapping transects were completed in 2025 within the project footprint and surrounding area (Figure 9‑5), providing a more detailed foundation for ongoing geological assessment and future lithological model updates. This mapping work is planned to continue in 2026, expanding outward from the main project area to further refine regional geological understanding.

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Figure 9‑3: Rock Sample Locations from the 2021 to 2025 Exploration Programs

Source: First Mining, 2025

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Figure 9‑4: Soil Sample Locations from the 2021 to 2025 Exploration Programs

Source: First Mining, 2025

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Figure 9‑5: Bedrock Mapping Locations from the 2025 Exploration Program

Source: First Mining, 2025

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Table 9‑2: Grab Sampling Highlights from 2022 to 2024 Field Programs

Area	Sample Type	Sample ID	Au (g/t)	Program

| Swain | Grab | B1055032 | 2.93 | 2022 |

| Swain | Grab | B1055051 | 9.34 | 2022 |

| Wagner | Grab | B1055002 | 42.40 | 2022 |

| Wagner | Grab | C261738 | 2.22 | 2023 |

| Wagner | Grab | C261739 | 3.90 | 2023 |

| Wagner | Grab | C261740 | 2.36 | 2023 |

| Wagner | Grab | C261743 | 9.59 | 2023 |

| OMJ | Grab | C262558 | 2.41 | 2024 |

| OMJ | Grab | C262559 | 3.03 | 2024 |

| OMJ | Grab | C262564 | 4.02 | 2024 |

| OMJ | Grab | C262578 | 7.29 | 2024 |

| OMJ | Grab | C262593 | 49.00 | 2024 |

| Sol D'Or | Float | B1055352 | 34.70 | 2022 |

| Challenger | Grab | C261397 | 4.42 | 2023 |

| Challenger | Grab | C261398 | 25.60 | 2023 |

| Challenger | Grab | C261801 | 2.81 | 2023 |

| Challenger | Grab | C261803 | 2.90 | 2023 |

| Challenger | Grab | C261809 | 7.10 | 2023 |

| Challenger | Grab | C261817 | 3.76 | 2023 |

| Challenger | Grab | C262503 | 2.85 | 2024 |

| Challenger | Grab | C262506 | 26.60 | 2024 |

| Charger | Grab | C262511 | 20.30 | 2024 |

| Charger | Grab | C262513 | 3.76 | 2024 |

| Charger | Grab | C262637 | 7.73 | 2024 |

| Challenger | Grab | C262701 | 3.74 | 2024 |

| Saddle | Grab | C262741 | 3.23 | 2024 |

| Bronco | Grab | B1055294 | 2.67 | 2022 |

| Bronco | Grab | B1055423 | 15.40 | 2022 |

| Bronco | Grab | C261379 | 2.19 | 2023 |

| Horseshoe | Grab | C261698 | 3.59 | 2023 |

| Bronco | Grab | C261769 | 9.30 | 2023 |

| Bronco | Grab | C255912 | 3.65 | 2024 |

| Bronco | Grab | C262664 | 38.50 | 2024 |

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9.2.6 Methodologies

9.2.6.1 Soil Sampling

Soil samples for all First Mining programs were collected by steel spade shovel and steel trowel or aluminum soil auger, depending on the depth to the B-horizon soil profile. The B-horizon was the targeted soil horizon, with the C-horizon being the next favourable horizon, followed by the A-horizon. The soil samples were collected in paper kraft bags labeled with a unique sample ID; all samples were approximately 500 g or a ¾ full kraft bag. Each soil sample location was marked with the sample ID, sample date, and sampler name written on flagging tape and tied to a tree above or within 2 m of the sample site. Once dried, the soil samples were placed in cardboard boxes and taped, to be shipped to the assay lab for gold fire assay and multi-element analysis by ICP method with an aqua regia digest. QAQC samples were inserted into the sample stream with standard and blanks included at a 10% insertion rate, and field duplicates at a 3% insertion rate.

9.2.6.2 Rock Grab Sampling

Rock grab samples were collected in polyethylene bags and labeled with a unique sample ID from a sample booklet. Samples were selected based on favourable mineralization, favourable alteration or favourable structural setting for gold mineralization. Samples were collected with a hammer and chisel, and all samples were about two-fist sized. Samples were zip tied and placed in rice bags to be shipped to the assay lab for gold fire assay and multi-element analysis by ICP method with a four-acid digest. QAQC samples were inserted into the sample stream at a 15% insertion rate for standard and blanks, and a 5% insertion rate for field, crush and lab duplicates.

9.2.6.3 Channel Sampling

Areas designated for channel sampling were detail mapped and captured at a high resolution focusing on lithology, alteration, mineralization and structure before commencing with channel sampling. Channel samples were marked on the outcrop with a crayon and photographed. Sample intervals were measured with a tape measure ensuring they comply to sample length ranges (0.3 m - 1 m), and an average width of 0.05 m wide and 0.1 m deep. Once a sample length was cut, the sample was chipped out with a hammer and chisel and subsequent sampling and assaying protocols were the same as for the rock grab sampling programs (see Section 9.2.6.2). Dymo tags were stamped with the channel number, sample ID, and length and placed at the start of each sample cross-cut.

9.3 Airborne Survey

In 2022, First Mining entered into an agreement with SkyTEM Canada Inc. (“SkyTEM”) to carry out a high resolution airborne electromagnetic and magnetic survey across a large portion of its mineral tenure in the Birch-Uchi Greenstone Belt. The purpose of the survey was to collect time domain electromagnetic and magnetic data as part of a data gap analysis, providing First Mining with a modern, continuous geophysical dataset across its consolidated tenure over key structures and deformation zones in the belt, supporting exploration targeting on a district scale.

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The survey was flown over a total of 3,849-line km at a traverse interval spacing of 100 m. Mira Geosciences was contracted to perform a QAQC assessment on the collected dataset, and to further support advanced plate and EM conductor modelling on the SkyTEM dataset in key areas of interest, merge the SkyTEM & legacy magnetic data, and generate a first derivative (1D) inversion model on a targeted area. This work was completed and delivered to First Mining in March 2023, and included a series of EM targets for further review and analysis along key structures and zones of deformation.

The location of the airborne survey and total magnetic field imagery is presented in Figure 9‑6.

The SkyTEM survey provided a strong foundation for subsequent exploration activities, offering a key dataset to guide follow-up field campaigns and target prioritization in the coming years.

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Figure 9‑6: Airborne Geophysics over Gold Canyon Property Area showing 2022 Survey Area

Source: First Mining, 2025

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10 DRILLING

10.1 Introduction

This section details the drilling completed by both Gold Canyon between 2007 and 2013 and First Mining since 2016; earlier drilling has been summarized in Section 6 of the report. Any Gold Canyon drilling completed prior to First Mining’s acquisition of the Project is classed as historical work.

The majority of drill holes targeted the Portage zone, which is the most significant mineralized zone in the Springpole deposit, with many drilled from Springpole Lake. The Portage zone is largely comprised of disseminated mineralization striking 135° and extending from the surface to a depth of over 400 m, and is on average approximately 150 m in width and over 1,500 m in length. Mineralization in this zone is generally very continuous and also contains silver in close association with gold. Exploration drilling through the deposit has predominantly been completed along drill sections oriented southwest-northeast and spaced at approximately 50m. The dominant drill directions are approximately 220° azimuth, or vertical, with some holes oriented at azimuths between 18° and 40°. For angled holes, the dip is generally between -40 and -60.

Drilling to the north and northeast of the Portage zone intersects the Camp zone, which strikes approximately 120°, and the East Extension zone, which strikes approximately 135°. These two zones exhibit higher variability in gold and silver grades than the main Portage zone.

Locations of drilling discussed in this section are presented in Figure 10‑1, Figure 10‑2, Figure 10‑3 and Figure 10‑4, and a representative cross section through the Portage and East Extension zones looking northwest is presented in Figure 14‑3 in Section 14.

10.2 Historical Programs

10.2.1 Historical Drilling by Gold Canyon

Between 2007 and 2013, a total of 269 diamond drill holes were completed on the property by Gold Canyon, for a total of 90,305 m. Diamond drill programs were completed each year during this period, with the exception of 2009. A summary of the total holes and metres drilled in each program is presented in Table 10‑1, and drill procedures are discussed in Section 11.2.1 and Section 11.3.1. Further details on the work programs can be found in the report references listed in the table and summarized below.

Table 10‑1 : Gold Canyon Diamond Drilling Summary 2007 - 2013

Year	Total Holes	Total Metres	Report Reference

| 2007 | 11 | 2,122 | Smith, 2008a |

| 2008 | 7 | 2,452 | Smith, 2008b |

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Year	Total Holes	Total Metres	Report Reference

| 2010 | 29 | 10,437 | Roberts, 2010; Camier, 2012a |

| 2011 | 82 | 29,787 | Camier, 2012b, 2012c, 2012d, 2012e, 2012f |

| 2012 | 98 | 39,392 | Muntzert, 2013 |

| 2013 | 42 | 6,115 | Muntzert, 2014 |

Holes were primarily drilled for exploration purposes, with the following exceptions:

·	Four condemnation holes drilled in the Johnson Island area in 2012 (SC12-194, SC12-195, SC12-196 and SC12-198).
·	Eight holes (SG12-188 through SG12-193A) drilled for geotechnical purposes in 2012.
·	18 shallow holes (‘SV13’ series) between 15 m and 71 m length, targeting lake bottom sediments and till above bedrock surface and utilizing a ‘Vibracore’ drilling technique for better recovery (Muntzert, 2014).
·	Seven holes (SG13-200 to SG13-206) drilled in 2013 as part of the oriented core program, to obtain structural and geotechnical data around the proposed open-pit area.

Historical hole locations are presented on Figure 10‑1 and a summary of significant intercepts from these drill programs is provided in Table 10‑2.

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Figure 10‑1: Springpole Gold Project Historical Drill Hole Collar Location Map

Source: First Mining, 2025

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Table 10‑2: Drill Hole Intercepts, Gold Canyon 2007 – 2013

Year	Hole ID	From (m)	To (m)	Interval (m)	Au (g/t)	Ag (g/t)

| 2010 | SP10-001 | 12.5 | 64 | 51.5 | 0.93 | 1.1 |

| | SP10-002 | 242 | 335 | 93 | 2.4 | 11.2 |

| | SP10-004 | 31 | 182 | 151 | 0.72 | 2.6 |

| | SP10-006 | 278 | 363 | 85 | 0.93 | 6.7 |

| | SP10-007 | 33 | 250 | 217 | 1.57 | 7.5 |

| | SP10-008 | 257 | 451 | 194 | 1.22 | 7.6 |

| | SP10-009 | 3 | 167 | 164 | 1.02 | 2.7 |

| | SP10-011 | 229 | 323 | 94 | 2.51 | 10.7 |

| | SP10-012 | 275 | 408 | 133 | 0.79 | 8.6 |

| | SP10-016 | 206 | 511 | 305 | 1.03 | 4.7 |

| | SP10-019 | 182 | 489 | 307 | 1.44 | 5.5 |

| | SP10-024 | 166 | 391 | 225 | 1.48 | 4.7 |

| | SP10-026 | 54 | 407 | 353 | 1.17 | 3.9 |

| 2011 | SP11-030 | 14 | 73 | 59 | 2.54 | 2.02 |

| | SP11-033 | 13 | 315 | 302 | 1.38 | 7.17 |

| | SP11-034 | 37 | 110.5 | 73.5 | 1.18 | 6.18 |

| | | 162 | 331 | 169 | 1.08 | 6.29 |

| | SP11-035 | 37 | 68 | 31 | 1.02 | 3.63 |

| | | 105 | 200.5 | 95.5 | 1.22 | 3.26 |

| | SP11-036 | 204 | 394.5 | 190.5 | 0.9 | 3.97 |

| | SP11-037 | 54 | 316.5 | 262.5 | 0.92 | 4.67 |

| | SP11-039 | 60 | 117 | 57 | 0.41 | 3.11 |

| | | 132 | 165 | 33 | 0.53 | 4.48 |

| | SP11-040 | 51 | 151.5 | 100.5 | 7.59 | 8.82 |

| | SP11-041 | 161 | 237 | 76 | 1.5 | 5.6 |

| | SP11-042 | 9 | 411 | 402 | 0.76 | 2.86 |

| | SP11-043 | 42 | 153 | 111 | 2.03 | 7.01 |

| | SP11-044 | 132 | 351 | 219 | 0.71 | 11.81 |

| | SP11-045 | 36 | 90 | 54 | 2.15 | 19.13 |

| | SP11-045A | 63 | 213 | 150 | 2.59 | 12.4 |

| | SP11-046 | 34 | 63 | 29 | 0.57 | 5.49 |

| | | 238 | 306.5 | 68.5 | 0.82 | 6.74 |

| | SP11-047 | 22.7 | 177 | 154.3 | 0.99 | 8.69 |

| | SP11-048 | 121 | 315 | 194 | 1.11 | 13.79 |

| | SP11-049 | 20 | 152 | 132 | 1.37 | 7.6 |

| | SP11-050 | 139 | 247 | 108 | 0.54 | 3.3 |

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Year	Hole ID	From (m)	To (m)	Interval (m)	Au (g/t)	Ag (g/t)

| | | 304 | 328 | 24 | 0.63 | 3.96 |

| | SP11-051 | 14 | 164 | 150 | 1.14 | 3.8 |

| | SP11-052 | 19 | 158 | 139 | 1.04 | 10.83 |

| | SP11-053 | 11.4 | 21 | 9.6 | 2.95 | 13.32 |

| | SP11-054 | 23 | 165 | 142 | 0.81 | 17.63 |

| | SP11-056 | 55.5 | 228 | 172.5 | 0.93 | 21.38 |

| | SP11-057 | 91.5 | 312 | 220.5 | 0.84 | 4.91 |

| | SP11-058 | 48.4 | 159 | 110.6 | 2.48 | 4.57 |

| | SP11-059 | 72 | 364.5 | 292.5 | 1.13 | 4.15 |

| | SP11-060 | 51 | 255 | 204 | 1.15 | 4.89 |

| | SP11-066 | 16 | 40 | 24 | 17.48 | 3.2 |

| | SP11-067 | 15 | 54 | 39 | 2.93 | 1.04 |

| | SP11-070 | 93 | 401 | 308 | 1.29 | 1.4 |

| | SP11-071 | 149 | 435 | 286 | 1.06 | 7.73 |

| | SP11-072 | 63 | 382 | 319 | 0.97 | 2.52 |

| | SP11-073 | 17 | 267 | 250 | 1.46 | 3 |

| | SP11-074 | 121 | 490 | 369 | 0.91 | 5.58 |

| | SP11-075 | 113 | 319 | 206 | 0.91 | 2.85 |

| | SP11-076 | 28 | 149 | 121 | 0.71 | 1.53 |

| | | 295 | 387 | 92 | 0.6 | 2.15 |

| | SP11-077 | 10 | 87 | 77 | 0.73 | 0.51 |

| | | 130 | 236 | 106 | 3.35 | 2.15 |

| | SP11-078 | 249 | 363 | 114 | 0.58 | 4.09 |

| | SP11-079 | 3 | 177.5 | 174.5 | 0.56 | 1.97 |

| | | 312 | 416 | 104 | 0.59 | 2.13 |

| | SP11-080 | 48 | 124 | 76 | 0.62 | 1.9 |

| | SP11-081 | 92 | 321 | 229 | 0.82 | 2.42 |

| | SP11-082 | 85 | 171 | 86 | 1.07 | 17.95 |

| | | 262 | 403 | 141 | 0.72 | 5.93 |

| | SP11-083 | 24 | 155 | 131 | 0.77 | 3.12 |

| | SP11-084 | 15 | 349.5 | 334.5 | 0.81 | 5.14 |

| | SP11-087 | 159 | 353 | 194 | 0.96 | 5.97 |

| | SP11-088 | 7 | 36 | 29 | 0.62 | 1.2 |

| | | 300 | 346 | 46 | 0.58 | 6.69 |

| | | 364 | 441 | 77 | 0.72 | 4.74 |

| | SP11-091 | 66 | 376 | 310 | 1.89 | 6.6 |

| | SP11-092 | 109 | 177 | 68 | 0.58 | 1 |

| | SP11-093 | 122 | 316.5 | 194.5 | 0.85 | 3.73 |

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Year	Hole ID	From (m)	To (m)	Interval (m)	Au (g/t)	Ag (g/t)

| | SP11-094 | 312.5 | 455 | 142.5 | 0.71 | 4.7 |

| | SP11-096 | 66 | 323 | 257 | 1.48 | 5.82 |

| | SP11-097 | 27 | 60 | 33 | 0.71 | 0.78 |

| | | 200 | 291 | 91 | 0.79 | 4.67 |

| | SP11-098 | 3 | 124 | 121 | 1.68 | 3.67 |

| | | 311.5 | 401.5 | 90 | 2 | 7.17 |

| | SP11-099 | 254 | 430 | 176 | 0.8 | 7.63 |

| | SP11-100 | 404.5 | 482 | 77.5 | 0.62 | 5.38 |

| | SP11-104 | 279 | 427 | 148 | 1.66 | 6.11 |

| | SP11-106 | 344.5 | 472 | 127.5 | 3.52 | 10.72 |

| | SP11-107 | 247 | 377 | 130 | 0.73 | 2.4 |

| | SP11-108 | 381 | 446 | 65 | 3.14 | 11.02 |

| | SP11-109 | 509.5 | 540 | 30.5 | 1.09 | 4.52 |

| 2012 | SP12-127 | 251 | 398 | 147 | 1.14 | 5.47 |

| | SP12-128 | 230 | 549 | 319 | 1.02 | 4.81 |

| | SP12-131 | 301.3 | 546 | 244.7 | 0.8 | 5.84 |

| | SP12-146 | 77 | 91 | 14 | 5.03 | 99.69 |

| | SP12-158 | 16.7 | 60.2 | 43.5 | 1.81 | 3.23 |

| | SP12-160 | 23 | 384 | 361 | 1.08 | 8.83 |

| | SP12-163 | 130.9 | 265 | 134.1 | 0.91 | 13.57 |

| | SP12-181 | 157 | 225 | 68 | 0.72 | 8.74 |

| | SP12-183 | 202 | 385 | 183 | 0.61 | 3.47 |

| | SP12-186 | 114 | 240 | 126 | 1.17 | 2.63 |

| 2013 | SG13-201 | 148 | 196 | 48 | 0.66 | 4.76 |

| | SG13-205 | 226 | 327 | 101 | 0.93 | 12.36 |

| | SG13-206 | 108 | 175 | 67 | 1.94 | 2.94 |

| | including | 145 | 153 | 8 | 4.44 | 6.35 |

10.2.2 2013 Geotechnical and Structural Program

In 2013, SRK were contracted by Gold Canyon to complete a geotechnical and structural/geological program. This program was conducted at two separate times of the year, but data collected from each part was combined into one final report by C. Nagy (SRK, 2013).

During the winter drilling program, seven holes were drilled using HQ3 equipment (triple tube) and Reflex ACT 2 orientation tools. Further details of the drilling procedures for these holes can be found in Muntzert (2014). SRK geotechnical engineers were on site to train, assist and supervise the drilling crews and Gold Canyon geologists. Details of the seven holes are provided in Table 10‑3.

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Further details on the structural modelling work undertaken by SRK can be found in the internal report by C. Nagy (SRK, 2013).

Table 10‑3: 2013 Oriented-Core Drilling Program

Hole ID	Easting (UTM)	Northing (UTM)	Elevation (m)	Length (m)	Azimuth (°)	Dip (°)

| SG13-200 | 549,649 | 5,693,248 | 391 | 351 | 200 | -70 |

| SG13-201 | 549,801 | 5,693,265 | 391 | 339 | 135 | -70 |

| SG13-202 | 549,500 | 5,693,423 | 391 | 351 | 270 | -70 |

| SG13-203 | 549,300 | 5,693,676 | 391 | 349.5 | 225 | -65 |

| SG13-204 | 549,521 | 5,693,730 | 391 | 351 | 019 | -65 |

| SG13-205 | 549,782 | 5,693,447 | 391 | 349.5 | 030 | -70 |

| SG13-206 | 549,140 | 5,693,889 | 391 | 310.5 | 040 | -50 |

Note:

1.	Coordinates in Universal Transverse Mercator (UTM) and World Geodetic System 1984 (WGS84).

| 2. | Elevations from LiDAR. |

10.3 First Mining Drilling, Springpole Project

10.3.1 Introduction

First Mining completed drilling between 2016 and 2024. Details of the drill procedures for the exploration, metallurgical, geotechnical and hydrological drill programs are provided in Section 11.2.2 and Section 11.3.2.

10.3.2 2016 Drill Program

The 2016 drill program was implemented by First Mining to collect additional material from the Portage zone so that additional metallurgical testing could be carried out. In total, 1,712 m were drilled in four holes (PM-DH-01 to PM-DH-04).

The 2016 drill hole locations are illustrated on Figure 10‑2 and significant drill intersections are summarized in Table 10‑4.

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Figure 10‑2: Location of first Mining 2016 to 2020 Drill Holes

Source: First Mining, 2021

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Table 10‑4: Significant Intercepts from 2016 Metallurgical Drilling Program

Hole	From (m)	To (m)	Length (m)	Au (g/t)	Ag (g/t)

| PM-DH-01 | 16.5 | 60 | 43.5 | 1.57 | 11.59 |

| and | 84 | 371 | 287 | 1.25 | 5.48 |

| including | 107 | 118 | 11 | 4.95 | 5.68 |

| and including | 201 | 236 | 35 | 2.79 | 13.8 |

| PM-DH-02 | 74 | 332 | 258 | 1.87 | 8.37 |

| including | 171 | 175.5 | 4.5 | 17.97 | 47.57 |

| and including | 193 | 211 | 18 | 3.22 | 14.42 |

| and including | 233 | 237 | 4 | 14.66 | 53.15 |

| and | 401 | 406 | 5 | 3.15 | 21.64 |

| PM-DH-03 | 50 | 294 | 244 | 1.2 | 6.96 |

| and | 312 | 357 | 45 | 2.7 | 10.42 |

| including | 332 | 338 | 6 | 10 | 26.98 |

| and | 371 | 409 | 38 | 0.65 | 4.15 |

| PM-DH-04 | 3.95 | 11 | 7.05 | 2.21 | 0.58 |

| and | 16 | 18 | 2 | 15.66 | 5.1 |

| including | 16 | 17 | 1 | 30.36 | 9.94 |

| and | 22 | 23 | 1 | 4.95 | 1.32 |

| and | 88 | 90 | 2 | 3.91 | 0.35 |

| and | 103 | 109 | 6 | 3.39 | 0.93 |

| including | 104 | 105 | 1 | 17.96 | 3.81 |

| and | 134 | 141 | 7 | 0.59 | 0.73 |

| and | 146.43 | 154.88 | 8.45 | 0.82 | 0.8 |

| and | 186.32 | 333 | 146.68 | 2.15 | 4.88 |

| including | 199.15 | 205 | 5.85 | 11.22 | 4.79 |

| and including | 203 | 204 | 1 | 30.97 | 13.50 |

| and including | 245 | 265 | 20 | 3.9 | 4.53 |

| and | 274 | 285 | 11 | 3.21 | 9.31 |

Note: Reported widths are drilled core lengths; assay values are uncut

10.3.3 2018 Drill Program

In 2018, First Mining carried out a short geotechnical drill program to test the integrity of ground relevant to potential dike construction and characterize the dike foundation materials. Eleven short holes were drilled totalling 243 m. The drill hole locations are illustrated on Figure 10‑2.

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10.3.4 2020 Drill Program

The 2020 Springpole drill program consisted of a combination of metallurgical, geotechnical/hydrogeological and condemnation drilling. A total of 48 drill holes were completed during this program over 7,892 m.

10.3.4.1 Metallurgical Drill Program

The 2020 metallurgical drilling program was implemented to collect additional material for metallurgical testwork within the proposed open pit. Approximately 1,182 m of drilling was completed in three drill holes (SM20-001 to SM20-003). Drill hole locations are illustrated on Figure 10‑2 and significant drill intersections are summarized in Table 10‑5.

Table 10‑5: Significant Intercepts from the 2020 Metallurgical Drilling Program

Hole ID	From (m)	To (m)	Length (m)	Au (g/t)	Ag (g/t)

| SM20-001 | 5 | 68 | 63 | 1 | 4.28 |

| including | 20 | 30 | 10 | 2.48 | 10.3 |

| and | 97 | 153 | 56 | 4.88 | 9.98 |

| including | 135 | 137 | 2 | 122.79 | 103.11 |

| and | 221 | 253 | 32 | 0.52 | 3.99 |

| and | 279 | 286 | 7 | 0.72 | 6.48 |

| and | 293 | 325 | 32 | 1.31 | 5.19 |

| and | 334 | 347 | 13 | 0.47 | 1.39 |

| SM20-002 | 129 | 136 | 7 | 0.67 | 2.67 |

| and | 228 | 307.5 | 79.5 | 0.67 | 3.78 |

| and | 331 | 334.5 | 3.5 | 1.35 | 7.96 |

| and | 336 | 389 | 53 | 1.55 | 10.09 |

| including | 372 | 374 | 2 | 7.94 | 31.01 |

| SM20-003 | 130 | 157 | 27 | 0.66 | 10.17 |

| Including | 130 | 136 | 6 | 1.71 | 32.66 |

| and | 186 | 432 | 246 | 1.53 | 6.49 |

| including | 214 | 244 | 30 | 3.62 | 7.71 |

| and including | 277 | 315 | 38 | 2.37 | 13.92 |

| and including | 343 | 350 | 7 | 3.2 | 8.77 |

Note: Reported widths are drilled core lengths; assay values are uncut

10.3.4.2 Condemnation Drill program

Condemnation drilling was undertaken in 2020 in areas of proposed mine infrastructure. All drill holes intersected basalt/andesite with minor tuff horizons, and only minor sulphides in very narrow horizons (pyrite, some chalcopyrite) lacking favourable alteration. Approximately 2,218 m of drilling was completed in 20 drill holes. Sampling was limited to zones of visible mineralization and in total only 47.45 m were sampled. Results varied from <0.005 ppm Au (less than detection limit) to a maximum of 0.21 ppm Au.

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The 2020 condemnation drill hole locations are shown on Figure 10‑2.

10.3.4.3 Geotechnical and Hydrogeological Drill Program

In 2020, Fracflow Consultants Inc. were retained by First Mining to complete a geotechnical-hydrogeological investigation on the pit wall area and the areas of proposed mine infrastructure.

The 2020 geotechnical drill program comprised 48 diamond drill holes over 7,892 m. Ten of these holes (SGH20-001 to SGH20-010) targeted the pit wall area and consisted of drilling and logging of inclined HQ size boreholes, packer tests, fracture surveys using acoustic televiewer, rock testing (point load tests and Brazilian tests), and multi-level piezometer installation. In addition, a detailed geotechnical field testing and sampling program was completed over the areas of proposed mine infrastructure, which included test pit excavations (for overburden investigation), hand auguring, NQ-size borehole drilling, and ground penetrating radar (GPR) surveys in selected locations.

One hydrogeological drill hole (SPW20-001) was drilled to characterize the aquifer on the north side of the current camp area.

None of the holes from the geotechnical program were assayed, with the exception of 3 holes which intersect the proposed pit (SGH20-006 and SGH20-009, and a portion of SGH20-001) which were assayed at later dates in 2021 and 2023.

Locations of the geotechnical holes are illustrated on Figure 10‑2.

10.3.5 2021 Drill Program

The 2021 Springpole drill program consisted of drilling for metallurgical, exploration, geotechnical, hydrogeological, ARD and condemnation purposes (Table 10‑6). A total of 55 drill holes were completed, totaling 8,748 m. Figure 10‑3 shows the location of the 2021 drill hole locations.

Table 10‑6: Summary of 2021 Drill Program

Purpose	No. of Holes	Metres

| ARD | 3 | 231 |

| Condemnation | 5 | 1,030 |

| Exploration | 13 | 4,366 |

| Metallurgical | 10 | 2,632 |

| Geotechnical | 23 | 449 |

| Hydrogeological | 1 | 40 |

| Total | 55 | 8,748 |

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10.3.5.1 Metallurgical Drill Program

The 2021 metallurgical drilling program was implemented to collect additional material within the proposed open pit for ongoing metallurgical testwork to support the Feasibility Study on the Project. Approximately 2,632 m of drilling was completed in ten drill holes (SM21-001 to SM21-010). Samples from the ten drill holes, all located within the 2021 PFS pit shell, were selected to represent the major lithologies and mineralized zones in the target mining area and were provided to the metallurgical lab for the additional testwork, which comprised flowsheet optimization, variability testing, additional flotation studies and materials handling testing.

The 2021 metallurgical drill hole locations are illustrated on Figure 10‑3 and select significant mineralized intersections are summarized in Table 10‑7. Half core samples were submitted for fire assay and ICP assaying, with the exception of selected whole core samples which were taken for comminution testwork and waste sampling.

Table 10‑7: Significant Intercepts, 2021 Metallurgical Drilling Program

Hole ID	From (m)	To (m)	Length (m)	Au (g/t)	Ag (g/t)

| SM21-001 | 17 | 33 | 16 | 2.68 | 2.25 |

| including | 17 | 25 | 8 | 4.88 | 3.91 |

| and including | 18 | 19 | 1 | 17.2 | 3.27 |

| SM21-002 | 26 | 29 | 3 | 19.03 | 39.3 |

| including | 27 | 28 | 1 | 49.5 | 78.2 |

| SM21-003 | 208 | 221 | 13 | 0.59 | 4.02 |

| and | 231.5 | 285.4 | 53.9 | 0.68 | 5.83 |

| and | 310.5 | 318 | 7.5 | 19.57 | 50.17 |

| including | 312 | 314 | 2 | 60.55 | 86.85 |

| and | 331.5 | 337.5 | 6 | 2.56 | 21.07 |

| including | 334.5 | 337.5 | 3 | 4.12 | 34.63 |

| SM21-004 | 252 | 313.5 | 61.5 | 0.79 | 5.4 |

| and | 315 | 324 | 9 | 2.4 | 32.87 |

| SM21-005 | 22.4 | 45.1 | 22.7 | 2.71 | 0.85 |

| including | 24 | 34.7 | 10.7 | 4.74 | 1.15 |

| and | 55.1 | 58 | 2.9 | 8.2 | 1.52 |

| including | 56 | 57 | 1 | 15.9 | 2.5 |

| and | 191 | 283 | 92 | 1.42 | 2.28 |

| including | 223 | 227 | 4 | 5.15 | 3.55 |

| and including | 232 | 240 | 8 | 3.24 | 2.31 |

| and including | 261 | 267 | 6 | 3.79 | 4.22 |

| and | 311 | 323 | 12 | 0.52 | 2.17 |

| SM21-006 | 32 | 42 | 10 | 3.59 | 73.29 |

| including | 41 | 42 | 1 | 16.1 | 478 |

| and | 65.8 | 151 | 85.2 | 0.99 | 14.92 |

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Hole ID	From (m)	To (m)	Length (m)	Au (g/t)	Ag (g/t)

| and including | 77 | 89 | 12 | 2.87 | 54.26 |

| and | 158 | 176 | 18 | 0.5 | 1.14 |

| and | 198 | 219 | 21 | 0.4 | 1.77 |

| and | 223 | 294 | 71 | 1.24 | 7.05 |

| SM21-007 | 140 | 151 | 11 | 0.65 | 2.12 |

| and | 154 | 171 | 17 | 0.54 | 4.59 |

| and | 185 | 193 | 8 | 0.69 | 11.38 |

| and | 196 | 238 | 42 | 0.66 | 9.53 |

| and | 246 | 285 | 39 | 0.77 | 21.79 |

| including | 265 | 274 | 9 | 2.03 | 57.5 |

| SM21-008 | 37.5 | 47 | 9.5 | 2.29 | 4.08 |

| and | 57.5 | 173.3 | 115.8 | 1.67 | 4.4 |

| including | 92 | 95 | 3 | 4.3 | 9 |

| and including | 98 | 102 | 4 | 5.21 | 1.23 |

| and including | 125 | 126 | 1 | 8.17 | 2.4 |

| and | 187.3 | 199.32 | 12.02 | 1.2 | 3.19 |

| and | 230 | 244 | 14 | 0.63 | 1.71 |

| and | 260 | 282 | 22 | 0.41 | 2.02 |

| SM21-009 | 28.5 | 95 | 66.5 | 1.21 | 24.52 |

| including | 37 | 41 | 4 | 3.56 | 20.38 |

| and | 105 | 120 | 15 | 1.32 | 7.95 |

| and | 164.3 | 214.5 | 50.2 | 0.98 | 9.48 |

| SM21-010 | 11.4 | 64 | 52.6 | 1.35 | 9.99 |

| including | 40 | 41 | 1 | 9.45 | 128 |

| and | 71 | 88 | 17 | 0.35 | 1.55 |

| and | 114 | 281 | 167 | 2.24 | 9.33 |

| including | 114 | 124 | 10 | 5.93 | 7.26 |

| and including | 225 | 238 | 13 | 6.58 | 39.42 |

| and | 288.5 | 308 | 19.5 | 0.44 | 4.92 |

| and | 309.5 | 372.5 | 63 | 0.64 | 4.8 |

Note: Reported widths are drilled core lengths; assay values are uncut

10.3.5.2 Exploration Drill Program

10.3.5.2.1 Eastern Trachyte Targets

This exploration drilling was undertaken approximately 1 km east of the 2021 PFS pit location in an area of outcropping trachyte along a structural trend subparallel with the dominant structural trend associated with mineralization at the Springpole Project. First Mining geologists observed the presence of pyrite in trachyte on surface, and the 2004 Fugro geophysical survey indicates an associated potassium signature and locally anomalous magnetic response in this area. A series of six holes were drilled from three pads to test the trachyte targets. The eastern-most holes (SP21-005 and SP21-006) encountered a series of interlaminated andesite and tuff units consistent with the local basement suite, prior to encountering the favourable host units of trachyte, trachyte breccia, and heterolithic breccia, which the other four SP21 holes (SP21-001 to SP21-004) encountered from the beginning of the hole. All the exploration holes terminated in andesitic basement rock. Veining consisted of regionally consistent mixed quartz/carbonate in the basement units, with multiple stages of quartz veining in both the basement suite and the units of interest. Moderate to intense potassic and silicious alteration was encountered through the favourable units in all of the 2021 exploration holes, with variably strong sulphide mineralization (disseminated pyrite ranging from 2-10%, with locally up to 25% dominantly around iron formation clasts within the heterolithic breccias) and local presence of base metals such as molybdenum, galena and sphalerite as well as fracture-controlled fluorite.

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Approximately 1,545 m of drilling was completed in 6 drill holes. In total, 1,497 m of core was sampled and there were no significant intercepts of gold or silver mineralization; assay results ranged from <0.005 ppm Au (less than detection limit) to a maximum of 0.095 ppm Au.

The 2021 exploration drill hole locations from the eastern trachyte targets are illustrated on Figure 10‑3.

10.3.5.2.2 Southwest Pit Area

In 2021, seven diamond drill holes totaling 2,820.5 m (SG21-001 to SG21-007) were completed to investigate the south-western side of the proposed pit. The initial two holes were planned as geotechnical holes targeting the pit wall area, so were drilled at HQ size. However, gold and silver mineralization was encountered in these holes, so an additional five angled exploration holes were drilled at NQ size to further delineate this area of new mineralization on the edge of the proposed pit. These holes were treated as exploration holes and First Mining completed the logging and sampling. No detailed geotechnical logging or testwork was completed on any of these holes. One of the seven drill holes was also utilized to collect hydrogeological data (SGH21-006).

Hole SG21-007 did not intersect favourable geology and there was no visible mineralization, so this hole was not sampled during the 2021 program. The other six drill holes did intersect favourable geology and mineralization. The assays results include wide intersections of gold mineralization, for example 69 m of 0.66 g/t Au (in SG21-002 from 299 m to 368 m), as well as high grade individual samples (e.g. 19.7 g/t Au from 141-142 m in SGH21-006). Locations of the drill holes are shown on Figure 10‑3 and a summary of significant assay intercepts is provided in Table 10‑8.

Table 10‑8: Significant Intercepts from 2021 Exploration Drilling (SW Pit Area)

Hole ID	From (m)	To (m)	Length (m)	Au (g/t)	Ag (g/t)

| SG21-001 | 214 | 238 | 24 | 0.45 | 4.55 |

| SG21-002 | 247 | 259 | 12 | 0.92 | 5.79 |

| including | 258 | 259 | 1 | 4.98 | 35.9 |

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Hole ID	From (m)	To (m)	Length (m)	Au (g/t)	Ag (g/t)

| and | 299 | 368 | 69 | 0.66 | 6.37 |

| including | 330 | 331 | 1 | 4.11 | 23.4 |

| and including | 344 | 345 | 1 | 3.39 | 28.3 |

| and | 377 | 384 | 7 | 1.25 | 8.70 |

| SG21-003 | 202 | 219 | 17 | 0.40 | 3.26 |

| and | 293 | 326 | 33 | 0.56 | 3.59 |

| including | 298 | 306 | 8 | 1.1 | 6.63 |

| and including | 337 | 345 | 8 | 0.73 | 5.50 |

| SG21-004 | 316 | 334 | 18 | 0.41 | 6.49 |

| and | 341 | 367 | 26 | 0.52 | 7.30 |

| including | 359 | 360 | 1 | 3.30 | 40.30 |

| SG21-005 | 248 | 258 | 10 | 0.38 | 3.98 |

| SGH21-006 | 141 | 144 | 3 | 7.02 | 82.6 |

| including | 141 | 142 | 1 | 19.7 | 233 |

| and | 185 | 201 | 16 | 0.40 | 5.51 |

| and | 246 | 254 | 8 | 0.51 | 3.71 |

| and | 264 | 286 | 22 | 0.39 | 2.64 |

| and | 305 | 313 | 8 | 1.08 | 9.33 |

| including | 307 | 308 | 1 | 2.76 | 17.1 |

| and | 336 | 357 | 21 | 0.82 | 8.99 |

| including | 338 | 343 | 5 | 2.12 | 20.8 |

| and including | 339 | 340 | 1 | 6.00 | 70.30 |

Note: Reported widths are drilled core lengths; assay values are uncut

10.3.5.3 Geotechnical Drill Program

In 2021, Ausenco Engineering Canada Inc. (Ausenco) was retained by First Mining to complete a geotechnical-hydrogeological investigation in support of the Feasibility Study for the Project by collecting data to support the design of the following mine components:

·	Tailings Management Facility
·	Fish habitat development area.
·	Camp site.
·	Ore stockpiles.
·	Other facilities and infrastructure at the process plant area.

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Rodren Drilling Services Ltd. conducted this drilling program using HQ Triple Tube coring with a diamond drill rig running 12 h/day, under the supervision of First Mining and Ausenco. Drilling started in June 2021 and finished in November 2021 with a small break during the summer due to wildfire in close proximity to the camp. A total of 23 diamond HQ geotechnical drill holes between 5 and 41 m deep were completed during the 2021 field season, for a total metreage of approximately 449 m. Selected drill holes were equipped with geotechnical-hydrogeological monitoring equipment (standpipe and vibrating wire piezometers).

For each geotechnical drill hole, as well as detailed logging, Ausenco also carried out Standard Penetration Tests (where possible), in-situ packer tests to collect hydrogeological data, and a geotechnical rock record to determine the rock mass rating (RMR), and quality and strength of bedrock. Representative core samples were also selected for laboratory testing programs.

Some of the core material collected in this geotechnical drilling program was also used to augment the ongoing metal leaching and acid rock drainage (ML/ARD) testwork on the Project.

First Mining sampled approximately 58 m of this drill core for geochemical analysis by fire assay and ICP from drill holes BH-Q-18 and BH-Q-19, both located in the proposed fish habitat development area. Assays did not return any significant gold or silver values for these two holes. No significant mineralization or favourable geology was observed in the other geotechnical holes, and subsequently they were not assayed as part of this program.

The 2021 geotechnical drill hole locations are shown on Figure 10‑3.

10.3.5.4 Condemnation Drill Program

Condemnation drilling was continued by First Mining in 2021, in areas of proposed mine infrastructure. Approximately 1,030 m of drilling was completed in 5 drill holes. All drill holes intersected basalt/andesite with minor tuff horizons, and only minor sulphides in very narrow horizons (pyrite, some pyrrhotite) lacking favourable alteration.

Sampling was limited to one drill hole, SC21-051, the only drill hole containing visible sulphide mineralization above background. In total, 169 m were sampled. Results varied from <0.005 ppm Au (less than detection limit) to a maximum of 0.059 ppm Au.

Two of the five condemnation drill holes, SCH21-049 and SCH21-052, were also utilized to collect hydrogeological data.

The 2021 condemnation drill hole locations are shown on Figure 10‑3.

10.3.5.5 Acid Rock Drainage (ARD) Drill Program

In 2021, three diamond drill holes totaling 231 m (SP-ARD-001 to SP-ARD-003) were completed in order to obtain material for the ongoing ARD testwork program. Drill hole locations are shown on Figure 10‑3. These holes targeted ARD data gaps in the East Extension and Camp zones of the deposit, where rock could potentially be used for construction and the additional data could help address uncertainties related to metal leaching. Additional ARD samples were also taken from the geotechnical holes in the proposed fish habitat development area, as discussed in Section 10.3.4.3.

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Samples from the ARD drill holes were assayed for gold and multi-element (ICP) including silver. Select assays are presented in Table 10‑9.

Table 10‑9: Intercepts from 2021 ARD Drill Program

Hole ID	From (m)	To (m)	Length (m)	Au (g/t)	Ag (g/t)

| SP-ARD-002 | 29 | 29.7 | 0.7 | 0.32 | 0.50 |

| and | 33 | 34.2 | 1.2 | 1.18 | 2.3 |

| and | 36 | 37 | 1 | 0.55 | 0.40 |

| and | 39 | 40 | 1 | 2.11 | 0.90 |

| and | 65 | 78 | 13 | 3.59 | 1.68 |

| Including | 65 | 66 | 1 | 35.7 | 3.1 |

| SP-ARD-003 | 15 | 16 | 1 | 0.34 | 1.20 |

Note: Reported widths are drilled core lengths; assay values are uncut

10.3.5.6 Long Term Monitoring Well

In 2021, an additional vertical hole (SPW20-002) was drilled as a potential long-term monitoring well site. This hole was drilled outside of the area affected by the mine plan to provide background water data for the site in the future, if required. This hole is cased and capped ready for monitoring well installation. The drill hole was drilled to a depth of 40 m and was not sampled. The drill hole location is shown on Figure 10‑3.

10.3.6 2022 Drill Program

The 2022 Springpole drill program consisted of drilling for geotechnical and hydrogeological site investigation purposes, as well as for sample collection for the metal leaching and acid rock drainage (ARD) testwork program. A total of 63 holes were completed totalling 5,839 m (Table 10‑10). Figure 10‑3 shows the location of the 2022 drill hole locations.

Table 10‑10: Summary of 2022 Drill Program

Purpose	No. of Holes	Metres

| ARD | 10 | 1,671.2 |

| Geotechnical | 34 | 3,756.3 |

| Hydrogeological | 19 | 411.2 |

| Total | 63 | 5,838.7 |

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Figure 10‑3: Location of 2021 and 2022 Drill Holes

Source: First Mining, 2025

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10.3.6.1 Geotechnical Drill Program

A total of 34 diamond drill holes were completed during 2022 for geotechnical site investigation purposes, for a total metreage of approximately 3,756 m. Rodren Drilling Services Ltd. conducted these drilling programs using HQ Triple Tube diamond drill rigs under the supervision of First Mining, as well as Knight Piésold Ltd., WSP Canada Ltd. (formerly Wood plc), and SRK.

10.3.6.1.1 Dikes and TMF Investigation

Knight Piésold Ltd. and WSP Canada Ltd. were retained by First Mining to conduct a geotechnical site investigation to support the PFS level design of both the proposed dikes and tailings management facility (TMF).

For the site investigation on the dikes, ten drill holes totalling 527.43 m were completed, with Knight Piésold personnel on site monitoring the field program. The dike area drill holes were predominantly shallow (between 13.5 m and 41 m depth), with the exception of two deeper holes of 101 m length (SG22-020) and 201 m length (SG22-012). The drill holes were positioned along the footprints of the two proposed dikes, with the exception of hole SG22-024 which was positioned midway between the eastern dike and the southeast wall of the proposed pit. Drill hole SG22-020 was positioned approximately 400 m to the south of the western dike. This hole intersected a 50 m thick clay rich overburden layer above alternating andesite and highly altered trachyte units. Low grade gold mineralization is present throughout this hole, with the highest grades occurring within an intersection of 6 m averaging 0.53 g/t Au at 65 m – 71 m depth. The deepest drill hole in this program, SG22-012, intersected andesite down to 125 m, before transitioning into alternating tuffs and potassic-altered porphyry units. Gold mineralization is present throughout this hole, with a notable intersection of 7 m at 0.73 g/t Au from 165 to 172 m depth, including 2 m at 1.73 g/t from 170 to 172 m. Four holes in this program were not assayed (SG22-011, SG22-015, SG22-022 and SG22-024), and the remaining holes (SG22-009, SG22-010, SG22-013 and SG22-014) intersected andesites or andesite porphyries, with low grade gold mineralization (<0.1 g/t Au) present throughout.

A further 17 drill holes totalling 199.73 m were completed to investigate the overburden and bedrock properties within the proposed TMF area. Of these, ten holes were drilled in spring 2022 under the supervision of Knight Piésold, and the remaining seven holes were completed in the fall under the supervision of Wood plc (now WSP Canada Ltd). Drill holes in this program were vertical, with shallow depths ranging from 5 m to a maximum of 21 m. All drill holes intersected andesites and basalts, and of the nine holes in this program that were assayed, none encountered any gold mineralization.

Geotechnical work completed as part of the dike and TMF site investigation program included the following:

·	Diamond drilling of overburden and bedrock.
·	Geotechnical logging of all recovered material, including overburden and bedrock.
·	Collection of representative soil samples for laboratory testing.
·	In situ geotechnical testing of soil encountered.
·	Packer tests conducted in the bedrock.

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10.3.6.1.2 Open Pit Investigation

SRK was contracted by First Mining to carry out a geotechnical site investigation in support of the pit slope design at Feasibility Study level. The field program was designed to collect detailed geotechnical and hydrogeological data from drill holes, and targeted areas of low-RQD material identified in the central and southern parts of the deposit and the proposed open pit. Seven angled drill holes totalling 3,029 m were completed between February and April 2022.

Drilling in this southwest area of the deposit identified additional mineralization (the “SW Extension Zone”) beyond the ore zone boundary as defined in the 2021 PFS, highlighting extension of the deposit continuity outside the existing mineral resource area.

Results from the 2022 program further supported drilling results from the 2021 site investigation and have established expansion opportunities for additional resource delineation. Highlighted intercepts supporting the SW Extension Zone include 1.01 g/t Au over 30.0 m in hole SG22-008 from 300 to 330 m; 1.13 g/t Au over 25.0 m in hole SG22-021 from 252 m to 277 m, (including 2.87 g/t Au over 7.0 m from 264 – 271 m); and 1.95 g/t Au over 12 m from 228 to 240 m, and 0.51 g/t Au over 30.0 m from 255 m – 285 m in hole SG22-027.

Mineralization encountered in these drill holes includes heavily disseminated pyrite with local concentrations of semi-massive sulphide within strongly altered and leached host rocks. The favourable lithologies present for gold and silver mineralization are trachytes, porphyries, volcanic breccias, and altered andesites.

The SW Extension Zone has been delineated over approximately 150 m in the southwesterly direction and 550 m in the strike direction, with mineralization remaining open along strike and at depth for further testing. Since the boundaries of this zone have yet to be fully defined, future opportunities exist for step-out drilling and additional exploration to extend the identified mineralization.

Some of the core material collected in this geotechnical drilling program was also used to augment the metal leaching and acid rock drainage (ML/ARD) testwork on the Project.

First Mining sampled approximately 3,125 m of the geotechnical drill core for gold fire assay and ICP analysis, and assay highlights are presented in Table 10‑11.

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Table 10‑11: Select Assay Highlights, 2022 Geotechnical Drilling Program

Hole ID	From (m)	To (m)	Length (m)	Au (g/t)	Ag (g/t)

| SG22-008 | 268 | 279 | 11 | 0.51 | 2.73 |

| and | 300 | 330 | 30 | 1.01 | 8.44 |

| including | 307 | 308 | 1 | 11.10 | 116.00 |

| and | 334 | 347.15 | 13.15 | 0.57 | 4.53 |

| SG22-012 | 165 | 172 | 7 | 0.73 | 2.63 |

| SG22-017 | 8 | 9 | 1 | 17.30 | 17.80 |

| and | 112 | 114 | 2 | 5.12 | 8.95 |

| including | 113 | 114 | 1 | 9.40 | 16.20 |

| and | 521 | 531 | 10 | 0.54 | 1.47 |

| SG22-019 | 89 | 99 | 10 | 0.65 | 1.75 |

| SG22-021 | 78 | 87 | 9 | 1.01 | 23.63 |

| and | 203 | 221.33 | 18.33 | 0.48 | 4.75 |

| and | 222 | 238 | 16 | 0.41 | 3.46 |

| and | 252 | 277.02 | 25.02 | 1.13 | 7.59 |

| including | 264 | 271 | 7 | 2.87 | 15.90 |

| and including | 268 | 271 | 3 | 4.17 | 19.0 |

| and | 530 | 544 | 14 | 0.35 | 0.99 |

| SG22-023 | 134 | 157.2 | 23.2 | 0.35 | 5.23 |

| including | 158 | 164 | 6 | 1.02 | 18.05 |

| and | 211.09 | 222 | 10.91 | 0.85 | 12.61 |

| SG22-027 | 228 | 240 | 12 | 1.95 | 15.45 |

| and | 255 | 285 | 30 | 0.51 | 3.92 |

| and | 295.14 | 296.33 | 1.19 | 6.14 | 30.17 |

| and | 305 | 326.65 | 21.65 | 0.49 | 3.87 |

| SG22-029 | 102.71 | 122.63 | 19.92 | 0.77 | 2.92 |

Note: Reported widths are drilled core lengths; assay values are uncut

10.3.6.2 Acid Rock Drainage (ARD) Drill Program

In 2022, ten diamond drill holes totaling 1,671.2 m were completed to obtain sample material for the ongoing metal leaching/ARD testwork program. These holes targeted data gaps in the available ARD data within the Portage and Camp zones of the deposit, where the additional data could help address uncertainties related to metal leaching. Additional ARD samples were also taken from selected geotechnical holes which intersected the Portage zone.

This drilling primarily targeted areas of waste and low grade within the PFS pit for ARD sample collection. Gold and ICP analysis was completed on 1,557 samples from the ARD drill holes and assay highlights from mineralized intersections include 1.17 g/t Au over 19.0 m from 85 – 104 m in drill hole SP22-ARD-010, 1.75 g/t Au over 16.0 m from 212.0 – 228.0 m in drill hole SP22-ARD-012, and 0.57 g/t Au over 44.0 m from 73.0 – 117.0 m in drill hole SP22-ARD-004. Select assay results are presented in Table 10‑12.

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Table 10‑12: Select Assay Results, 2022 ARD Drilling Program

Hole ID	From (m)	To (m)	Length^1^ (m)	Au (g/t)	Ag (g/t)

| SP22-ARD-004 | 73 | 117 | 44 | 0.57 | 3.65 |

| SP22-ARD-005 | 47 | 55 | 8 | 0.61 | 7.04 |

| and | 91 | 99.4 | 8.4 | 0.67 | 3.19 |

| and | 126 | 133.83 | 7.83 | 0.41 | 2.61 |

| and | 137 | 151 | 14 | 0.63 | 4.99 |

| SP22-ARD-006 | 18 | 25 | 7 | 0.49 | 2.88 |

| and | 94 | 110 | 16 | 0.38 | 2.58 |

| SP22-ARD-007 | 27 | 32.55 | 5.55 | 1.66 | 5.33 |

| including | 30.87 | 31.5 | 0.63 | 6.8 | 39.5 |

| SP22-ARD-008 | 8 | 12 | 4 | 2.39 | 1.82 |

| and | 17 | 21 | 4 | 1.7 | 1.35 |

| including | 19.45 | 21 | 1.55 | 3.66 | 2.9 |

| SP22-ARD-010 | 85 | 104 | 19 | 1.17 | 15.28 |

| SP22-ARD-011 | 129 | 130 | 1 | 9.4 | 3 |

| SP22-ARD-012 | 212 | 228 | 16 | 1.75 | 13.95 |

| including | 224 | 225 | 1 | 13.9 | 178 |

Note: Reported widths are drilled core lengths; assay values are uncut

10.3.6.3 Hydrogeological Drill Program

In 2022, 19 drill holes totalling 411 m were completed as part of ongoing hydrogeological site investigations required to support the Environmental Assessment (EA) at Springpole. Two of these holes were for characterization of the hydrogeological environment in the area between Birch Lake and the northern wall of the proposed open pit (SH22-001 and SH22-002), and the remaining 17 drill holes ('SH22-MW’ hole series) were shallow monitoring well installations which were positioned around proposed mine infrastructure and intended for long-term groundwater data collection.

Gold and ICP analysis were completed on 135 samples from the hydrogeological drill holes, and a total of 130 m was assayed for gold and multi-element. Drill holes intersected unmineralized volcanics units composed of andesites, basalt and tuffs. For the holes that were assayed, no gold mineralization was reported.

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10.3.7 2024 Drill Program

In the fall of 2024, First Mining carried out hydrogeological and exploration drill programs that included 31 diamond drill holes totalling 3,651 m (Table 10‑13).

Table 10‑13: Summary of 2024 Drill Program

Purpose	No of Holes	Metres

| Hydrogeological | 26 | 1,358.76 |

| Exploration | 5 | 2,292.7 |

| Total | 31 | 3,651.46 |

10.3.7.1 Hydrogeological Drill Program

A total of 26 drill holes were completed for hydrogeological purposes, with a total metreage of 1358.76 m. The majority were vertical holes drilled for the installation of shallow monitoring wells around proposed mine infrastructure areas and intended for long-term groundwater data collection. One drill hole (SH24-PW-016C) was drilled for the purposes of completing a pumping test, and additional monitoring wells, SH24-MW-016D to SH24-MW-016G, were installed to act as monitoring points for the pumping test. Two drill holes, SH24-022 and SH24-023, targeted the area between the proposed pit and CDF area. These were deeper, angled holes drilled to depths of approximately 401m and 500 m, respectively.

Under the supervision of First Mining and WSP Canada Ltd. (formerly Wood plc), Maple Leaf Drilling Ltd. conducted the drilling for the shallow hydrogeological holes and Rodren Drilling Services Ltd. conducted the drilling for the deep hydrogeological boreholes. Core was recovered in triple-tube HQ3 for most of the drill holes, except for drill holes SH24-022 and SH24-023 which were drilled with NQ3.

Locations of the hydrogeological holes are illustrated on Figure 10‑4. Holes were assayed for gold and multi-element (ICP); no gold mineralization was encountered in any of the holes except for SH24-023 which returned sporadic gold values up to a maximum of 0.98 g/t.

10.3.7.2 Exploration Drill Program

Five holes totalling 2,293 m were drilled for exploration purposes in a Phase 1 drilling campaign focusing on a 150 m strike area at the Southeast Extension target located at the southeastern boundary of the current mineral resource and proposed open pit wall design. First Mining identified this target through advanced 3D target modelling, which highlighted a theorized easterly adjusted trend to the southern end of the main Portage zone. Data supporting this target model includes televiewer foliation data, geophysical data, historical drill hole data, geological mapping, as well as data from an advanced structural mapping campaign. Drill holes SP24-007, SP24-009, and SP24-011 were designed to test the Phase 1 target shape, and drill holes SP24-008 and SP24-010 were to test a potential extension of the mineralization and key stratigraphic units towards the northeast of the Phase 1 target shape.

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Mineralization in drill holes SP24-007, SP24-009, and SP24-011 is hosted within a silicified and sericitized intermediate volcanic/breccia unit, and medium to coarse-grained trachyte units. Mineralization consists of 2-5% disseminated and fracture-controlled pyrite mineralization occurring coincident with sheared and brecciated structural zones. Minor amounts of galena, sphalerite, chalcopyrite, and tellurides are hosted within quartz carbonate veining that comprises 1-5% of the core. Higher grade intervals (>3 g/t Au) are associated with mineralized brecciated quartz carbonate veining with up to 10% fine-grained disseminated pyrite. Drill holes SP24-008 and SP24-010 intersected key stratigraphic units consisting of megacrystic trachyte and heterolithic breccia. Mineralization intersected in these holes is hosted within carbonate- and silica-altered andesite, trachyte, and heterolithic breccia units, with up to 2% disseminated fine-grained pyrite mineralization.

Highlights from this drilling campaign are from drill holes SP24-007, SP24-011, SP24-009, all of which returned favourable gold and silver grades that are representative of the established resource grade profile. The most northerly hole of the program, SP24-011, returned two broad mineralization intervals including 0.75 g/t Au and 3.30 g/t Ag over 134.2 m, and 0.67 g/t Au and 12.79 g/t Ag over 105.4 m. Assay results for this program are presented in Table 10‑14.

The Springpole Southeast Extension Phase 1 drilling campaign intersected significant widths of continuous mineralization in all holes and validated the exploration target model with further delineation of key stratigraphical units defining forward exploration opportunities. The Southeast Extension Target remains open along strike towards the south and southeast of the main Portage zone and has the potential to add meaningful mineralization extension or additional zones within or near the current PFS proposed open pit shell.

Table 10‑14: Assay Results, 2024 Exploration Drilling Program

Hole ID		From (m)	To (m)	Length (m)	Au g/t	Ag g/t

| SP24-007 | | 50.0 | 51.1 | 1.1 | 6.46 | 34.87 |

| SP24-007 | including | 50.6 | 51.1 | 0.5 | 13.70 | 73.10 |

| SP24-007 | | 53.1 | 56.6 | 3.5 | 0.65 | 5.86 |

| SP24-007 | | 66.2 | 71.0 | 4.8 | 0.73 | 7.46 |

| SP24-007 | | 73.0 | 119.5 | 46.5 | 0.61 | 4.66 |

| SP24-007 | | 125.0 | 139.7 | 14.7 | 0.65 | 1.71 |

| SP24-007 | including | 128.0 | 129.0 | 1.0 | 3.20 | 0.80 |

| SP24-007 | | 148.9 | 149.9 | 1.0 | 0.48 | 0.54 |

| SP24-007 | | 152.9 | 162.5 | 9.6 | 0.40 | 1.99 |

| SP24-007 | | 166.1 | 175.1 | 9.0 | 0.37 | 5.36 |

| SP24-007 | | 179.0 | 187.0 | 8.0 | 0.47 | 4.68 |

| SP24-007 | | 200.0 | 203.0 | 3.0 | 0.31 | 3.61 |

| SP24-007 | | 213.0 | 224.0 | 11.0 | 0.49 | 7.64 |

| SP24-007 | | 237.0 | 250.0 | 13.0 | 0.56 | 5.04 |

| SP24-007 | | 254.0 | 262.0 | 8.0 | 0.41 | 3.64 |

| SP24-007 | | 264.0 | 265.0 | 1.0 | 0.46 | 4.45 |

| SP24-007 | | 270.5 | 272.5 | 2.0 | 0.45 | 3.14 |

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Hole ID		From (m)	To (m)	Length (m)	Au g/t	Ag g/t

| SP24-007 | | 287.35 | 291.0 | 3.6 | 0.47 | 10.39 |

| SP24-007 | | 302.5 | 355.0 | 52.5 | 0.83 | 8.02 |

| SP24-007 | Including | 320.35 | 331.7 | 11.35 | 1.03 | 13.20 |

| SP24-007 | and including | 335.3 | 340.8 | 5.5 | 2.06 | 15.74 |

| SP24-007 | | 359.6 | 364.6 | 5.0 | 0.38 | 8.21 |

| SP24-007 | | 371.3 | 372.8 | 1.5 | 0.84 | 14.52 |

| SP24-007 | | 378.8 | 380.8 | 2.0 | 0.54 | 6.38 |

| SP24-007 | | 404.5 | 405.0 | 0.5 | 0.42 | 4.49 |

| SP24-008 | | 349.0 | 350.0 | 1.0 | 0.35 | 2.17 |

| SP24-008 | | 371.0 | 373.0 | 2.0 | 0.33 | 2.85 |

| SP24-008 | | 392.0 | 393.0 | 1.0 | 0.32 | 2.18 |

| SP24-009 | | 8.0 | 8.75 | 0.8 | 0.57 | 4.66 |

| SP24-009 | | 38.3 | 45.1 | 6.8 | 0.44 | 3.49 |

| SP24-009 | including | 42.6 | 43.1 | 0.5 | 2.41 | 28.50 |

| SP24-009 | | 56.0 | 57.0 | 1.0 | 0.67 | 5.88 |

| SP24-009 | | 64.4 | 69.4 | 5.0 | 0.42 | 1.89 |

| SP24-009 | | 74.2 | 94.0 | 19.8 | 0.38 | 1.69 |

| SP24-009 | | 105.0 | 114.5 | 9.5 | 0.33 | 1.72 |

| SP24-009 | | 118.5 | 136.0 | 17.5 | 1.01 | 10.41 |

| SP24-009 | including | 121.6 | 122.3 | 0.7 | 9.30 | 184.00 |

| SP24-009 | | 139.7 | 150.0 | 10.3 | 0.48 | 1.79 |

| SP24-009 | | 165.8 | 166.8 | 1.0 | 0.32 | 1.56 |

| SP24-009 | | 172.8 | 174.8 | 2.0 | 0.43 | 1.24 |

| SP24-009 | | 186.0 | 187.0 | 1.0 | 0.56 | 0.50 |

| SP24-009 | | 207.0 | 211.0 | 4.0 | 0.22 | 0.83 |

| SP24-009 | | 216.0 | 221.0 | 5.0 | 0.31 | 1.53 |

| SP24-009 | | 226.0 | 228.0 | 2.0 | 0.33 | 2.16 |

| SP24-009 | | 231.0 | 232.0 | 1.0 | 0.44 | 2.21 |

| SP24-009 | | 234.7 | 235.7 | 1.0 | 0.36 | 1.10 |

| SP24-009 | | 238.7 | 240.7 | 2.0 | 0.41 | 6.75 |

| SP24-009 | | 273.0 | 297.7 | 24.7 | 0.89 | 6.95 |

| SP24-009 | including | 280.95 | 287.8 | 6.85 | 1.66 | 13.45 |

| SP24-009 | | 301.7 | 305.7 | 4.0 | 0.51 | 3.12 |

| SP24-009 | | 308.7 | 310.7 | 2.0 | 0.45 | 4.28 |

| SP24-009 | | 315.2 | 337.0 | 21.8 | 0.44 | 4.34 |

| SP24-009 | | 342.0 | 344.0 | 2.0 | 0.68 | 6.88 |

| SP24-009 | | 354.4 | 355.4 | 1.0 | 0.30 | 4.62 |

| SP24-009 | | 374.9 | 375.45 | 0.6 | 0.47 | 0.30 |

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Hole ID		From (m)	To (m)	Length (m)	Au g/t	Ag g/t

| SP24-010 | | 4.1 | 5.0 | 0.9 | 0.31 | 0.32 |

| SP24-010 | | 61.5 | 62.3 | 0.8 | 0.40 | 1.90 |

| SP24-010 | | 75.6 | 77.0 | 1.4 | 0.57 | 6.72 |

| SP24-010 | | 141.5 | 142.5 | 1.0 | 0.55 | 2.55 |

| SP24-010 | | 150.7 | 151.35 | 0.7 | 0.43 | 8.62 |

| SP24-010 | | 204.0 | 205.0 | 1.0 | 0.30 | 3.10 |

| SP24-010 | | 337.9 | 339.0 | 1.1 | 1.72 | 37.77 |

| SP24-010 | | 371.9 | 374.0 | 2.1 | 0.55 | 1.51 |

| SP24-011 | | 23.5 | 157.7 | 134.2 | 0.75 | 3.30 |

| SP24-011 | including | 108.0 | 120.4 | 12.4 | 1.01 | 6.24 |

| SP24-011 | | 170.65 | 185.7 | 15.05 | 0.51 | 2.68 |

| SP24-011 | | 189.1 | 213.3 | 24.2 | 0.79 | 7.41 |

| SP24-011 | | 217.5 | 256.3 | 38.8 | 0.44 | 3.91 |

| SP24-011 | | 261.9 | 262.75 | 0.85 | 0.38 | 6.98 |

| SP24-011 | | 265.1 | 370.5 | 105.4 | 0.67 | 12.79 |

| SP24-011 | Including | 265.1 | 278.0 | 12.9 | 1.00 | 21.18 |

| SP24-011 | and including | 352.2 | 370.5 | 18.3 | 1.12 | 16.33 |

| SP24-011 | | 411.5 | 416.0 | 4.5 | 1.84 | 60.90 |

| SP24-011 | including | 411.5 | 412.2 | 0.7 | 7.22 | 263.00 |

| SP24-011 | | 422.0 | 423.0 | 1.0 | 0.32 | 0.73 |

Note: Reported widths are drilled core lengths; assay values are uncut

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Figure 10‑4: Location of First Mining 2024 Drill Holes

Source: First Mining, 2025

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10.4 First Mining Drilling, Regional Programs

In 2022 and 2023, First Mining completed targeted drill programs at regional prospects within its wider Birch-Uchi mineral tenure. The 2022 exploration drill program was completed at the Swain target, located approximately 7 km southwest of First Mining’s western-most property boundary at the Project. Drilling comprised five diamond drill holes totalling 1,557 m. The purpose of these holes was to drill test areas of historically identified mineralization as well as favourable structural features for gold mineralization. Drill holes were strategically placed to test along the Grace Deformation Zone (GDZ) in areas of prospective gold mineralization based on geology, grab samples, favourable geophysics, and structural architecture. Rodren Drilling of Winnipeg was commissioned by First Mining to complete the drilling. Figure 10‑5 shows the location of the 2022 Swain drill hole locations and drill assay results are presented in Table 10‑15.

The exploration drilling campaign carried out on the Swain property was successful in advancing base datasets and the identification of favourable host rock lithologies over zones of meaningful width with increased alteration, veining and mineralization. Results delineated an area of interest over one kilometre of strike length coincident with the influence of the GDZ structure, which includes drill intercepts of 0.64 g/t Au over 5.6 m, and 0.34 g/t Au over 14.9 m (hole SWL22-001). This initial work identified important indicators that would support follow-up targeting down strike of the GDZ focusing on anomalous soil geochemical values (including 724 ppb Au) at the Hogwarts target and rock grab sample values including 34.7 g/t Au at the Sol d’Or target.

Table 10‑15: Assay Results, 2022 Swain Drilling Program

Hole ID	From (m)	To (m)	Length (m)	Au Grade (g/t)

| SWL22-001 | 63.46 | 78.33 | 14.87 | 0.34 |

| including | 65.68 | 66.07 | 0.39 | 2.51 |

| | 118.86 | 124.5 | 5.64 | 0.64 |

| including | 124 | 124.5 | 0.5 | 4.59 |

| SWL22-002 | 116.1 | 118.3 | 2.2 | 0.69 |

| | 126.3 | 134.9 | 8.6 | 0.26 |

| | 326 | 329 | 3 | 0.23 |

| | 414.7 | 417 | 2.3 | 0.27 |

| SWL22-003 | 163.97 | 165.2 | 1.23 | 1.18 |

| SWL22-004 | 208 | 215.3 | 7.3 | 0.09 |

| | 245.65 | 249.2 | 3.55 | 0.27 |

| SWL22-005 | 16 | 18.33 | 2.33 | 0.15 |

| | 42 | 43.06 | 1.06 | 0.43 |

| | 54 | 57 | 3 | 0.13 |

| | 62 | 66 | 4 | 0.14 |

| | 75 | 89 | 14 | 0.16 |

| | 94 | 104 | 10 | 0.23 |

Note: Reported widths are drilled core lengths; assay values are uncut

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Figure 10‑5: Location of 2022 Drilling, Swain Target

Source: First Mining, 2022

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During 2023, First Mining completed exploration drilling on its regional Horseshoe and Saddle targets located in the Birch-Uchi mineral tenure adjacent to the Springpole Project. A total of 10 diamond drill holes were completed, for a total of 2,430.7 m of drilling. Drill holes were strategically placed to test key structures that are in close proximity to the SDZ and HDZ. These areas are indicative of prospective gold mineralization, based on geology, grab samples, favourable geophysics, and structural architecture. Drilling activities were conducted by Rodren Drilling of Winnipeg.

The Saddle target is located approximately 12 km southwest of the Springpole deposit. This was the first drilling by the Company at this target and comprised five drill holes totalling 842 m. Results indicated the presence of a significant gold mineralization system, with assay highlights including 0.92 g/t Au over 114.0 m in hole SAT23-001, and 0.75 g/t Au over 57.7 m in drill hole SAT23-002.

The Horseshoe target is located approximately 10 km east-southeast of the Springpole Project. Assay highlights from the five holes drilled at Horseshoe include 0.48 g/t Au over 48.5 m in drill hole HOR23-001, 0.54 g/t Au over 57 m in drill hole HOR23-002, and 0.64 g/t Au over 24 m and 0.45 g/t Au over 14.5 m in drill hole HOR23-003.

Hole locations for the 2023 regional drilling are shown on Figure 10‑6 and Figure 10‑7, and assay results are in Table 10‑16.

Table 10‑16: Assay Results, 2023 Regional Drilling Program

Hole ID	From (m)	To (m)	Length (m)	Au (g/t)

| SAT23-001 | 4 | 118 | 114 | 0.92 |

| including | 7 | 16 | 9 | 1.24 |

| including | 23.65 | 24.88 | 1.23 | 4.41 |

| including | 33 | 46 | 13 | 1.36 |

| including | 107 | 116.56 | 9.56 | 1.23 |

| and | 123 | 133.95 | 10.95 | 0.54 |

| SAT23-002 | 51 | 108.7 | 57.7 | 0.75 |

| including | 93 | 104.5 | 11.5 | 1.11 |

| and | 119 | 121 | 2 | 0.55 |

| SAT23-003 | 132 | 132.7 | 0.7 | 1.12 |

| SAT23-004 | 7.5 | 62 | 54.5 | 0.58 |

| including | 49 | 50 | 1 | 5.36 |

| including | 61 | 62 | 1 | 3.23 |

| and | 139 | 143 | 4 | 0.47 |

| SAT23-005 | 27 | 32.65 | 5.65 | 0.3 |

| and | 50 | 69 | 19 | 0.36 |

| HOR23-001 | 22 | 25 | 3 | 2.01 |

| and | 70.6 | 74 | 3.4 | 0.3 |

| and | 159 | 194 | 35 | 0.57 |

| and | 221 | 269.5 | 48.5 | 0.48 |

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Hole ID	From (m)	To (m)	Length (m)	Au (g/t)

| including | 256 | 256.5 | 0.5 | 4.29 |

| HOR23-002 | 27 | 29 | 2 | 0.7 |

| and | 68 | 72 | 4 | 0.49 |

| and | 90 | 92 | 2 | 0.48 |

| and | 105 | 107 | 2 | 1.35 |

| and | 158 | 215 | 57 | 0.54 |

| including | 193 | 203 | 10 | 0.9 |

| including | 211.06 | 215 | 3.94 | 1.05 |

| HOR23-003 | 48.4 | 49.5 | 1.1 | 2.02 |

| and | 74 | 75 | 1 | 1.06 |

| and | 82 | 83 | 1 | 1.41 |

| and | 100 | 101 | 1 | 1 |

| and | 134 | 140 | 6 | 1.38 |

| including | 135 | 136 | 1 | 4.49 |

| and | 150 | 174 | 24 | 0.64 |

| including | 153.8 | 154.4 | 0.6 | 5.43 |

| including | 173 | 174 | 1 | 4.86 |

| and | 183 | 197.5 | 14.5 | 0.45 |

| including | 195.5 | 197.5 | 2 | 2.23 |

| and | 217 | 221 | 4 | 0.94 |

| and | 234 | 236 | 2 | 0.76 |

| and | 263 | 268.4 | 5.4 | 0.32 |

| and | 355.85 | 357.25 | 1.4 | 0.57 |

| HOR23-004 | 80 | 81 | 1 | 0.96 |

| and | 272 | 273.65 | 1.65 | 0.46 |

| HOR23-005 | No significant assays | | | |

Note: Reported widths are drilled core lengths; assay values are uncut

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Figure 10‑6: Location of 2023 Drilling at Saddle Target

Source: First Mining, 2023

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Figure 10‑7: Location of 2023 Drilling at Horseshoe Target

Source: First Mining, 2023

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10.5 Core Resampling Program 2019 – 2020

During the winter of 2019 - 2020, First Mining initiated a program of core re-sampling in order to quantify the sulphur content of the in-pit material. A total of 8,610 samples were collected for total sulphur assays, along with 471 samples collected for bulk density determination.

10.6 Drill Collar Surveying

All historical holes drilled prior to 2010 were surveyed using various earth projections, either NAD27 (North American Datum 1927) Canada, WGS or NAD83 projections. In September 2006, W.J. Bowman Ltd. of Dryden, Ontario, surveyed 275 historic drill hole collars from collar numbers BL-1 through BL-373. Historical collar locations collected in NAD83 have been converted to the UTM WGS84 projection in the current drill database.

For the 2007 and 2008 drill programs, the drill hole collars were located and surveyed using a handheld GPS and recorded in UTM NAD27 Canada projection. All the collar survey information was converted to WGS84 and field checked against collar locations using handheld Trimble GeoXH DGPS.

The 2010 to 2012 drill hole collars were initially surveyed using handheld GPS devices. During the initial phases of the offshore 2010 drill program, drill hole collars on the lake ice were surveyed by handheld, real-time differential GPS with an average accuracy of 4 to 5 m and recorded in UTM NAD27 Canada projection. On-shore drill holes were initially located with handheld GPS and once the drill hole was complete, the hole location was temporarily marked; subsequently, the collars were surveyed using a Trimble GeoXH handheld DGPS device with an external antenna giving sub-metre (~10 cm) location accuracy.

For the offshore drill programs (2011 to 2013), drills were mounted on barges, the drill sites were marked by floating buoy and located using the Trimble GeoXH from a boat. All onshore drill collars were located and subsequently surveyed using the Trimble GeoXH. At the beginning of the winter 2011 drill program, the UTM WGS84 projection was adopted as the standard for surveying drill collars and other surface landmarks. All previously recorded UTM measurements were converted accordingly.

All drill site locations for inclined drill holes, onshore or offshore on the ice, were marked using two to four painted laths aligned along strike either side of the proposed drill hole location. These laths were used as fore- and back-sights for setting the drill location and orientation. Inclination of the drill hole was checked on the drill head, prior to commencing drilling, using either a Brunton compass or inclinometer accurate to half of one degree.

All drill holes completed by First Mining from 2016 to 2024 have been surveyed using the UTM WGS84 projection.

10.7 Oriented Core Surveying

For the historical drilling, oriented core measurements were collected from a total of 44 drill holes (added to which are the six holes drilled in 2013). Oriented core is used to evaluate the structural geology by allowing the geologists to measure the real angular relationships, as opposed to apparent angles. The tool used was the ACT 2 from Reflex Technologies. This system is fully digital, using solid-state three-axis accelerometers to record the orientation of the core-barrel when the core is taken off-bottom at the end of each drill run. There were significant problems encountered during the winter 2011 drill program due to tool failures. Some oriented core information was collected, but too little to be of widespread use. The 2013 program successfully collected data from all six holes, although the data was not processed as part of this work program.

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Where down-hole poor ground conditions were encountered, the oriented core tool proved to be of little value due to the incompetent nature of intensely altered and mineralized rock. Wherever competent rock was encountered, oriented core data was collected.

First Mining collected oriented core information using a Reflex ACTIII tool for all holes in its 2022 and 2023 regional drill programs, as well as for all exploration holes and the two deep hydrogeological holes (SH24-022 and SH24-023) from the 2024 drill programs.

10.8 Down Hole Surveying

10.8.1 Gold Canyon Programs (Historical)

All drill holes during the 2010 drill program were surveyed using a Reflex Technologies single EZ-Shot or EZ-Trax down-hole survey system. Drill holes were surveyed once completed – this procedure was used because of the chance that bad ground conditions encountered in the drill holes increased the risk of cave-in when pulling the drill string backwards to conduct a survey. A cave-in can result in increased cost due to time spent reaming the drill hole clean back to the bottom, or from the possibility of sticking the drill string, causing loss of drilling tools. The presence of magnetite in banded iron formation and relatively unaltered trachyte or greenstone caused interference with respect to azimuth readings and also the azimuth of the drill traces. This required many repetitions of the down-hole survey readings, which in some cases resulted in lesser consistent data readings.

For the 2011, 2012, and 2013 programs, the Reflex Down-Hole Gyro survey system was adopted with the EZ-Trax or EZ-Shot down-hole survey tools as back up. The Reflex Gyro is built around a digital micro-gyro, which consists of a silicon sensor chip and an integrated circuit assembled in a ceramic (non-magnetic) package. The gyro provides directional data (azimuth and dip) at any interval from inside the drill rods. This system is used to provide azimuth and inclination data in rocks with strong magnetic fields, because the gyros operate independently of the earth’s magnetic field. The system also records ambient temperature as well as collecting basic gravity measurements. The Reflex gyro system was successfully applied to the majority of the 2011 and 2013 drill programs.

Data recorded from the down-hole surveys was incorporated into 3D planning and modelling.

10.8.2 First Mining Programs

The Reflex Gyro tool was used in the 2016 drilling program with vertical checks conducted on the vertical holes to ensure that there were no large deviations in dip during drilling.

No downhole surveying was completed in 2018 on the geotechnical holes targeting the dike foundations.

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In the 2020 program, down hole surveying was conducted on drill holes SGH20-001 to SGH20-009, SM21-001 to SM20-001, and SPW20-001. Surveying was conducted using either Reflex EZ-Gyro or Reflex EZ-Trac tools.

In the 2021 drill program, downhole surveying was conducted on drill holes SM21-001 to SM21-007, SP21-001 to SP21-006, SC21-048, SC21-050, SC21-051, SCH21-049, SCH21-052, SG21-001 to SG21-005, SG21-007, SGH21-006 and SP-ARD-001 to SP-ARD-003. No downhole surveys were conducted for the other vertical drill holes in the program. Surveying was conducted using either a Reflex EZ-Trac or a Reflex EZ-Shot tool approximately every 50 m with an additional survey at the start of the hole within 15 m of surface.

Downhole surveys during the 2022 drill programs at the Springpole Project were conducted on angled drill holes. For 9 drill holes (SG22-008, SG22-017, SG22-019, SG22-021, SG22-023, SG22-027, SG22-029, SP22-ARD-005, SP22-ARD-009) an EZ Shot was used at a 30 m frequency, with the initial measurement taken between 15-20 m. For drill holes SP22-ARD-011, SP22-ARD-012, and SP22-ARD-013, an EZ-Gyro was used, also conducted at 30 m intervals. Upon completion of drilling, a continuous, more detailed Sprint Gyro survey was completed on select holes (SG22-019, SG22-021, SG22-023, SG22-027, SG22-029, SP22-ARD-011, and SP22-ARD-012).

For the 2022 regional drilling at the Swain property, Reflex’s TN14 drill aligner was used for drill hole alignment prior to drilling. Downhole surveys were conducted on all the five holes drilled at the Swain property. Single shot gyro points were used for downhole deviation monitoring, and then each hole was completed with a continuous gyro survey for the entire hole length. Reflex’s Gyro SPRINT-IQTM was used for surveying each hole.

For the 2023 regional drilling program, downhole surveys were completed on all holes. Survey tools provided by Devico were the DeviGyro and DeviCounter for the downhole surveying, and the DeviAligner for the rig alignment. Downhole surveying was undertaken with a continuous gyro survey and the end-of-hole survey was completed with the DeviGryo and DeviCounter tools. The data was collected using IMDEX’s cloud-based collection tool called the HUB. Data was uploaded direct the HUB from the DeviAligner at the drill rig and available for the geologist on and off site immediately. The logging geologist imported the data into Fusion through DHLogger for inclusion in the drill database.

In the 2024 drill program, downhole surveys were completed on the 2 deep hydrogeological holes SH24-022 and SH24-023, and exploration holes SP24-007 to SP24-011. Survey tools were the same as the 2023 program (DeviGyro and DeviCounter) and downhole surveying was done at depths of 15 m, 50 m, and every 50 m downhole, with a continuous gyro survey completed at the end of each hole, and an end-of-hole survey completed with the DeviGryo and DeviCounter tools. Data collection was the same as for the 2023 program, i.e. via the IMDEX Hub and subsequently imported into the Fusion drill database.

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A summary of the downhole survey status for all the First Mining drill programs is provided in Table 10‑17.

Table 10‑17: Downhole Survey Summary, First Mining Drill Programs

Year	Program	Test Type	Holes

| 2020 | Geotechnical | No downhole surveys | BH-PS-07R1, BH-PS-08R1, BH-PS-09R1, BH-TMF-04, BH-TMF-05, BH-TMF-15, BH-WSF1-01, BH-WSF1-02, BH-WSF1-03, BH-WSF1-11, BH-WSF1-12, BH-WSF2-06, BH-WSF2-13, BH-WSF2-14, SGH20-010 |

| 2020 | Condemnation | No downhole surveys | SC20-010, SC20-012, SC20-013, SC20-014, SC20-015, SC20-016, SC20-019, SC20-020, SC20-023, SC20-024, SC20-025, SC20-026, SC20-027, SC20-028, SC20-037, SC20-038, SC20-039, SC20-040, SC20-046, SC20-047 |

| 2020 | Hydrogeology | Reflex EZ Trac | SPW20-001 |

| 2021 | Geotechnical | No downhole surveys | BH-CAMP-32, BH-CAMP-33, BH-PP-24, BH-PP-25, BH-PP-26, BH-PP-27, BH-PP-28, BH-PP-29, BH-PP-30, BH-PP-31, BH-PP-34, BH-SP-039, BH-SP-040, BH-TMF-35, BH-TMF-36, BH-TMF-37, BH-TMF-38, BH-Q-18, BH-Q-19, BH-Q-20, BH-Q-21, BH-Q-22, BH-Q-23 |

| 2021 | ARD | Reflex EZ Shot | SP-ARD-001, SP-ARD-002, SP-ARD-003 |

| 2021 | Hydrogeology | No downhole surveys | SPW20-002 |

| 2022 | Geotechnical | No downhole surveys | SG22-009, SG22-010, SG22-011, SG22-012, SG22-013, SG22-014, SG22-015, SG22-022, SG22-024, SG22-025, SG22-026, SG22-028, SG22-030, SG22-031, SG22-032, SG22-034, SG22-035, SG22-036, SG22-038, SG22-039, SG22-040, SG22-041, SG22-042, SG22-043 |

| 2022 | Hydrogeology | No downhole surveys | SH22-001, SH22-002, SH22-MW-001A, SH22-MW-001B, SH22-MW-002A, SH22-MW-002B, SH22-MW-003A, SH22-MW-003B, SH22-MW-004A, SH22-MW-004C, SH22-MW-004D, SH22-MW-005A, SH22-MW-006B, SH22-MW-007A, SH22-MW-008A, SH22-MW-008B, SH22-MW-009A, SH22-MW-009B |

| 2022 | ARD | Reflex EZ Shot | SP22-ARD-005, SP22-ARD-009 |

| 2022 | ARD | Sprint Gyro | SP22-ARD-011, SP22-ARD-012, SP22-ARD-004^1^, SP22-ARD-006^1^, SP22-ARD-007^1^, SP22-ARD-008^1^, SP22-ARD-010^1^ |

| 2022 | ARD | Reflex EZ Gyro | SP22-ARD-013 |

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Year	Program	Test Type	Holes

| 2024 | Hydrogeology | No downhole surveys | SH24-MW-011, SH24-MW-012A, SH24-MW-012B, SH24-MW-013A, SH24-MW-013B, SH24-MW-014A, SH24-MW-014B, SH24-MW-015A, SH24-MW-015B, SH24-MW-016A, SH24-MW-016B, SH24-MW-016D, SH24-MW-016E, SH24-MW-016F, SH24-MW-016G, SH24-MW-017A, SH24-MW-017B, SH24-MW-018, SH24-MW-019, SH24-MW-020A, SH24-MW-020B, SH24-MW-021A, SH24-MW-021B, SH24-PW-016C |

Note:

1.	Vertical holes were checked with the gyro tools when available

10.9 Drilling Pattern and Density

The overall exploration drilling pattern over the Springpole deposit approximates a 50 m grid along the long axis of the Portage zone and approximately 45 to 65 m spacing down the dip of the mineralized zone. The QP is of the opinion that the drill spacing, and density is appropriate for this type of deposit and style of mineralization.

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11 SAMPLE PREPARATION, ANALYSES, AND SECURITY

11.1 Introduction

The following sections outlining sample preparation, analysis and security refer only to drill programs carried out by Gold Canyon and First Mining, and not to drilling conducted by prior operators. Any drilling completed prior to First Mining’s involvement in the Project is classed as historical.

11.2 Core Drilling and Sampling

11.2.1 Historical Programs (Gold Canyon)

Detailed descriptions of the drill core were carried out under the supervision of a senior geologist, a member in good standing of the APGO (Association of Professional Geologists of Ontario) and AIPG (American Institute of Professional Geologists). The core logging was carried out on-site in a dedicated core logging facility at the Springpole exploration camp. For Gold Canyon drill programs prior to 2012, and the 2013 program, drill log data was recorded onto paper logs that were later scanned and made digital. For the 2012 programs, drill holes appear to have only been digitally logged.

Core was laid out 30 to 40 boxes at a time. First, the core was photographed in 15 m batches prior to logging or sampling. This was followed by a geotechnical log that recorded quantitative and qualitative engineering data including detailed recovery data and rock quality designation. Any discrepancies between marker blocks and measured core length were addressed and resolved at this stage. The core was then marked up for sampling.

For Gold Canyon’s 2010 and 2011 drill programs, all the drill core intervals were sampled using sample intervals of 1 m. During the 2012 drilling program, Gold Canyon changed its standard sample length from 1 m to 2 m lengths. However, in zones of poor recovery, 1.5 m or 3 m samples were sometimes collected. Samples over the standard sample length were typically half core samples and whole core was generally only taken in intervals of poor core recovery across the sampled interval. Sampling marks were made on the core and sample tickets were stapled into the core boxes at the beginning of each sample interval.

Quality control samples were inserted into the sample stream. Inserting quality control samples involved the addition of certified blanks, certified gold standards, and field and laboratory duplicates. Field duplicates were collected by quartering the core in the sampling facility on-site. Laboratory duplicates were collected by splitting the first coarse reject and crushing and then generating a second analytical pulp. Blanks, standards, and duplicates made up on average 10% of the total sample stream. Sample tickets were marked blank, field or laboratory duplicate, or standard, and a sample tag was stapled into the core box within the sample stream.

Geological descriptions were recorded for all core recovered. Separate columns in the log allow description of the lithology, alteration style, intensity of alteration, relative degree of alteration, sulphide percentage, rock colour, vein type, and veining density. A separate column was reserved for written notes on lithology, mineralization, structure, vein orientations/relations etc. The header page listed the hole number, collar coordinates, final depth, start/end dates, and the name of the core logging geologist.

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11.2.2 First Mining Programs

Core preparation, quantitative measurements (metre marks, RQD, recovery, magnetic susceptibility and specific gravity) and core logging were completed on-site in a dedicated facility on site at the Springpole exploration camp. General procedures for core preparation and sampling were the same for all First Mining programs. The core preparation and quantitative measurements were undertaken by geotechnicians or logging geologists. Detailed geological logging and sample interval selection was carried out by logging geologists and reviewed by a senior geologist and a member in good standing of the APGO (Association of Professional Geologists of Ontario).

For the 2016 drill program, drill logging data was recorded onto paper logs and later scanned and digitized. For all drill programs since 2018, First Mining have completed drill core logging using Datamine ‘DHLogger’ software. Logging observations were recorded directly in DHLogger on laptop computers and uploaded into a centralized Fusion SQL drilling database (‘Fusion’). RQD, SG, magnetic susceptibility, and downhole survey data were recorded in Excel templates and then imported into Fusion through the DHLogger interface.

When core was received in the core shack it was laid out on the benches, checked for condition and any run block errors, and cleaned and pieced together. RQD and recovery measurements were then taken and recorded into an Excel sheet.

For programs where the drill core was oriented, after being checked for run block errors, the orientation marks were checked for quality, the core pieced together checking for ‘locking pieces’, and the angular difference between runs measured. If three or more consecutive runs locked together and had less than 10 degrees of cumulative difference an orientation line was drawn. This orientation line was then used for oriented measurements (alpha, betas, trend, plunge) to be recorded (contacts, veins angles, structural measurements). RQD was also taken along this orientation line and samples cut parallel to the line.

Detailed logging data collected directly in DHLogger includes lithology, alteration, veining, structure, and mineralization observations. Sample IDs and intervals were recorded by the geologist in an Excel spreadsheet that was later imported into DHLogger. The minimum sampling length for core samples was generally 0.3 m and maximum length was 1.5 m. Samples were selected taking lithological boundaries into account as well as reducing sample widths over areas with increased mineralization. Sample booklets were filled out with the sample intervals including inserted quality control samples. The sample booklets contain three tags, one remains in the book as a record, the second tag is stapled into the core box at the start of each sampling interval, and the third tag accompanies the core sample to the assay lab and only lists the sample ID and instructions for lab duplicates (crush and pulp). For QA/QC samples, stickers from the standard packaging were affixed to the corresponding sample tag that remains in the sample book, with a second tag stapled to the core box immediately after the prior sample tag.

QA/QC samples were inserted into the sample stream at preset positions, with blanks inserted every 30 samples, standards every 25 samples, and a duplicate between every standard and/or blank rotating between field, crush, and pulp duplicate types. Field duplicates were taken onsite by quartering the half core sample being sent to the lab for assay and placing each half in its own sample bag with its own sample ID, with the second sample in the sequence coded as the duplicate. For the crush and pulp duplicates, an empty poly bag with a sample tag was included within the poly bag of the sample to be duplicated. The type of duplicate required was communicated to the lab. The lab then created a split of the sample material at either the crushing or pulverising stage as indicated. One half of the split was assigned the original sample ID and the other half was assigned the sample ID from the tag provided in the empty bag.

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Beginning in 2024, First Mining began collecting magnetic susceptibility and specific gravity (relative density) data from drill core samples. Magnetic susceptibility was recorded every metre using a KT-10 magnetic susceptibility meter. These readings were entered into an Excel template and the data was reviewed by the logging geologist and imported into DHLogger. For density, water immersion measurements were collected every 25 m with a minimum of one measurement in each representative rock type per hole. Standard weights were used to calibrate the scale (AnD EJ-6100) before use for data collection. Every 25 m a piece of core approximately 15 cm in length was selected. The whole piece of core was weighed, the dry weight recorded, then the piece of core was weighed while fully submerged in water. The water temperature was recorded in the Excel table and the relative density of the rock was calculated. This data was then reviewed by the logging geologist and any ‘outlier’ values beyond the expected range were checked and retaken. The wet weight, dry weight, temperature and calculated density value were then imported into DHLogger.

In 2024, a check program was undertaken by First Mining to ensure the accuracy of these field measurements. For seven samples from the 2024 drill program, the half core remaining after sampling was re-measured for relative density then that half core sample was shipped to AGAT labs where immersion density testing was conducted on the same pieces of core. The corresponding lab and field measurements from these check samples showed an excellent correlation of 0.97.

After the logging and quantitative measurements were taken, the core was photographed in sets of four to five boxes both wet and dry. Aluminum tags embossed with the hole ID, box number, and box interval (from/to) were prepared and stapled onto the ends of each core box.

11.3 Core Sampling, Handling and Chain-of-Custody

11.3.1 Historical Programs (Gold Canyon)

Following the logging and core marking procedures described in the previous sub-section, the core was passed to the sampling facility. Core sampling was performed by experienced sampling technicians, (technicians were from Ackewance Exploration & Services of Red Lake, Ontario), and quality control was maintained through regular verification by on-site geologists. Core was broken, as necessary, into manageable lengths. Pieces were removed from the box without disturbing the sample tags, were cut in half lengthwise with a diamond saw, and then both halves were carefully repositioned in the box. When a complete hole was processed in this manner, one half was collected for assay while the other half remained in the core box as a witness. The remaining core in the boxes was then photographed. All logs and photographs were then submitted to the senior geologist/project manager for review and were backed up and archived.

The sampling technician packed one half of the split core sample intervals into transparent vinyl sample bags that were sequentially numbered to match the sample number sequences in the sample tag booklets used by the core-logging geologists. The numbered, blank portion of the triplicate sample tag was placed in the bag with the sample; the portion that was marked with the sample interval remained stapled into the bottom of the core box at the point where the sample interval begins. Sample bags were then sealed with plastic tags. Sealed sample bags were packed into rice sacks five samples to a bag. All sacks were individually labeled with the name of the company, number of samples contained therein, and the number sequence of the samples therein. Sacks were assigned sequential numbers on a per shipment basis. A project geologist then checked the sample shipment and created a shipping manifest for the sample batch. A copy of this manifest was given to the Project manager and a copy was sent along with the sample shipment. A copy of the sample shipment form was also sent via e-mail to the analytical laboratory.

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The project geologist prepared the sample submission form for the assay laboratory. This form identified the number of sample sacks as well as the sequence of sample numbers to be submitted. Due to the remote location, the shipment was then loaded on to a plane or helicopter and flown direct to Red Lake where representatives of the commercial analytical laboratory met the incoming flight and took the samples to the laboratory by pickup truck.

Once at the laboratory, a manager checked the rice sacks and sample numbers on the submission form. The laboratory then split the received sample manifest into batches for analysis, assigned a work order to the batch, and sent a copy of the mineral analysis acknowledgement form to the Project manager.

Aluminum tags embossed with the hole number, box number, and box interval (from/to) were prepared and stapled onto the ends of each core box. Core boxes were cross stacked on pallets and then moved to on-site storage.

11.3.2 First Mining Programs

Following the logging and core marking procedures described in Section 11.2, the core was moved to the core cutting facility. Core cutting and sampling were performed by experienced geotechnicians, and quality control was maintained through regular verification by on-site geologists. Core was broken, as necessary, into manageable lengths, and a red lumber crayon was traced along the line where the core meets the core box to create a guide-line. The core was cut in half lengthwise along this line using a core saw with a diamond blade. One half was returned to the core box as a permanent record and the other half was placed into a poly bag labeled with the corresponding sample ID and accompanied by the sample tag. Care was taken to ensure that the half remaining in the box was aligned. The second sample tag copy was stapled into the bottom of the core box where the sample interval begins. Sample bags were then sealed with zipties before being packaged into rice bags, five samples to a bag. A geologist then checked the sample shipment, attached a hard copy of the sample list and associated lab paperwork and arranged for transport offsite. A copy of the sample list and request for analysis was also sent via e-mail to the analytical laboratory. Once at the laboratory, the samples were unpacked and checked and an acknowledgement email was sent to First Mining.

Core boxes were cross-stacked on pallets and then moved to on-site storage locations.

11.4 Sample Security

For both historical and First Mining programs, core samples collected at the drill site were held in closed core boxes sealed with tape; at various times of day the boxes were delivered to the core logging facility. All core logging, sampling and storage took place at the Springpole Project site. Following the logging and marking of core (described in the preceding sections), all core preparation and sampling was performed by technicians (for Gold Canyon’s drill programs, technicians were from Ackewance of Red Lake, Ontario) under the supervision of the Project manager, or by company geologists. All on-site sampling activities were directly supervised by the Project manager or geologist.

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11.5 Sample Preparation and Analytical Procedures

11.5.1 Historical Programs (Gold Canyon)

11.5.1.1 Analytical Laboratories

All gold assay work from the Gold Canyon drill programs between 2010 and 2013 was performed by SGS Laboratories in Red Lake, Ontario. Silver and multi-element assays for the Gold Canyon drill programs were performed by the SGS Don Mills laboratory in Toronto, Ontario. The SGS facilities are certified and conform to requirements CAN-P-1579 and CAN-P-4E (ISO/IEC 17025:2005). Certification is accredited for precious metals including gold and silver and 52 element geochemical analyses.

11.5.1.2 Analytical Procedures

All samples received by SGS Red Lake were processed through a sample tracking system that is an integral part of the company’s laboratory information management system. This system utilizes bar coding and scanning technology that provides complete chain of custody records for every stage in the sample preparation and analytical process.

Samples were dried and then crushed to 70% of the sample passing 2 mm (-70 mesh). A 250 g sample was split off the crushed material and pulverized to 85% passing 75 microns (200 mesh). A 30 g split of the pulp was used for gold fire assay and a 2 g split was used for silver analysis. Crushing and pulverizing equipment was cleaned with barren wash material between sample preparation batches and, where necessary, between highly mineralized samples. Sample preparation stations were also equipped with dust extraction systems to reduce the risk of sample contamination. Once the gold assay was complete, a pulp was sent to the SGS Toronto facility for silver and possibly for multi-element geochemical analysis.

As part of the standard internal quality control procedures used by the laboratory, each batch of 75 Springpole core samples included four blanks, four internal standards, and eight duplicate samples. In the event that any reference material or duplicate result would fall outside the established control limits, the sample batches would be re-assayed.

Pulps and rejects from core samples from these historical programs, where still available, are being kept in storage by First Mining.

11.5.1.3 Gold, Silver and Multi-Element Analysis

Prepared samples were analyzed for gold by fire assay with atomic absorption finish. Samples returning assays in excess of 10 g/t Au were re-analyzed with a gravimetric finish.

Prepared pulp samples shipped from SGS Red Lake to SGS Toronto were analyzed for silver by three-acid digestion with atomic absorption finish.

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During the winter 2010 program, prepared samples were analyzed for 52 elements by acid digestion (3:1 HCl : HNO3). The list of elements is included in Table 11‑1.

Table 11‑1: SGS Multi-Element Analysis Method ICM14B – Detection Limits

Element	Limits	Element	Limits	Element	Limits

| Ag | 0.01 – 10 ppm | Hg | 0.01 ppm - 1% | Se | 1 ppm - 0.1% |

| Al | 0.01 - 15% | In | 0.02 ppm - 0.05% | Sn | 0.3 ppm - 0.1% |

| As | 1 ppm - 1% | K | 0.01 - 25% | Sr | 0.5 ppm - 1% |

| B | 10 ppm - 1% | La | 0.1 ppm - 1% | Ta | 0.05 ppm - 1% |

| Ba | 5 ppm - 1% | Li | 1 ppm - 5% | Tb | 0.02 ppm – 1% |

| Be | 0.1 ppm - 0.01% | Lu | 0.01 ppm - 0.1% | Te | 0.05 ppm - 0.1% |

| Bi | 0.02 ppm - 1% | Mg | 0.01 - 15% | Th | 0.1 ppm - 1% |

| Ca | 0.01 - 15% | Mn | 2 ppm - 1% | Ti | 0.01 - 15% |

| Cd | 0.01 ppm - 1% | Mo | 0.05 ppm - 1% | TI | 0.02 ppm - 1% |

| Ce | 0.05 ppm - 0.1% | Na | 0.01 - 15% | U | 0.05 ppm - 1% |

| Co | 0.1 ppm - 1% | Nb | 0.05 ppm - 0.1% | V | 1 ppm - 1% |

| Cr | 1 ppm - 1% | Ni | 0.5 ppm - 1% | W | 0.1 ppm - 1% |

| Cs | 0.05 ppm - 0.1% | P | 50 ppm - 1% | Y | 0.05 ppm - 1% |

| Cu | 0.5 ppm - 1% | Pb | 0.2 ppm - 1% | Yb | 0.1 ppm - 0.01% |

| Fe | 0.01% - 15% | Rb | 0.2 ppm - 1% | Zn | 1 ppm - 1% |

11.5.2 First Mining 2016, 2018 and 2020 (Metallurgical) Programs

11.5.2.1 Analytical Laboratories

All gold assay work in the 2016 and 2018 drill programs, as well as in the 2020 metallurgical program was completed by SGS Laboratories in Red Lake, Ontario. Multi-element assays including silver for these drill programs were completed by the SGS laboratory in Vancouver. The SGS facilities are certified and conform to requirements CAN-P-1579 and CAN-P-4E (ISO/IEC 17025:2005). Certification is accredited for precious metals including gold and silver and 52 element geochemical analyses.

First Mining has attested that there is no commercial nor other type of relationship between First Mining and SGS Laboratories that would adversely affect the independence of SGS Laboratories.

11.5.2.2 Analytical Procedures

All samples received by SGS Red Lake were processed through a sample tracking system which utilizes bar coding and scanning technology that provides complete chain of custody records for every stage in the sample preparation and analytical process.

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Samples were dried and then crushed to 70% of the sample passing 2 mm (-70 mesh). A 250 g sample was split off the crushed material and pulverized to 85% passing 75 microns (200 mesh). A 30 g split of the pulp was used for gold fire assay and a 2 g split was used for silver analysis. Crushing and pulverizing equipment was cleaned with barren wash material between sample preparation batches and, where necessary, between highly mineralized samples. Sample preparation stations were also equipped with dust extraction systems to reduce the risk of sample contamination. Once the gold assay was complete, a pulp split was sent to the SGS Vancouver facility for multi-element geochemical analysis including silver.

As part of their standard internal quality control procedures, for each batch of 75 Springpole core samples SGS would include four blanks, four internal standards, and eight duplicate samples. In the event that any reference material or duplicate result fell outside the established control limits, the sample batches would be re-assayed.

Pulps and rejects from the First Mining core samples from these programs are being kept in storage by First Mining.

11.5.2.3 Gold, Silver and Multi-Element Analysis

Prepared samples were analyzed for gold by fire assay with atomic absorption finish. Samples returning assays in excess of 10 g/t Au were re-analyzed with a gravimetric finish.

Samples from the 2016 to 2020 drill programs by First Mining were analyzed for multi-element by ICP with aqua regia digestion (ICM14B). A summary of the analytical methods used in all First Mining drill programs is presented in Table 11‑2.

Table 11‑2: Analytical Methods, First Mining Drill Programs

Program Year	Laboratory	Au Lab Package	Ag /Multi Element Lab Package

| 2016 | SGS | GEFAA313 (Exploration grade analysis, 30g, Fire Assay, AAS finish) | GEICP13B (Two Acid/Aqua Regia/ICP-AES Package, 34 Elements) |

| | | GOFAG303 (30g, Fire Assay, Gravimetric Finish) | GEICM14B (Two Acid/Aqua Regia/Combined ICP-AES and ICP-MS package, 51 Elements) |

| | | | GOICP95A (Sr) (Sodium peroxide fusion/ICP-AES) |

| | | | GEICM90A (Ce) (Sodium peroxide fusion/ICP-AES and ICP-MS, 55 elements) |

| 2018 | SGS | GEFAA313 (Exploration grade analysis, 30g, Fire Assay, AAS finish) | GEICM14B (Two Acid/Aqua Regia/Combined ICP-AES and ICP-MS package, 51 Elements) |

| 2020 | SGS | GEFAA313: Exploration grade analysis, 30g, Fire Assay, AAS finish | GEICM14B (Two Acid/Aqua Regia/Combined ICP-AES and ICP-MS package, 51 Elements) |

| | | GOFAG303 (30g, Fire Assay, Gravimetric finish) | GOICP13B (Ore grade, Aqua Regia/ICP-AES) |

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Program Year	Laboratory	Au Lab Package	Ag /Multi Element Lab Package

| | | 202-564 (Fire Assay – Au Ore Grade, Gravimetric Finish, 50g Charge) | 201-074 (Aqua Regia - ICP-OES/ICP-MS, 51 elements) |

| | | 202-564 (Fire Assay – Au Ore Grade, Gravimetric Finish (50g Charge) | 201-061 (Aqua Regia - ICP-OES/ICP-MS - Overlimit) |

11.5.3 First Mining 2020 (Condemnation), 2021 and 2022 Programs

11.5.3.1 Analytical Laboratories

Gold and silver assay work in the 2020 condemnation drill program and the 2022 drill programs was conducted by Activation Laboratories Ltd (ActLabs) at their laboratories in Ancaster, Ontario or Thunder Bay, Ontario. The ActLabs facilities are certified and conform to requirements ISO/IEC 17025:2017. ActLab’s quality system is accredited to international quality standards through the Standard Council of Canada (SCC) and the Canadian Association or laboratory accreditation (CALA). Certification is accredited for precious metals including gold and silver and 52 element geochemical analyses.

First Mining has attested that there is no commercial nor other type of relationship between First Mining and ActLabs that would adversely affect the independence of ActLabs.

11.5.3.2 Analytical Procedures

All samples received by ActLabs were processed through a sample tracking system which utilizes bar coding and scanning technology that provides complete chain of custody records for every stage in the sample preparation and analytical process.

As part of the standard internal quality control procedures used by the laboratory, each batch of samples included blanks, internal standards, and duplicate samples (taken both at the crush and pulp stages). In the event that any reference material or duplicate result fell outside the established control limits, the sample batches would be re-assayed.

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Pulps and rejects from the First Mining core samples from these programs are being kept in storage by First Mining.

11.5.3.3 Gold, Silver and Multi-Element Analysis

Prepared samples were analyzed for gold by fire assay with atomic absorption finish. Samples returning assays in excess of 10 g/t Au were re-analyzed with a gravimetric finish. Samples sent for multi-element analysis were digested with an aqua regia “partial” digestion and then finished using Inductively Coupled Plasma Mass Spectrometry (ICP-MS) or Inductively Coupled Plasma Spectrometry Optical Emission Spectrometry (ICP-OES).

11.5.4 First Mining 2023 and 2024 Programs

11.5.4.1 Analytical Laboratories

All gold and silver assay work from the 2023 and 2024 drill programs were performed by AGAT Laboratories Ltd (AGAT) at their Thunder Bay, Ontario laboratory (for Au fire assay) and Mississauga, Ontario or Calgary, Alberta (for ICP). The AGAT facilities are certified and conform to requirements ISO/IEC 17025:2017. AGAT quality system is accredited to international quality standards through the Standard Council of Canada (SCC) and the Canadian Association or laboratory accreditation (CALA). Certification is accredited for precious metals including gold and silver and 52 element geochemical analyses.

First Mining has attested that there is no commercial nor other type of relationship between First Mining and AGAT that would adversely affect the independence of AGAT.

11.5.4.2 Analytical Procedures

All samples received by AGAT Thunder Bay were processed through a sample tracking system that utilizes bar coding and scanning technology that provides complete chain of custody records for every stage in the sample preparation and analytical process.

Samples were dried and then crushed to 70% of the sample passing 2 mm (-70 mesh). A 250 g sample was split off the crushed material and pulverized to 95% passing (150 mesh). A 50 g split of the pulp was used for gold fire assay and a 2 g split was used for silver analysis. Crushing and pulverizing equipment was cleaned with barren wash material between sample preparation batches and, where necessary, between highly mineralized samples. Sample preparation stations were also equipped with dust extraction systems to reduce the risk of sample contamination. Once the gold assay was complete, a pulp was sent to the AGAT Calgary facility for silver and possibly for multi-element geochemical analysis.

During 2023, two holes from the 2020 hydrogeological program (SGH20-006 and SGH20-009) that intersect the pit area but were not assayed as part of that program, were sampled and sent for fire assay and ICP analysis at AGAT laboratories. Assay methodology was the same as for the 2024 drill program and is detailed below.

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Pulps and rejects from the 2023 and 2024 assay programs are being kept in storage by First Mining.

11.5.4.3 Gold, Silver and Multi-Element Analysis

Prepared samples were analyzed for gold by fire assay with atomic absorption finish. Samples returning assays in excess of 10 g/t Au were re-analyzed with a gravimetric finish. Samples sent for multi-element analysis were digested using aqua regia and then finished using Inductively Coupled Plasma-Atomic Emission Spectrometry (ICP-AES) or Inductively Coupled Plasma Mass Spectrometry (ICP-MS). Silver values over 100 g/t underwent overlimit analysis using an Inductively Coupled Plasma Optical Emission Spectrometry (ICP-OES) finish.

11.6 Bulk Density Data

Bulk density was obtained for select core samples using the water immersion method. The bulk density of a sample is the weight of the sample divided by the volume of the sample including voids.

In the 2012 Gold Canyon program, specific gravity pycnometer measurements (method PHY03V) were completed by SGS Lakefield Research Ltd. laboratory in Lakefield, Ontario on 145 pulverized samples.

The majority of the Project’s bulk density readings (472 samples) were completed by SGS at their Red Lake lab in 2020 as part of a historical core sampling program which was focused on sulphur data collection. Two lab methods were employed by SGS; GPHY17V, a bulk density immersion method where the core sample has a wax coating applied (used on incompetent or crumbly samples) and GPHY04V which uses the same bulk density, immersion method but the core sample is unwaxed. As a general rule in the 2020 bulk density collection program, core samples deemed competent enough not to require a wax coating were tested using the unwaxed method. Of the 472 total samples tested, 393 were waxed and 79 were unwaxed.

The procedure as applied by SGS laboratory for GHPY17V was as follows:

·	Oven-dry the samples and then cool to room temperature
·	Label and weigh each sample in grams
·	Coat the sample with paraffin wax heated in a container immersed in boiling water
·	Repeatedly immerse the sample in the wax until completely sealed
·	Avoid heating the sample
·	Weigh the waxed sample and record
·	Weigh the waxed samples (g) by suspending in water and recording the displaced volume (mL) and the water temperature (°C)
·	Remove the wax by placing in boiling water or freezing the core and chipping off if return of the sample is required

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Calculations:

·	Weight of wax = (weight of sample + wax) – (weight of sample)
·	Volume of wax = weight of wax /specific gravity (SG) of wax corrected for temperature
·	Volume of sample = (volume of sample + wax) – (volume of wax)
·	Bulk density (t/m^3^) = weight of sample (g)/volume of sample (mL)
·	Bulk Density (lb/ft^3^) = (t/m^3^)/0.0160

The remaining 156 samples in the bulk density dataset were taken on site by First Mining field personnel during the 2024 drill program using the wet immersion method. A check program on seven of the 2024 core samples was completed in 2025 by AGAT Laboratories in Thunder Bay, Ontario (see Section 11.2 for further details). AGAT’s wet immersion methodology is to dry the sample at 105°C and then cool to room temperature. It is weighed in air, then coated with a spray lacquer to close the pore space and air dried. The sprayed sample is weighed in water.

Results from selected analysis of bulk density are discussed in Section 14 of the Report.

11.7 Quality Assurance and Quality Control Programs

11.7.1 Historical Programs (Gold Canyon)

11.7.1.1 Pre-2007 QA/QC Programs

No documentation relating to sample handling and preparation or sample QA/QC documentation for the pre-2003 drilling was provided to the QP.

A total of 1,725 database entries were checked against the original certificates. Only a few data entry errors were observed and corrected; however, the total number of errors was not reported.

The QA/QC program for 2003 to 2007 consisted of:

·	resubmission of approximately 10% of the sample pulps to a second laboratory (ALS Chemex)
·	insertion of two commercial standard reference materials (standards submitted every 30th sample)
·	insertion of blanks

There were no field or coarse reject duplicates submitted. Also, no pulp duplicates were submitted to the primary laboratory.

Due to the lack of detailed documentation, particularly for pre-2003 drilling, in 2013, Gold Canyon decided to initiate a core resampling program focusing on the pre-2003 drilling. Specifically, Gold Canyon carried out silver assays for sections that were previously not assayed for silver, expanded the assay intervals to include previously unsampled intervals in the older core and carried out an extensive re-sampling of the previously sampled core to improve on the quality control procedures. A total of 3,352 samples were collected for assays, these include 2,768 check assays for gold, 2,179 check assays for silver, 457 new gold assays and 1,173 new silver assays. The QP reviewed the results of the re-sampling program and concluded that the results did not identify any bias with the pre-2003 drilling and that the pre-2003 drilling was acceptable for inclusion in the resource estimate. The East Extension and Camp zones as now defined correspond to the deposits estimated by P&E in their 2006 study (Armstrong, 2006).

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11.7.1.2 2007/2008 QA/QC Programs

A total of 18 drill holes were completed in 2007 and 2008 comprising a total of 1,374 assay intervals. These samples were assayed for gold only by the Accurassay Laboratories of Thunder Bay, Ontario. The QP checked a total of 137 samples representing 10% of the total against the original certificates. No errors were found.

No program was set up for duplicates, standards, or blanks for this drilling program. The laboratory ran their own set of duplicates for internal monitoring purposes; however, that data was not available to the QP. Accurassay Laboratories was an independent laboratory. The QP is unaware if the laboratory had a specific accreditation in 2007.

11.7.1.3 2010 to 2013 QA/QC Programs

In 2010, Gold Canyon instituted a QA/QC program consisting of commercial standard reference materials for gold, and, consistent with current industry practice, blanks, field duplicates, and pulp duplicates. In addition, a “round robin” program was instituted in 2011 with ACT Labs of Red Lake, Ontario, that compared pulp re-assay results against the original SGS results for 469 samples. ACT Labs was an independent laboratory. The QP is unaware if the laboratory had a specific accreditation in 2007.

SGS conducted their own program of internal duplicate analysis as well. The results of this program were also analyzed by the QP as a valuable comparison against the “blind” pulp duplicates submitted (SRK, 2019).

A summary of the blanks and standards submissions from the 2010 to 2013 assay programs are presented below:

·	A total of 1,605 field duplicates were submitted for gold.
·	A total of 1,532 field duplicates were submitted for silver.
·	A total of 1,591 pulp duplicates were submitted for gold.
·	A total of 1,515 pulp duplicates were submitted for silver.
·	A total of 1,173 commercial gold standards were submitted from a set of 19 different commercial standards.
·	No commercial standards were submitted for silver.
·	A total of 1,647 blanks were submitted with the gold assays.
·	A total of 660 blanks were submitted with the silver assays.

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The total submissions for gold duplicates, standards and blanks were 5,387; representing 10.1% of the samples assayed for gold. The total submissions for silver duplicates and blanks were 3,667; representing 7% of the total samples assayed for silver.

11.7.2 First Mining Programs

11.7.2.1 2016 - 2024 QA/QC Programs

For the First Mining QA/QC programs, blanks and standards were inserted at an approximate rate of one standard for every 25 samples (5% of total), and one blank for every 30 samples (4% of total). ‘Coarse’ duplicates taken from coarse reject, and ‘pulp’ duplicates taken from 250 g pulverized splits, were also inserted at regular intervals with an insertion rate of 4%. For the 2018 to 2024 assay programs, field duplicates from quartered core were also inserted at regular intervals, with an insertion rate of 4%.

In addition to the QA/QC program implemented by First Mining, the laboratories operate their own internal laboratory QA/QC system, inserting quality control materials, blanks, laboratory replicates and laboratory duplicates on each analytical run.

Blanks made of barren garden rock purchased from a local hardware store were used. A threshold of ten times the lower detection limit (LDL) was used as a guide to determine potential contamination. Any assays above this threshold were reviewed on a case-by-case basis to determine if any corrective action was required at that laboratory. If a single blank or standard within a zone of mineralization was deemed to have failed, that QA/QC sample plus five samples either side in the same batch were sent for reanalysis. If a blank/standard plus one or more consecutive standards were deemed to have failed, then the failed samples plus ten samples either side and all the samples between, were sent for reanalysis. For samples from unmineralized zones, if a single standard failed within a batch where the other standards or blanks passed, the entire batch was deemed to have passed and no corrective action was taken.

Twenty six different gold standards were used throughout the First Mining drill programs, spanning a range of gold grades from 0.6 ppm to 15.62 ppm. Eight of the standards were also used as silver standards in these assay programs. The standards were supplied by CDN Resource Laboratories Ltd. (CDN) of Vancouver, BC. A standard was deemed a failure if the result fell outside 3 standard deviations from its expected value as defined by the standard’s certificate. Any assays outside of this threshold were reviewed on a case-by-case basis to determine if any corrective action was required.

A summary of the blanks, standards, and duplicate submissions from the First Mining assay programs are presented below.

11.7.2.2 2016 QA/QC Program:

The total submissions for gold duplicates, standards and blanks in the 2016 assay program were 246, representing 15.7% of the samples assayed for gold. Coarse and pulp duplicates were taken at the rate of one in approximately every 25 samples, alternating between coarse and pulp within the sample stream.

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·	A total of 162 duplicates were submitted for gold, comprising 81 coarse duplicates and 81 pulp duplicates.
·	No duplicates were submitted for silver.
·	No field duplicates were taken in the 2016 program.
·	A total of 41 commercial gold standards were submitted from a set of 6 different commercial standards.
·	A total of 43 blanks were submitted with the gold assays.
·	No commercial standards or blanks were submitted for silver.

Overall, the laboratory performed well during the 2016 assay program.

11.7.2.3 2018 QA/QC Program

The total submissions for gold duplicates, standards and blanks in the 2018 assay program were eight, representing 9% of the samples assayed for gold. Coarse and pulp duplicates were taken at the rate of one in approximately every 25 samples, alternating between coarse and pulp within the sample stream.

·	A total of 12 duplicates were submitted for gold, comprising 4 coarse duplicates, 3 pulp duplicates and 5 field duplicates.
·	No duplicates were submitted for silver.
·	A total of 5 commercial gold standards were submitted from a set of 2 different commercial standards.
·	A total of 3 blanks were submitted with the gold assays.
·	No commercial standards or blanks were submitted for silver.

Overall, the laboratory performed well during the 2018 assay program.

11.7.2.4 2020-2021 QA/QC Program:

The total submissions of duplicates, standards reference material (SRM) and blanks in the 2020-2021 assay program were 755 samples, representing 10.3% of the samples assayed.

Duplicates were positioned between standards and/or blanks, alternating between field duplicates and coarse or pulp duplicates within the sample stream.

·	A total of 246 coarse and pulp duplicates were submitted for both gold and silver.
·	A total of 267 commercial standards for gold and silver were submitted.
·	A total of 242 blanks were submitted with both the gold and silver assays.

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Overall, the laboratory performed well with an average failure rate for the SRM of about 4%. All batches that were deemed failed were re-assayed and the new corrected data was used in the database for the resource estimate.

11.7.2.5 2022-2024 QA/QC Program

The total submissions for gold duplicates, standards and blanks in the 2022-2024 assay programs were 1,073 samples, representing 12.4% of the samples assayed.

In 2022, duplicates were positioned between standards and/or blanks, alternating between field duplicates and coarse or pulp duplicates within the sample stream. In 2024, the duplicates alternated between all three types (field, crush, and pulp) in equal amounts.

·	A total of 269 coarse and pulp duplicates were submitted for both gold and silver.
·	A total of 513 commercial standards for gold and silver were submitted.
·	A total of 291 blanks were submitted with both the gold and silver assays.

As part of the 2022 QA/QC program, First Mining noticed that some of the SRM for silver were consistently reporting higher than the expected value (Figure 11‑1).

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Figure 11‑1: Shewhart Control Chart for 2022 Silver Standards

Source: First Mining, 2022

This issue affected a total of 5,292 core samples over 45 workorders assayed between May 26 and November 16, 2022. Of these 5,292 samples, 464 were duplicate samples and 65 were reruns. Additionally, there were 459 standards and blanks – 451 original standards and 8 rerun standards. Of the 459 standards, 166 were certified for Ag.

To further investigate this issue, and as part of their 2022 check assay program, First Mining sent 21 consecutive samples for ICP reanalysis at AGAT laboratories in Thunder Bay. The results from the AGAT assays appeared to confirm the apparent bias for silver (Figure 11‑2).

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Figure 11‑2: XY Plot of Original ActLabs Silver Assays against AGAT Silver Re-assay Values

Source: First Mining, 2022

An internal investigation by ActLabs determined that the root cause of the high Ag bias noted in the 2022 assay results was related to an inaccuracy at their end caused by Ag spiked samples which were used to re-slope their calibrations, and that the bias was not consistent throughout the data set.

The lab subsequently reprocessed and reissued the 2022 assay certificates to First Mining who reviewed these assay results and found that the reprocessed results contained a greater bias than the original results. ActLabs was not able to express numerically the correction that was applied to the certificates and has not provided their methodology for calculating those results.

Since the Ag bias was not resolved by the re-issuance of the assay certificates, First Mining is unable to use this reprocessed data in their assay database.

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To quantify the potential error that can be attributed to the possible overestimation of silver assays, the QP carried out an investigation of the possible affected assay results (Table 11‑3).

Table 11‑3: Summary of Assays with Possible Bias from 2022 Drilling

Description	Count	Percent of Total

| Silver assays in Springpole database | 82,797 | 100 |

| Silver assays in Springpole Resource Domains | 39,941 | 48.24 |

| 2022 assays with possible high bias | 4,763 | 5.75 |

| 2022 assays with possible high bias used in Resource Domains | 1,368 | 3.43 |

| 2022 assays with possible high bias used in Resource Domains > 1 g/t | 876 | 2.19 |

| 2022 assays with possible high bias used in Resource Domains > 5 g/t | 204 | 0.51 |

Because the majority of the samples with possible high Ag bias are mostly low-grade, and only 204 assay values (0.51% of all assays used for the silver estimate) are greater than 5 g/t, the QP is of the opinion that the inclusion of the 2022 silver assay data will not materially affect the results of the mineral resource estimate. The relatively small amount of data with possible high bias will have a negligible impact on the silver estimate.

The QP is of the opinion that the 2022 silver data can be included in the 2025 mineral resource estimate.

11.8 QP Comments

In the opinion of QP, the sampling preparation, security and analytical procedures used in the drill programs conducted by Gold Canyon for gold analyses are acceptable but not fully consistent with generally accepted industry best practices because of the lack of standard reference material for silver for the earlier drill campaigns. However, because of the relatively low economic value of silver, the QP concludes that the assay data is adequate for use in resource estimation. First Mining has an established QA/QC protocol for the acceptance of assay batches with respect to the performance of standard reference material, duplicates, and blanks. The sampling preparation, security and analytical procedures used by First Mining in all their drilling programs are acceptable for inclusion in a PFS and are consistent with generally accepted industry best practices.

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12 DATA VERIFICATION

12.1 Data Verification by Dr. Arseneau (QP)

Of the 18 drill holes completed in 2007 and 2008, comprising a total of 1,374 assay intervals analyzed for gold, the QP checked a total of 137 samples representing 10% of the total against the original certificates. No errors were found.

A total of 3,135 assay values for gold and 3,161 assay values for silver in the database were compared against the original protected PDF assay certificates submitted by SGS Red Lake. These totals represent 10.1% and 10.4% of the total number of assays for gold and silver, respectively.

Of the original assay values checked against certificates, the focus was on values material to any resource estimate, either higher-grade intervals or very low-grade intervals in proximity to higher-grade intervals. The average grade of gold samples verified was 2.05 g/t Au. The average grade of silver samples checked was 8.27 g/t silver.

Only two errors were found for gold:

·	The gold value of sample interval SP10-028 from 433 m to 436 m (sample number 8287) was found to have an entered value of 5.96 g/t Au against a value on the assay certificate of 9.00 g/t Au.
·	The gold value of sample interval SP11-076 from 69 to 70 m (sample number 14583) having the value of 0.45 oz/t was incorrectly placed in the parts per billion column.

No errors were found with respect to silver assays.

This represents an error rate of 0.064% in gold assays and an error rate of 0.0% in silver assays. This error rate is well within acceptable industry standards.

In addition, a total of 15,965 assays collected during the 2020 to 2024 drill programs were checked against original assay certificates and no errors were noted.

12.1.1 Verification performed by the QP

12.1.1.1 Site Inspection

Dr. Arseneau carried out visits to the Springpole site on February 10 and 11, 2012, on August 8 and 9, 2012 and on June 20 to 22, 2022. During the site visits, core logging procedures were reviewed. Several sections of core from the Portage, Camp, and East Extension zones were examined. Sampling procedures and handling were observed. The deposit geology, alteration, and core recovery data were observed for the Portage zone. The QP was fully assisted during the site visits by Springpole personnel and was given full access to data during the site visits. Springpole field personnel were very helpful and fully cooperative during all site visits.

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During the site visits, Dr Arseneau re-logged mineralized sections of drill core from the Springpole deposit and checked geological units against the recorded written logs. Down-hole survey data entered in the digital database was checked against data entered on paper logs at the site and no errors were noted. Drill site locations could not be verified as most drill sites are situated under Springpole Lake, but two drill platforms drilling on Springpole Lake were observed during one visit.

12.1.1.1.1 Verifications of Analytical Quality Control Data

As part of the mineral resource estimation process, Dr. Arseneau reviewed the QA/QC data collected by Gold Canyon, reviewed the procedures in place to assure assay data quality, and verified the assay database against original assay certificates provided directly to the QP by SGS Red Lake, the assay laboratory. A total of 53,431 gold assays, 46% of the assay data, were checked against original assay certificates. No significant database errors were identified. About 143 minor rounding errors were observed. None of the rounding errors are deemed material or of any significance to the mineral resource estimate presented in this report.

In addition, all of the assay data collected during the 2020 to 2024 drill programs were checked against the original assay certificates for both gold and silver. The QP also reviewed the QA/QC data collected during these programs and found that the procedures established by First Mining and the results of the QA/QC data were in keeping with CIM Best Practice Guidelines (CIM 2018, CIM 2019) and acceptable for inclusions in the estimation of mineral resources and mineral reserves.

12.1.1.2 Independent Verification Sampling

A total of three mineralized quarter core samples were collected during the February 2012 site visit and five additional samples were collected during the 2022 site visit. The intent of the sampling program was only to determine if gold did occur in concentrations similar to what had been reported by Gold Canyon. Assays from the samples collected by Dr. Arseneau are presented in Table 12‑1. The re-sampling agrees with the original Gold Canyon sampling.

Table 12‑1: Assays from Duplicated Samples Collected During Site Visit

QP Check Assay		Gold Canyon Original

| 9135 | 8.64 | 9135 | 9.04 |

| 9136 | 7.49 | 9136 | 7.85 |

| 6152 | 2.37 | 6152 | 2.77 |

| 12618 | 0.15 | 308326 | 0.16 |

| 12619 | 0.66 | 308341 | 0.70 |

| 12620 | 0.58 | 308380 | 0.96 |

| 12621 | 1.10 | 98843 | 0.49 |

| 12622 | 1.60 | 98903 | 1.75 |

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12.2 Data Verification by David Ernest Bleiker (QP)

12.2.1 Geochemistry

The QP read the relevant geochemical appendix included in the Environmental Impact Statement/Environmental Assessment Summary for the Project prepared for First Mining by WSP (October, 2024). This appendix was prepared by WSP technical matter experts. Through discussion with the WSP subject matter experts and the contents of the reports, subject matter experts confirmed that standard industry methods were used for analyses and interpretations that are suitable for Pre-Feasibility studies and for the purpose used in this technical report. Additional discussions were related to interpreting and incorporating the geochemistry results into the CDF Pre-Feasibility designs presented in this document.

12.2.2 Hydrology

The QP read the relevant hydrology appendix included in the Environmental Impact Statement/Environmental Assessment Summary English for the Springpole Gold Project prepared for First Mining by WSP (October, 2024). This appendix was prepared by WSP subject matter experts. Through discussion with the WSP subject matter experts and the contents of the reports, subject matter experts confirmed that standard industry methods were used for the evaluations that are suitable for Pre-Feasibility studies and for the purpose used in this technical report. Additional discussions were related to interpreting and incorporating the hydrology results into the CDF Pre-Feasibility designs presented in this document.

12.3 Verification by Tommaso Roberto Raponi

The QP visited the property on July 29, 2025. During the site visit, inspection of the site layout for locations of planned plant and site infrastructure, available drill core, the core logging and core sampling facilities was completed.

Metallurgical test data was verified through a review of previous studies and testwork reports. Metallurgical testing has been completed at several specialist laboratories. Each laboratory has their own QA/QC procedures, which they adhere to in performing their testing on samples. All metallurgical data was verified and is adequate for this technical report as required by NI 43-101 guidelines.

There have been no limitations on the QP on his verification of any of the data presented in this report. The QP’s opinion is that all data presented in this report is adequate for the purposes of this report and is presented so that it is not misleading.

Inputs into operating costs were obtained through recent quotations or available data from other projects. Operations labour costs were provided by First Mining and reviewed to ensure their currency.

12.3.1 QP’s Comments

In the opinion of the QP, the integrity of the sample data for the Springpole Gold project is adequate for inclusion in mineral resource estimation and for the purpose that it is used in this technical report.

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13 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 Introduction

The Springpole deposit has been the subject of several metallurgical testwork programs and previous studies, as summarized in Section 13.2.

For the 2025 PFS, the flowsheet featured split tailings streams for flotation concentrate and flotation tailings to manage sulphur and potential acid generating (PAG) rock. A Merrill Crowe circuit was selected for precious metal recovery from the concentrate leach solution because of the high silver to gold ratio and variable flotation and leach extraction.

The 2021-2023 testwork program focused on optimizing the flotation and leach process flowsheet including:

·	Primary grind optimization
·	Conventional flotation – reagent optimization
·	Concentrate regrind size optimization and cyanide leach optimization
·	Variability testing for comminution, flotation, leaching and flotation tailings solid-liquid separation
·	Optimization of cyanide detoxification on concentrate and tailings streams
·	Solid-liquid separation tests on concentrate and tailings streams
·	Generation of samples for tailings characterization
·	Merrill Crowe feed solution characterization
·	Materials handling characterization for ore and tailings
·	Evaluation of variability sample metallurgical response to the selected flowsheet

13.2 Metallurgical Testwork

13.2.1 Historical Metallurgical Testwork Programs

A summary of historical testwork campaigns is presented in Table 13‑1. Testwork programs completed prior to 2020 were not considered for process design and are not described in this technical report.

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Table 13‑1: Metallurgical Testwork Summary

Year	Laboratory/Location	Testwork Performed

| 1989 | Lakefield Research, Lakefield, ON; Project No. LR3657 | Whole ore leach cyanide leach and carbon in leach (CIL) |

| 2011 | SGS Mineral Services, Vancouver, BC; Project No. 50138-001 | Whole ore cyanide leach |

| 2013 | SGS Mineral Services, Lakefield, ON; Project No. 13152-001 | Whole ore cyanide leach Flotation and concentrate regrind leach |

| 2013 | Process Mineralogical Consulting Ltd. Maple Ridge, BC; Project No. 201305 | Mineralogical analysis of two grab samples |

| 2017 | Base Metallurgical Laboratories, Kamloops, BC; BL0161 | Comminution testing Mineralogical assessment – BMA, TMS Whole ore leach Rougher flotation and concentrate regrind leach Viscosity |

| 2018 | ALS Metallurgy, Kamloops, BC; 180107 | Whole ore cyanide leach Flotation; Concentrate regrind leach and tail leach |

| 2018 | Jacobs Engineering Group, Lakeland Florida | Reverse flotation to float off mid-size mica to reduce comminution requirements |

| 2018 | Eriez Flotation Division, Erie Pennsylvania | Hydraulic classification to remove multiple size fractions of micas to reduce comminution requirement – cross flow and hydrofloat separation |

| 2020 | SGS Minerals, Lakefield, ON | Project No. 17802-01 November 27, 2020 Phase 1 (4 zone composites) Feed characterization Flotation with concentrate regrind and leach and tailings leach Whole ore leach Cyanide detoxification Solid-liquid separation for tailings |

| | | Project No. 17802-02 December 17, 2020 Phase 2 (4 zone and 12 variability composites) Feed assays Comminution characteristics Flotation with concentrate regrind and leach and tailings leach |

13.2.1.1 2020 SGS Program

The 2020 SGS testwork program was split into two phases:

·	Phase 1 – metallurgical testwork using coarse rejects sample from the 2016 drilling campaign.
·	Phase 2 – comminution and metallurgical testwork using new HQ drill core from the 2020 drilling campaign.

Four zone composites were selected based on spatial distribution, head grade and lithology for the Phase 1 metallurgical testwork campaign (Comp 1 to 4). For Phase 2, four more zone composites (Comp 5 to 8) were selected based on spatial zone, head grade and lithology, and 12 variability samples (Variability 1 to 12) were selected based on discrete intervals according to lithology and head grade. Test results were reported under project number 17802-01 and 17802-02.

The location of the selected drill hole intervals for the 2020 SGS programs are shown in Figure 13‑1 relative to the final pit shell outline.

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Figure 13‑1: Drill Holes Used for 2020 SGS Metallurgical Testing Programs

Source: First Mining, 2021

13.2.2 Current Metallurgical Testing Programs

A summary of current testwork campaigns is presented in Table 13‑2.

Table 13‑2: Current Metallurgical Testwork Summary

Year	Laboratory/Location	Testwork Performed

| 2021 | Base Metallurgical Laboratories, Kamloops, BC; BL758 April 18, 2022 | Comminution testing Mineralogical assessment – BMA, gold deportment Gravity concentration Rougher flotation and concentrate regrind leach Flotation tailings leach Merrill-Crowe precipitation testing Oxygen uptake testing Cyanide detox testing Thickening and filtration testing |

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Year	Laboratory/Location	Testwork Performed

| 2022 | FLSmidth Inc. Salt Lake City, UT; Oracle Project Reference 9252312729 January 7, 2022 | Flotation tailings thickening and filtration testing |

| 2022 | Woodgrove Technologies, Toronto, ON; Report No. 21075-00013319-PR-RPT-10000 March 2022 | DFR pilot plant testing using proprietary Direct Flotation Reactor (DFR) technology for rougher flotation and Benchtop Flotation Reactor (BFR) technology for cleaner flotation |

| 2022 | Jenike and Johanson Ltd. Mississauga, ON; Flow Property Test report 72117-1 and 72117-2 25 March 2025 | Flow property test results for ore and tailings |

| 2023 | Base Metallurgical Laboratories, Kamloops, BC; BL1073 May 4, 2023 | Rougher flotation and concentrate regrind leach Flotation tailings leach Cyanide detox testing Thickening and filtration testing |

13.2.2.1 2021-2023 Base Metallurgical Laboratories Programs

The 2021-2023 Base Metallurgical Laboratories (BaseMet) testwork programs used HQ drill core from the 2021 drilling campaign. Samples were sent to Base Metallurgical Laboratory in Kamloops, Canada, for testing and creating composites. Testing was carried out under two programs, BL758 and BL1073.

As part of the 2021-2023 testwork programs, ten variability samples were tested for ore competency and hardness/grindability characteristics. The ten comminution samples were chosen to represent different zones and lithologies/rock groups within the pit.

In program BL758, three master composites were used to investigate flowsheet optimization for primary grind size and flotation reagents. The Production Composite was made using samples from all five zones to be representative of feed throughout the life of the mine. Master Composite 1 was made up of samples from only zones A, B, and C. Master Composite 2 was comprised of samples from Zones D and E. Mass and metal recovery were investigated for variability samples (spatial, lithology and grade variability samples). Three variability samples were also created to investigate the effects of areas containing high levels of copper, zinc and arsenic in the deposit.

In Program BL1073, four additional master composites were used for Merrill Crowe precipitation investigations and leach and cyanide detoxification optimization, Master Composites 3 and 5 were made up of samples from Zones A, B and C. Master Composites 4 and 6 were comprised of samples from Zones D and E. Six variability composites were assembled for tailings characterization testing. The location of the zones within the 2021 PFS pit shell are presented below in Figure 13‑2.

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Figure 13‑2: Spatial Zones for 2022 and 2023 BaseMet Metallurgical Testing Programs

Source: First Mining, 2021

The locations of the drill holes used for the 2021-2023 BaseMet program are shown Figure 13‑3.

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Figure 13‑3: Drill Hole and Interval Locations for Samples in the 2021-2023 BaseMet Program

Source: First Mining, 2022

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13.2.2.2 2022 FLS Dewatering Program

FLSmidth Inc. (FLS) conducted characterization, thickening, and filtration testing on a tailings sample from the Springpole Project. The sample tested was the Production Composite from BaseMet program BL758. The testing was conducted at the FLS Dawson Laboratories in Salt Lake City, UT.

13.2.2.3 2022 Woodgrove DFR Pilot Program

Woodgrove Technologies performed pilot scale testing on a feed sample from the Springpole Project. The sample tested was the Production Composite from BaseMet program BL758. The testing was conducted at Base Metallurgical Laboratory in Kamloops, BC under the supervision of Woodgrove Technologies staff.

13.2.2.4 2022 J&J Materials Flow Testing Program

Jenike and Johanson Ltd (J&J) performed material flow property testing on seven samples from the Springpole Project. The feed samples tested were six of the comminution samples representing different zones from BaseMet program BL758. The tailings tested was detoxified leach tailings generated from the Production Composite from BaseMet program BL758.

13.3 Metallurgical Testing Programs

13.3.1 Sample Head Analyses

13.3.1.1 2020 SGS Program

Observations from the head assay results for both Phase 1 and 2 composites and Phase 2 variability samples:

·	Gold assays ranged from 0.60 to 2.0 g/t.
·	Silver assays ranged from <0.5 to 20 g/t.
·	All samples assayed low levels of copper and nickel, with low to mid levels of lead, which contribute to cyanide consumption. Composite 8 was high in zinc and lead (0.7% and 0.3% respectively), Var 9 was high in zinc (0.2%).
·	Most sulphur occurs as sulphide sulphur. Sulphide sulphur assays ranged from 0.14% to 5.5%.
·	All samples had low levels of organic carbon indicating low potential of preg-robbing.
·	Composite sample 8 showed high mercury content 8 g/t; variability samples 3, 6, 8 and 9 showed high mercury content ranging 1.2 to 5.7 g/t the remaining samples contained around 0.5 g/t. These levels warrant mercury control in the process flowsheet.
·	Tellurium assays ranged <4 to 18 g/t. This is consistent with the mineralogy of Springpole telluride ore. The presence of tellurides causes refractory behaviour in processing, this is overcome by fine grinding of flotation concentrates for improved liberation of fine-grained gold.

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13.3.1.2 2021-2023 BaseMet Programs

Composites were submitted for a full suite of assays which included gold, silver, copper, arsenic and zinc by direct fire assay, sulphur by Leco analysis and Inductively Coupled Plasma (ICP) scan for 39 elements.

Key assays for the composites tested in programs BL758 and BL1073 are shown below in Table 13‑3 and Table 13‑4.

Table 13‑3: Head Assays for Samples Tested in Program BL758

Composite	Au (g/t)	Ag (g/t)	As (g/t)	Cu (g/t)	Fe (%)	Hg (g/t)	S (%)	Te (g/t)	Zn (g/t)

| Production | 1.27 | 6.96 | 118.9 | 59.2 | 5.56 | 1.21 | 3.17 | 6.45 | 174.2 |

| Master Comp 1 (A+B+C) | 0.96 | 5.95 | 78.5 | 68.6 | 4.19 | 1.03 | 2.60 | 6.95 | 316.9 |

| Master Comp 2 (D+E) | 1.01 | 7.75 | 70.6 | 174.6 | 5.55 | 0.84 | 4.65 | 13.07 | 349.9 |

| Zone A | 1.34 | 1.02 | 126.7 | 31.0 | 2.83 | 0.13 | 0.79 | 0.56 | 105.9 |

| Zone B | 1.15 | 5.67 | 71.9 | 65.0 | 3.12 | 0.45 | 2.58 | 8.24 | 386.9 |

| Zone C | 1.36 | 9.26 | 74.1 | 88.8 | 5.40 | 1.46 | 3.42 | 10.44 | 190.8 |

| Zone D | 1.14 | 6.52 | 113.7 | 101.0 | 5.31 | 0.96 | 4.38 | 8.98 | 279.2 |

| Zone E | 1.03 | 7.12 | 48.7 | 213.7 | 5.73 | 0.64 | 4.71 | 14.38 | 341.1 |

| Grade Bin 1 | 0.44 | 3.59 | 84.6 | 92.5 | 4.73 | 0.82 | 3.33 | 5.52 | 195.4 |

| Grade Bin 2 | 0.66 | 4.83 | 91.1 | 105.7 | 5.18 | 0.93 | 3.75 | 8.51 | 325.2 |

| Grade Bin 3 | 0.98 | 8.55 | 58.9 | 118.2 | 4.99 | 0.85 | 4.10 | 11.90 | 309.6 |

| Grade Bin 4 | 1.70 | 8.41 | 73.9 | 134.4 | 4.88 | 0.76 | 3.65 | 11.57 | 232.9 |

| Grade Bin 5 | 3.15 | 13.86 | 115.4 | 126.2 | 4.84 | 1.28 | 3.66 | 17.42 | 196.4 |

| HG Sample 1 | 1.38 | 3.81 | 99.9 | 140.7 | 5.04 | 0.54 | 3.96 | 7.89 | 202.4 |

| HG Sample 2 | 2.23 | 13.09 | 126.3 | 136.7 | 5.51 | 1.30 | 4.17 | 17.70 | 250.7 |

| LG Sample 1 | 0.50 | 6.87 | 78.2 | 69.7 | 3.80 | 1.09 | 2.79 | 6.78 | 403.1 |

| LG Sample 2 | 0.62 | 4.74 | 89.5 | 85.0 | 4.17 | 0.68 | 3.05 | 6.97 | 179.0 |

| LG Sample 3 | 0.60 | 5.31 | 102.2 | 69.4 | 4.81 | 0.94 | 3.88 | 8.31 | 348.8 |

| HG Breccia | 2.37 | 16.56 | 114.9 | 130.1 | 5.06 | 1.51 | 3.81 | 16.63 | 325.9 |

| LG Breccia | 0.67 | 6.62 | 69.1 | 67.0 | 4.26 | 0.99 | 3.04 | 8.23 | 220.1 |

| HG MS/And/Tuff | 1.99 | 3.85 | 100.4 | 61.4 | 5.09 | 1.04 | 3.48 | 5.05 | 386.7 |

| LG MS/And/Tuff | 0.47 | 4.36 | 68.1 | 68.6 | 5.31 | 0.92 | 2.70 | 4.62 | 164.6 |

| HG Trachyte | 1.46 | 6.78 | 68.4 | 138.0 | 5.28 | 0.49 | 4.67 | 13.66 | 275.1 |

| LG Trachyte | 0.50 | 4.82 | 91.6 | 74.1 | 4.56 | 0.89 | 3.55 | 7.43 | 385.5 |

| High Cu & Zn 1 | 4.66 | 16.03 | 108.1 | 548.5 | 5.66 | 0.73 | 4.71 | 31.24 | 320.6 |

| High Cu & Zn 2 | 1.79 | 20.33 | 61.0 | 164.7 | 5.35 | 1.60 | 3.95 | 25.21 | 779.3 |

| High As | 3.59 | 4.91 | 278.4 | 82.2 | 5.78 | 0.80 | 4.80 | 5.47 | 136.6 |

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Table 13‑4: Head Assays for Samples Tested in Program BL1073

Sample	Head grades (calculated)

| MC3 | 1.04 | 5 | 2.60 | 3.62 | 1.00 | 77 | 458 | 77 |

| MC4 | 0.98 | 7 | 3.67 | 4.65 | 1.00 | 111 | 334 | 187 |

| MC5^1^ | 1.03 | 4 | 2.72 | - | - | - | - | - |

| MC6^1^ | 0.89 | 5 | 3.37 | - | - | - | - | - |

| Zone A^1^ | 1.54 | 0.5 | 0.61 | - | - | - | - | - |

| Zone B^1^ | 1.09 | 1 | 3.22 | - | - | - | - | - |

| Zone C^1^ | 0.69 | 2 | 2.90 | - | - | - | - | - |

| Zone D^1^ | 0.52 | 3 | 3.17 | - | - | - | - | - |

| Zone E^1^ | 0.48 | 6 | 2.52 | - | - | - | - | - |

| High Cu/Zn Comp^1^ | 3.42 | 30 | 5.26 | - | - | - | - | - |

Note:

1.	No head assays were performed for Fe, Hg, Cu, Zn, Pb for these composites.

13.3.2 Sample Mineralogy

13.3.2.1 2020 SGS Program

Bulk mineralogy of Composites 2 and 4 was determined with QEMSCAN, as summarized in Table 13‑5.

Table 13‑5: Mineral Proportions

Mineral Mass %	Comp 2	Comp 4

| Pyrite/marcasite | 5.25 | 7.76 |

| Other sulphides | 0.02 | 0.03 |

| Quartz | 5.53 | 9.09 |

| K-Feldspar | 32.7 | 23.0 |

| Plagioclase | 10.5 | 6.45 |

| Muscovite/Illite | 19.5 | 19.8 |

| Biotite | 20.1 | 29.9 |

| Chlorite | 0.53 | 0.66 |

| Kaolinite | 0.49 | 0.77 |

| Other silicates | 0.10 | 0.06 |

| Fe-oxides | 1.20 | 0.49 |

| Rutile | 1.12 | 1.17 |

| Calcite | 0.09 | 0.06 |

| Ankerite | 0.73 | 0.10 |

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Mineral Mass %	Comp 2	Comp 4

| Siderite | 1.38 | 0.02 |

| Apatite | 0.55 | 0.61 |

| Other | 0.23 | 0.06 |

| Total | 100 | 100 |

Bulk mineral composition of Composite 2 and 4 shows:

·	major potassium feldspar 23-33%;
·	major micas (biotite, muscovite, illite) 40-50%;
·	both composites are characterized by pyrite/marcasite sulphide mineralization; and
·	trace amount of clays (kaolinite) are present.

A full gold deportment study was completed on two composites as shown in Table 13‑6.

Table 13‑6: Summary of Gold Deportment

Sample ID	Association	Average Size (µm)	Size Range (µm)	Overall Gold Distribution (%)

| Comp 2 | Liberated | 3.2 | 0.6 - 12.8 | 21.8 |

| | Exposed | 2.8 | 0.6 - 25.1 | 64.3 |

| | Locked | 2.1 | 0.6 - 11.5 | 8.48 |

| | Sub-microscopic-Au | - | - | 5.37 |

| | Total | 2.7 | 0.5 - 7.4 | 100 |

| Comp 4 | Liberated | 3.9 | 0.6 - 38.3 | 32.0 |

| | Exposed | 2.4 | 0.6 - 25.1 | 42.1 |

| | Locked | 2.2 | 0.6 - 6.7 | 14.1 |

| | Sub-microscopic-Au | - | - | 11.7 |

| | Total | 2.4 | 0.5 - 7.4 | 100 |

The gold deportment study shows, for composites 2 and 4:

·	5 to 12% of the gold is sub-microscopic (<0.6 µm) or refractory
·	8 to 14% of the gold is locked in the 0.6 to 11 µm size range
·	42 to 64% of the gold is exposed
·	22 to 32% of the gold is liberated

The gold distribution by gold minerals by grain size is presented in Table 13‑7.

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Table 13‑7: Summary of Gold Carrier Minerals

Comp 2	Gold Minerals	Gold Distribution (%)	Comp 4	Gold Minerals	Gold Distribution (%)

| | Au-Ag-Pb-Te | 14.4 | | Au-Ag-Pb-Te | 16.6 |

| | Pb-Petzite | 7.69 | | Sylvanite | 11.3 |

| | Sylvanite | 6.27 | | Electrum | 5.58 |

| | Electrum | 5.58 | | Pb-Petzite | 2.01 |

| | Au-Hessite | 4.01 | | Muthmannite | 1.73 |

| | Native Gold | 3.72 | | Native Gold | 1.47 |

| | Au-Ag-Te | 1.04 | | Au-Hessite | 1.00 |

| | Au-Pb-Hessite | 0.89 | | Pb-Sylvanite | 0.75 |

| | Muthmannite | 0.60 | | Calaverite | 0.47 |

| | Calaverite | 0.37 | | Au-Ag-Te | 0.45 |

| | Au-Altaite | 0.28 | | Au-Ag-Hg | 1.74 |

| | Au-Ag-Hg | 0.17 | | - | - |

| | Chalcopyrite | 0.01 | | Chalcopyrite | 0.05 |

| | Hematite | 0.32 | | Hematite | 0.23 |

| Total | | 100 | | | 100 |

Gold occurs primarily as telluride minerals in the microscopic size range, with petzite the most dominant. While not refractory, gold telluride minerals require more aggressive leach conditions. Gold and electrum occur in minor amounts (<10%). Pyrite is the dominant carrier in the sub-microscopic size range.

13.3.2.2 2021-2023 BaseMet Program

Bulk mineralogy of the five zone composites was determined using QEMSCAN as summarized below in Table 13‑8.

Table 13‑8: Bulk Mineral Composition of Zone Composite Samples (Program BL758)

Mineral Mass (%)	Zone A	Zone B	Zone C	Zone D	Zone E

| Pyrite | 1.68 | 5.91 | 7.02 | 9.19 | 10.6 |

| Other Sulphides | 0.07 | 0.07 | 0.17 | 0.08 | 0.13 |

| Quartz | 13.6 | 3.73 | 10.7 | 5.87 | 4.80 |

| Plagioclase | 4.15 | 1.31 | 14.5 | 12.1 | 6.58 |

| K-Feldspar | 29.3 | 70.3 | 24.3 | 39.2 | 41.8 |

| Sericite/Muscovite | 21.6 | 8.51 | 15.3 | 9.41 | 11.0 |

| Biotite | 5.52 | 5.98 | 9.49 | 10.1 | 15.6 |

| Clays | 2.20 | 1.47 | 2.44 | 2.59 | 2.84 |

| Other Silicates | 0.84 | 0.26 | 0.86 | 1.04 | 0.24 |

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Mineral Mass (%)	Zone A	Zone B	Zone C	Zone D	Zone E

| Oxides | 1.12 | 1.55 | 2.50 | 2.28 | 2.09 |

| Calcite | 3.30 | 0.12 | 5.17 | 3.67 | 2.71 |

| Ankerite | 14.5 | 0.12 | 5.68 | 3.21 | 0.18 |

| Other Carbonates | 1.37 | 0.12 | 1.36 | 0.69 | 0.22 |

| Apatite | 0.66 | 0.07 | 0.40 | 0.29 | 0.53 |

| Fluorite | 0.02 | 0.01 | 0.03 | 0.03 | 0.28 |

| Other | 0.06 | 0.46 | 0.17 | 0.25 | 0.31 |

| Total | 100 | 100 | 100 | 100 | 100 |

Pyrite is the predominant sulphide mineral (>98%).

A visible gold deportment study on the five zone composites is summarized below in Table 13‑9.

Table 13‑9: Gold Deportment and Association in Zone Composite Samples (Program BL758)

Gold Association	Zone A	Zone B	Zone C	Zone D	Zone E

| Pure gold minerals | 12.1 | 0.04 | 0.00 | 15.4 | 0.00 |

| Free gold minerals | 28.7 | 0.00 | 0.00 | 0.00 | 0.00 |

| Liberated gold minerals | 11.7 | 0.00 | 5.28 | 45.3 | 0.00 |

| Gold:Silver minerals | 0.00 | 0.09 | 0.01 | 0.00 | 0.03 |

| Gold:Pyrite | 38.7 | 3.84 | 0.72 | 16.0 | 20.2 |

| Gold:Arsenopyrite | 0.00 | 0.00 | 0.00 | 0.00 | 0.00 |

| Gold:Other Sulphides | 0.00 | 0.01 | 0.00 | 0.00 | 0.00 |

| Gold:Silicates | 0.00 | 70.7 | 0.23 | 0.00 | 0.00 |

| Complex | 8.78 | 25.3 | 93.8 | 23.3 | 79.8 |

| Total Liberated | 52.5 | 0.04 | 5.28 | 60.7 | 0.00 |

| Total Associated | 47.5 | 100 | 94.7 | 39.3 | 100 |

Most free gold observed were found in Zone A and Zone D samples both as pure gold grains and free/liberated (<5 and 80% of surface area associations respectively). The balance of associations were with pyrite or complex categories, with Zone B containing 70% of visible gold in association with silicates.

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13.3.3 Comminution Testing

13.3.3.1 2020 SGS Program

13.3.3.1.1 Comminution Characterization

Twenty-seven samples were tested for ore competency and hardness/grindability characteristics. Of these samples, twenty-four composite samples were prepared for the comminution testwork only, the remaining three samples were based on metallurgical variability samples, namely Variability 10 to 12. Results are included in Table 13‑10.

13.3.3.1.2 Regrind Testing - IsaMill

IsaMill is one technology used for ultrafine grinding, which is based on a high-intensity attrition mechanism. The relationship between the product size and energy input remains constant during the scale-up of the bench-scale testing unit. The log of size plotted against the log of energy is referred to as a signature plot to the tested sample, which is used to support the selection of a full-scale IsaMill.

The flotation concentrate sample from test F-39 was submitted for IsaMill testing with a target product size 80% passing size (P80) of 15 µm. Grinding was completed in two stages. Stage 1 used 4.5 mm grinding media with an intermediate target P80 of approximately 30 µm while stage 2 used 2.5 mm grinding media to reach the final target P80. The results are plotted in Figure 13‑4.

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Figure 13‑4: IsaMill Signature Plot

Source: SGS, 2020

The energy required to grind the sample from a feed 80% passing size (F80) of 173 µm to the target product size was measured at 46.9 kWh/t of concentrate.

13.3.3.2 2021-2023 BaseMet Program

Comminution testwork was carried out on ten variability samples to assess variability in ore competency and hardness/grindability characteristics by both zone and lithology.

Due to the crumbly nature of some of the drill core samples, it was not possible to obtain a representative set of core pieces to conduct SMC Test for most of the Springpole samples. Hence, the SAG Power Index (SPI) test was utilized, as this test uses whole ore samples rather than selected size fraction. Ten parallel tests (SMC and SPI) were conducted for validation of the relationship between the results of these two tests. As shown in Figure 13‑5 the parallel test results indicate that the Springpole samples are well aligned with the relationship between SPI and drop weight index (DWi) developed for a large database. The data in Figure 13‑5 includes previous historical data as well as the 2021-2023 comminution variability samples.

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Figure 13‑5: SPI and DWi Relationship (Adapted from Bailey et al. 2008)

Source: Ausenco, 2025

In addition to the ten parallel tests (SMC and SPI), conventional Bond Work Index tests were carried out to determine Bond Rod Mill Work Index (RWi), Bond Ball Work Index (BWi) at a closing screen size of 106 µm, and Bond Abrasion Index (Ai). Table 13‑10 shows the full breakage database, including historical data.

Table 13‑10: Breakage Data

Program	Sample	Rock Group	Ai (g)	RWi (kWh/t)	BWi (kWh/t)	SPI (min)	Axb (calculated)	SG

| SGS 2020 | Comp 1 (PM-DH-01) | Not available | 0.142 | - | 12.7 | - | - | 2.57 |

| | Comp 2 (PM-DH-02) | Not available | 0.080 | - | 13.4 | - | - | 2.64 |

| | Comp 3 (PM-DH-03) | Not available | 0.108 | - | 13.3 | - | - | 2.60 |

| | SP11-061 | Not available | - | - | 7.3 | - | - | - |

| | SP11-065 | Not available | - | - | 13.9 | - | - | - |

| | SP11-066 | Not available | - | - | 17.1 | - | - | - |

| | SP11-069 | Not available | - | - | 11.9 | - | - | - |

| | SP11-090 | Not available | - | - | 13.2 | - | - | - |

| | Tra 1 - Ore A | Not available | - | - | - | 49.6 | 59.0 | 2.70 |

| | Tra 1 - Ore B | I | 0.033 | 10.3 | 15.2 | 19.8 | 111.5 | - |

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Program	Sample	Rock Group	Ai (g)	RWi (kWh/t)	BWi (kWh/t)	SPI (min)	Axb<br> <br>(calculated)	SG

| | Tra 1 - Ore C | I | 0.079 | 10.0 | 12.0 | 14.2 | 140.3 | - |

| | Bx 1 - Ore A | I | - | - | - | 7.2 | 224.6 | - |

| | Bx 2 - Ore A | X | 0.063 | 9.6 | 11.9 | 17.8 | 120.0 | - |

| | Bx 2 - Ore B | X | - | - | 16.2 | 16.3 | 127.5 | - |

| | Tra 2 - Ore A | X | - | 9.1 | 13.3 | 17.1 | 123.4 | 2.46 |

| | Tra 3 - Ore A | I | 0.386 | 15.0 | 16.5 | 67.3 | 47.8 | 2.73 |

| | Por 3 - Ore A | I | 0.141 | 13.1 | 15.4 | 37.6 | 71.5 | - |

| | Por 3 - Ore B | I | 0.172 | - | 17.9 | 31.7 | 80.4 | - |

| | Bx 3 - Ore A | I | 0.033 | - | 11.5 | 22.0 | 103.6 | - |

| | Tra 3 - Ore B | X | 0.025 | 8.9 | 8.4 | 31.3 | 81.2 | - |

| | Tra 3 - Ore C | I | - | 11.8 | 12.8 | 35.7 | 74.1 | 2.70 |

| | Tra 1 - Ore D | I | - | - | 11.0^1^ | - | - | - |

| | Tra 3 - Ore D | I | - | - | 15.6^1^ | - | - | - |

| | Tra 3 - Ore E | I | - | - | 12.0^1^ | - | - | - |

| | Por 3 - Ore C | I | - | - | 16.8^1^ | - | - | - |

| | Bx 2 - Ore C | I | - | - | 12.4^1^ | - | - | - |

| | Var 10 - and DH4 | X | - | - | 13.5^1^ | - | - | - |

| | Var 11 - Msed DH4 | B | - | - | 12.7^1^ | - | - | - |

| | Var 12 - Bx DH4 | S | - | - | 16.4^1^ | - | - | - |

| 2021-2023 Basemet (BL758) | Zone A - Bx 1 | X | 0.234 | 17.1 | 18.3 | 123.8 | 31.3 | - |

| | Zone A - And 1 | B | 0.051 | 16.1 | 13.7 | 116.0 | 32.8 | - |

| | Zone B - Tra 1 | B | 0.075 | 9.3 | 11.9 | 10.8 | 169.7 | - |

| | Zone B - Bx 1 | I | 0.303 | 14.8 | 17.4 | 97.6 | 36.9 | - |

| | Zone C - And 1 | X | 0.357 | 15.7 | 13.5 | 137.2 | 29.2 | - |

| | Zone C - Tra 1 | B | 0.283 | 16.7 | 18.6 | 82.0 | 41.6 | - |

| | Zone D - Tra 1 | I | 0.316 | 17.6 | 17.4 | 143.4 | 28.3 | - |

| | Zone D - Tra 2 | I | 0.043 | 8.6 | 12.7 | 14.2 | 140.1 | - |

| | Zone E - Tra 1 | I | 0.193 | 15.9 | 19.3 | 62.5 | 50.3 | - |

| | Zone E - Tra 2 | I | 0.184 | 13.9 | 16.2 | 41.5 | 66.7 | - |

Note:

1.	Test done on assay reject material and was finer than standard Bond ball mill feed

| 2. | Dashes in the table above for RWI, SPI, Axb and SG values indicate that the relevant parameter was not tested for this sample. |

Where rock groups are defined as follows:

·	B - Basement Suite
·	I - Intrusive Complex
·	X -Volcaniclastic Breccias and Associated Exhalatives
·	S - Overlying Sedimentary Series

Average abrasion index is 0.16 g which is considered low.

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BWi ranges from 7.3 to 19.3 kWh/t which is considered very soft to hard, ore hardness is highly variable. The 75^th^ percentile value of 15.9 kWh/t was selected for design.

Calculated Axb ranges from 28-224 which is considered competent to soft, ore competence is highly variable. The Axb value for design was determined for each rock group (Basement, Intrusive and Volcanic) as the inverse of the 75^th^ percentile value. The overall Axb value for SAG mill design of 65.2 was calculated as the weighted average by rock group over the life of mine.

Figure 13‑6 shows examples of crumbly and competent drill core that would account for the variability in comminution characteristics observed.

Figure 13‑6: Example of Crumbly and Competent Drill Core

Source: Ausenco, 2025

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Ore hardness (BWi) does not vary significantly per rock group, however ore competence (Axb) varies widely by rock group, with average values of Axb for Basement, Intrusive and Volcanic of 35, 87 and 128 respectively. The Basement rock group is considered competent while Intrusive and Volcanic rock groups are soft. For optimal mill design and operation, the proportion of Basement rock group in the mill feed will need to be managed.

13.3.4 Flotation

13.3.4.1 2020 SGS Program

The 2020 SGS test program investigated rougher kinetics in Phase 1 and 2, and a suite of different flotation reagents over a range of primary grind sizes P80 from 100 µm to 200 µm. Drilling compound used to aid in drill core recovery negatively impacted flotation results for some samples, causing foaming and high mass pulls. Optimized conditions selected for subsequent testing included a primary grind P80 size of 150 µm, natural pH, collector addition; 100 g/t to 130 g/t potassium amyl xanthate (PAX) and frother addition; methyl isol butyl carbinol (MIBC) as required.

Using 8 minutes rougher flotation residence time, gold recovery for all rougher flotation tests at P80 150 µm primary grind on Phase 1 composites ranged from 72% to 78%. Despite high sulphur recovery, gold recovery was not sufficiently high, requiring leaching of flotation tailings was required to achieve acceptable overall gold recovery.

Phase 2 samples Composites 5 and 6 showed high sulphur recovery (>95%). However, composite 6 was foaming and resulting in increased mass pull with expected entrainment of non-sulphide gangue. Composites 7 and 8 showed sulphur recovery ranging from 80% to 88%. Gold recovery for composites 5, 6 and 8 was relatively high compared with Phase 1 tests ranging from 78% to 82%, at 8 minutes flotation time. Composite 7, with a low sulphide head grade (0.14 %) and gold head grade (0.8 g/t), showed reduced gold and silver recovery.

13.3.4.2 2021-2023 BaseMet Program

The 2021-2023 BaseMet Program (BL758) investigated the effect of primary grind size on overall flotation and leach extraction. The production composite and two master composites were assessed and the primary grind sizes tested were P80 of 75, 100, 125 and 150 µm. After primary grinding, the mineralized material was concentrated by means of conventional flotation. Flotation concentrate was reground to a P80 of 17 µm before leaching, and flotation tailings were leached without regrinding. Overall gold extraction was used to compare the different primary grind sizes. The results of these tests are summarized in Table 13‑11.

Table 13‑11: Overall Flotation and Leach Extraction at Different Grind Sizes

Composite	Test Number	Grind Size (P80, µm)	Overall Au Extraction (%)	Combined Leach Tailings Grade (g/t)

| Production Comp | BL758-24 | 100 | 88.1 | 0.14 |

| Production Comp | BL758-34 | 125 | 88.2 | 0.13 |

| Production Comp | BL758-35 | 150 | 89.2 | 0.12 |

| MC 1 | BL758-25 | 75 | 92.4 | 0.08 |

| MC 1 | BL758-20 | 100 | 90.3 | 0.10 |

| MC 1 | BL758-36 | 125 | 88.8 | 0.11 |

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Composite	Test Number	Grind Size (P80, µm)	Overall Au Extraction (%)	Combined Leach Tailings Grade (g/t)

| MC 1 | BL758-37 | 150 | 89.0 | 0.12 |

| MC 2 | BL758-26 | 100 | 82.2 | 0.16 |

| MC 2 | BL758-38 | 125 | 83.6 | 0.14 |

| MC 2 | BL758-39 | 150 | 81.1 | 0.16 |

There were small improvements in extraction as grind size decreased for Master Composites 1 and 2, and a slight decrease in extraction with decreasing grind size for the Production Composite. These findings did not support a target primary grind size smaller than a P80 of 150 µm. The difference in extraction was not substantial enough to offset the increased costs of finer grinding.

The 2020 SGS program identified gangue reporting to the flotation concentrate, so dispersants were tested to limit mass pull and prevent floatable mica from appearing the concentrate product. Results are shown in Table 13‑12.

Table 13‑12: Effect of Dispersant Addition on Mass Recovery to Concentrate

Test Number	Composite	Reagent	Reagent Addition (g/t)	Flotation Concentrate Mass Pull (%)	Concentrate Au Distribution (%)	Concentrate Ag Distribution (%)	Concentrate S Distribution (%)

| BL758-26 | MC 2 | N/A | N/A | 22.30 | 80.5 | 79.6 | 96.0 |

| BL758-29 | MC 2 | Disp 40E | 500 | 18.37 | 79.0 | 81.3 | 95.0 |

| BL758-30 | MC 2 | Disp 40E | 250 | 19.45 | 81.7 | 87.7 | 95.0 |

| BL758-31 | MC 2 | Dep CMC | 250 | 20.66 | 81.9 | 70.4 | 96.6 |

The use of slimes dispersants showed little benefit and subsequent tests were performed without dispersants.

Flotation tests were performed on Master Composites 1 and 2 to determine optimal PAX and MIBC dosages. Addition rates of 100 g/t PAX and 14 g/t MIBC were selected and used in subsequent tests.

Cleaner flotation was removed from the flowsheet due to the requirement to maximize sulphur recovery to the flotation concentrate for separate disposal of potentially acid generating (PAG) tailings.

Flotation tests in program BL758 were carried out on the variability composites. Tests were carried out in a 4L Denver cell at a primary grind P80 size of 150 µm, at natural pH. Rougher flotation time was 12 minutes, and PAX addition was 100 g/t, MIBC was added as required. Results are summarized in Table 13‑13 and shown in Figure 13‑7.

Flotation tests in Program BL1073 were carried out at the same conditions as Program BL758 and are summarized in Table 13‑14.

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Figure 13‑7: Distribution to Rougher Flotation Concentrate - Variability Samples (Program BL758)

Source: Ausenco, 2025

Mass and metal recovery to concentrate was highly variable, ranging from 5% to over 30% for mass and 60% to 95% for gold.

First Mining engaged Woodgrove Technologies to perform pilot scale rougher flotation testwork for Springpole using Direct Flotation Reactor (DFR) technology and cleaner testwork using its Benchtop Flotation Reactor (BFR) technology. The highlights of the program were:

·	DFR roughers were able to repeatedly achieve a 93% sulphur recovery to concentrate with a 10% mass pull
·	DFR roughers were able to achieve a maximum sulphur recovery of 95% at a mass pull of 14%
·	BFR cleaners were able to achieve a stage mass recovery of 60% at 97-98% sulphur and 94% gold recovery

These tests demonstrated that DFR and BFR technology was successful at upgrading the feed with high recoveries to concentrate and a low mass recovery. Economic evaluation recommended selection of conventional flotation tank cells for the project.

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13.3.5 Leaching

13.3.5.1 2020 SGS Program

Phase 1 examined concentrate leach extraction over a range of regrind sizes and selected a target P80 grind size of 15 µm to 17 µm for subsequent tests. Selected leach test conditions were 40% solids (by weight), pH of 11-11.5, dissolved oxygen of 20 mg/L (using oxygen) and NaCN concentration maintained at 2 g/L.

Phase 2 flotation concentrate leach results for composites 5, 6 and 8 and variability samples 1 to 12 at 48 h showed gold extraction ranging from 62% to 97% and silver extraction from 80% to 96%. Sodium cyanide consumption ranged 1.3 to 11.6 kg/t of concentrate and lime consumption ranged 2.2 to 7.8 kg/t of concentrate. Similarly, to Phase 1 composites, Phase 2 samples demonstrated most of the leach is complete within 30 h.

Flotation tailings cyanide leach conditions were 40% solids, pH of 11-11.5, dissolved oxygen of 20 mg/L. Leaching of gold was generally complete within 12 h. For Phase 2 composites and variability samples, gold extraction ranged from 51% to 93% and silver extraction from 26% to 90%. Sodium cyanide consumption ranged 0.13 to 0.68 kg/t of tailings and lime consumption ranged 0.6 to 1.3 kg/t of tailings.

13.3.5.2 2021-2023 BaseMet Program

Flotation concentrate produced from the Production Composite and Master Composite 1 were tested to determine optimum regrind size for concentrate leaching. Two regrind sizes were selected for comparison at P80 of 10 versus 17 µm. The results were evaluated on combined gold and silver extractions, and the results are presented in Table 13‑13.

Table 13‑13: Flotation and Leach Extraction Test Results at Different Regrind Sizes (Program BL758)

Test Number	Composite	P80 (µm)	Leach Concentrate Au Extraction (%)	Leach Concentrate Ag Extraction (%)	Combined Leach Au Extraction (%)	Combined Leach Ag Extraction (%)

| BL758-24C | Production Comp | 17 | 88.1 | 97.0 | 88.1 | - |

| BL758-24D | Production Comp | 10 | 90.9 | 97.0 | 90.4 | - |

| BL758-34A | Production Comp | 17 | 89.7 | 92.29 | 88.2 | 87.8 |

| BL758-35A | Production Comp | 17 | 90.2 | 93.3 | 89.2 | 87. 9 |

| BL758-14A | MC 1 | 10 | 90.8 | 96.1 | 91.0 | 96.5 |

| BL758-20A | MC 1 | 10 | 92.9 | 96.4 | 90.3 | 96.3 |

| BL758-25A | MC 1 | 17 | 93.0 | 96.6 | 92.4 | 96.8 |

| BL758-36A | MC 1 | 17 | 89.3 | 93.9 | 88.9 | 88.2 |

| BL758-37A | MC 1 | 17 | 90.2 | 92.4 | 89.0 | 86.4 |

| BL758-14A | MC 1 | 10 | 90.8 | 96.1 | 90.9 | 96.7 |

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Using an Isamill signature plot determined during the 2021 PFS, at a P80 grind size of 17 µm the specific energy is 38.8 kWh/t of concentrate, this increases to 86.6 kWh/t of concentrate at 10 µm. The finer regrind size did not provide a significant enough improvement to recovery to justify the increased of the regrind. A P80 regrind size of 17 µm was selected for subsequent variability testwork.

The products from the flotation variability tests were advanced to concentrate and tailings leaching. This testing was conducted to investigate the effects of zone, head grade, lithology and high copper, zinc and arsenic head grades on concentrate leach residue and overall extraction. All variability tests targeted a concentrate regrind P80 size of 17 µm.

Flotation concentrate leach test conditions were bottle roll tests using 2 g/L NaCN, oxygen sparged, pH 11.0, 20% solids for 48 h. Flotation tailings leach test conditions were bottle roll tests using 1 g/L NaCN, oxygen sparged, pH 11.0, 40% solids for 48 h.

The combined flotation and concentrate and tailings leach performance for the Production, Master and Variability Composites are shown in Figure 13‑8 and Table 13‑14. Leach kinetics for concentrate and tailings are shown in Figure 13‑9 and Figure 13‑10.

Figure 13‑8: Leach Extraction for Flotation Concentrate and Tailings Composites and Variability Samples (Program BL758)

Source: Ausenco, 2025

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Figure 13‑9: Concentrate Leach Gold Extraction vs Cumulative Time (Program BL758)

Source: Ausenco, 2025

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Figure 13‑10: Tailings Leach Gold Extraction vs Cumulative Time (Program BL758)

Source: Ausenco, 2025

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Table 13‑14: Flotation and Leach Extraction for Composites and Variability Samples (Program BL758)

Composite	Test Number	Rougher Flotation Concentrate Distribution			Rougher Concentrate Mass (%)	Flotation Concentrate Leach Extraction		Flotation Tailings Leach Extraction		Overall Flotation Leach Extraction		Calculated<br> <br>Head Grade			Overall<br> <br>Residue

| | | Au (%) | S (%) | Ag (%) | | Au (%) | Ag (%) | Au (%) | Ag (%) | Au (%) | Ag (%) | Au (g/t) | Ag (g/t) | S (%) | Au (g/t) | Ag (g/t) |

| Production Comp | BL758-35 | 81.3 | 76.4 | 94.8 | 15.7 | 90.2 | 93.3 | 84.7 | 70.4 | 89.2 | 87.9 | 1.26 | 8.58 | 2.23 | 0.12 | 1.18 |

| MC 1 | BL758-37 | 71.6 | 70.2 | 94.0 | 11.9 | 90.1 | 92.3 | 86.2 | 72.2 | 89.0 | 86.3 | 1.09 | 8.87 | 2.18 | 0.12 | 1.27 |

| MC 2 | BL758-39 | 84.9 | 87.9 | 96.8 | 26.5 | 79.8 | 93.3 | 88.7 | 86.1 | 81.1 | 92.4 | 0.93 | 10.92 | 5.55 | 0.16 | 1.25 |

| Zone A | BL758-42 | 94.2 | 69.3 | 99.4 | 5.9 | 95.2 | 82.2 | 83.4 | 52.0 | 94.5 | 72.9 | 2.12 | 0.92 | 0.76 | 0.14 | 0.22 |

| Zone B | BL758-43 | 69.8 | 60.6 | 93.6 | 12.2 | 86.8 | 88.7 | 92.0 | 92.2 | 88.4 | 90.1 | 1.16 | 5.57 | 2.47 | 0.13 | 0.55 |

| Zone C | BL758-44 | 71.6 | 88.7 | 93.8 | 13.3 | 88.9 | 93.3 | 86.9 | 76.1 | 88.4 | 91.4 | 1.37 | 9.12 | 2.89 | 0.12 | 0.89 |

| Zone C | BL758-44R^1^ | 80.4 | 79.1 | 95.3 | 19.0 | | | | | | | 1.49 | 8.54 | 3.61 | | |

| Zone D | BL758-45 | 62.0 | 58.9 | 70.9 | 12.2 | 84.9 | 87.2 | 79.0 | 74.3 | 83.5 | 85.2 | 1.08 | 10.18 | 3.38 | 0.21 | 1.23 |

| Zone D | BL758-45R^1^ | 83.6 | 75.4 | 94.7 | 23.8 | | | | | | | 1.16 | 5.26 | 4.22 | | |

| Zone E | BL758-46 | 69.8 | 70.7 | 92.4 | 16.2 | 91.5 | 89.5 | 85.4 | 90.6 | 89.6 | 89.8 | 1.00 | 6.01 | 4.48 | 0.16 | 0.62 |

| Grade Bin-1 | BL758-47 | 72.5 | 60.3 | 92.2 | 12.8 | 78.3 | 85.7 | 82.3 | 84.7 | 79.4 | 85.3 | 0.47 | 3.51 | 2.94 | 0.10 | 0.49 |

| Grade Bin-2 | BL758-48 | 69.9 | 64.7 | 91.3 | 14.0 | 82.2 | 87.6 | 81.1 | 85.9 | 81.9 | 87.0 | 0.63 | 3.89 | 3.59 | 0.11 | 0.52 |

| Grade Bin-3 | BL758-49 | 61.6 | 62.8 | 77.4 | 12.4 | 83.2 | 90.1 | 80.7 | 78.0 | 82.2 | 85.6 | 0.82 | 8.49 | 3.72 | 0.15 | 1.15 |

| Grade Bin-3 | BL758-49R^1^ | 82.3 | 78.1 | 96.4 | 23.0 | | | | | | | 0.92 | 6.32 | 4.19 | | |

| Grade Bin-4 | BL758-50 | 60.2 | 63.4 | 79.9 | 12.4 | 87.7 | 93.3 | 91.7 | 82.1 | 89.3 | 89.2 | 1.58 | 13.39 | 3.55 | 0.16 | 1.32 |

| Grade Bin-4 | BL758-50R^1^ | 75.6 | 73.0 | 92.8 | 19.5 | | | | | | | 1.55 | 11.34 | 4.17 | | |

| Grade Bin-5 | BL758-51 | 60.4 | 56.2 | 79.0 | 11.6 | 89.9 | 93.5 | 84.4 | 86.5 | 87.7 | 90.4 | 2.86 | 13.72 | 3.77 | 0.35 | 1.23 |

| Grade Bin-5 | BL758-51R^1^ | 78.2 | 71.2 | 94.9 | 17.8 | | | | | | | 2.71 | 12.55 | 4.44 | | |

| HG MS & Tuff | BL758-56 | 89.7 | 67.3 | 84.2 | 18.3 | 89.7 | 92.4 | 81.6 | 67.4 | 88.9 | 84.2 | 2.37 | 4.50 | 4.15 | 0.22 | 0.58 |

| LG MS & Tuff | BL758-57 | 84.3 | 83.8 | 99.6 | 15.5 | 87.8 | 93.0 | 75.6 | 87.5 | 85.9 | 92.1 | 0.49 | 5.20 | 2.32 | 0.08 | 0.36 |

| HG Cu + Zn 1 | BL758-58 | 82.6 | 80.1 | 98.2 | 24.0 | 95.2 | 96.6 | 84.5 | 92.5 | 93.3 | 95.8 | 6.02 | 29.86 | 6.00 | 0.41 | 1.64 |

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Composite	Test Number	Rougher Flotation Concentrate Distribution			Rougher Concentrate Mass (%)	Flotation Concentrate Leach Extraction		Flotation Tailings Leach Extraction		Overall Flotation Leach Extraction		Calculated<br> <br>Head Grade			Overall<br> <br>Residue

| | | Au (%) | S (%) | Ag (%) | | Au (%) | Ag (%) | Au (%) | Ag (%) | Au (%) | Ag (%) | Au (g/t) | Ag (g/t) | S (%) | Au (g/t) | Ag (g/t) |

| HG Cu + Zn 2 | BL758-59 | 82.9 | 83.8 | 96.8 | 21.5 | 91.2 | 95.7 | 83.3 | 85.7 | 89.8 | 94.1 | 1.84 | 19.33 | 4.28 | 0.20 | 1.41 |

| LG Breccia | BL758-60 | 86.2 | 85.5 | 99.1 | 20.5 | 85.4 | 93.5 | 90.4 | 91.1 | 86.1 | 93.2 | 0.58 | 5.49 | 2.99 | 0.09 | 0.41 |

| HG Breccia | BL758-61 | 84.1 | 81.6 | 97.4 | 23.9 | 91.8 | 97.2 | 84.2 | 83.2 | 90.6 | 94.6 | 2.25 | 14.88 | 4.06 | 0.22 | 1.01 |

| HG Sample 1 | BL758-62 | 84.2 | 61.9 | 97.5 | 17.9 | 87.0 | 90.2 | 86.2 | 90.8 | 86.9 | 90.4 | 1.35 | 4.74 | 3.79 | 0.18 | 0.37 |

| HG Sample 2 | BL758-63 | 84.2 | 80.8 | 97.9 | 26.2 | 89.8 | 96.5 | 83.9 | 85.9 | 88.8 | 94.5 | 2.20 | 11.55 | 4.53 | 0.24 | 0.79 |

| LG Sample 1 | BL758-64 | 79.7 | 72.6 | 97.4 | 12.2 | 82.2 | 91.3 | 87.9 | 93.4 | 83.4 | 91.8 | 0.52 | 6.42 | 2.80 | 0.09 | 0.57 |

| LG Sample 2 | BL758-65 | 81.1 | 67.5 | 86.2 | 15.0 | 80.9 | 91.5 | 71.8 | 85.7 | 79.2 | 89.6 | 0.58 | 4.97 | 3.56 | 0.14 | 0.47 |

| LG Sample 3 | BL758-66 | 84.0 | 78.4 | 97.9 | 20.9 | 82.5 | 96.7 | 84.5 | 81.2 | 82.9 | 93.4 | 0.54 | 4.77 | 3.81 | 0.10 | 0.28 |

| HG Trachyte | BL758-67 | 73.1 | 67.1 | 95.3 | 16.9 | 87.1 | 92.0 | 90.9 | 92.9 | 88.2 | 92.3 | 1.42 | 7.58 | 4.41 | 0.17 | 0.64 |

| LG Trachyte | BL758-68 | 88.5 | 78.9 | 97.8 | 22.8 | 82.5 | 92.6 | 94.0 | 88.5 | 83.8 | 91.7 | 0.47 | 4.75 | 3.43 | 0.09 | 0.35 |

| High As | BL758-69 | 89.06 | 81.45 | 96.83 | 20.3 | 95.1 | 96.0 | 88.4 | 92.3 | 94.4 | 95.3 | 3.06 | 4.30 | 5.01 | 0.16 | 0.24 |

Note:

1.	Repeat tests were carried out to improve sulphur recovery to concentrate

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The leach variability tests in Program BL758 show that:

·	leach extraction of gold and silver from flotation concentrate and flotation tailings was highly variable;
·	concentrate leach is essentially complete after 8 h and tailings leach after 10 h. No relationship is apparent between the zone or lithology of the composite and the overall gold extraction expected;
·	the overall gold extraction varied with head grade, and the low grade samples performed slightly worse than their higher grade counterparts;
·	the leach concentrate residue grades remained consistent except for the very high grade sample in test BL758-51; and
·	there was no significant reduction in overall gold extraction in the samples with high base metal content or arsenic

Average sodium cyanide consumption from the Production and Master Composite samples (tests BL758-35, 37 and 39) was 0.12 kg/t of tailings for the tailings leach circuit and 2.87 kg/t of concentrate for the concentrate leach circuit. Average lime consumption was 1.33 kg/t of tailings and 3.98 kg/t of concentrate.

Overall sodium cyanide consumption of the Variability Composites ranged between 0.6 and 1.4 kg/t, with one of the high copper composite samples (High Cu & Zn 1) consuming 1.7 kg/t. Lime consumption ranged between 0.9 and 1.4 kg/t.

A sample of the Production Composite underwent a multi-stage diagnostic leach to assess the distribution of gold in the two leach tailings streams. The diagnostic leach test on the leach products from the Production Composite sample showed that about 78% of the gold in the flotation tailings leach residue was amenable to leaching with a high concentration of cyanide, however this is unlikely to be economic. About 72% of the gold in the concentrate leach residue was locked in sulphides and about 11% was locked in carbonates indicating a refractory component unrecoverable by the selected flowsheet. About 17% was locked in silicates which may be recovered by finer grinding; however this is unlikely to be economic.

Preg-robbing was investigated through comparative leach tests carried out over 48 h with and without the addition of carbon, on the Zone Composite samples. Similar leach extractions were achieved with and without carbon, indicating that preg-robbing is not likely to be of concern for this project.

The oxygen uptake rates were determined for rougher concentrate leach and rougher tailings leach on the Production Composite sample. The concentrate is a high consumer of oxygen at 4 kg/t of leach feed, while the tailings is a low consumer of oxygen at 0.07 kg/t of leach feed.

Flotation and leach extraction test results for samples tested in Program BL1073 are shown in Table 13‑15.

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Table 13‑15: Flotation and Leach Extraction for Composites and Variability Samples (Program BL1073)

Test	Comp	Rougher Conc.			Rougher Tailings		Leach Recovery (24 hours)

| | | (%) | Au (%) | Ag (%) | Au (%) | Ag (%) | Au (%) | Ag (%) | Au (%) | Ag (%) | Au (%) | Ag (%) |

| R01/01B/01D | MC3 | 14.2 | 78.7 | 72.5 | 21.3 | 27.5 | 86.2 | 869 | 87.5 | 86.4 | 81.6 | 88.3 |

| R02/02B/02D | MC4 | 12.1 | 66.6 | 70.7 | 33.4 | 29.3 | 88.7 | 91.4 | 89.9 | 94.1 | 86.4 | 85.1 |

| R03/03M/03N | MC5 | 14.0 | 75.6 | 57.4 | 24.4 | 42.6 | 85.7 | 86.3 | 85.8 | 86.4 | 85.4 | 86.2 |

| R04/04L/04M | MC6 | 15.4 | 81.0 | 87.5 | 19.0 | 12.5 | 87.3 | 90.9 | 86.4 | 91.4 | 91.0 | 87.4 |

| R05/05C/05B | Zone A | 7.6 | 87.4 | 88.4 | 12.6 | 11.6 | 70.0 | 34.7 | 8.4 | 29.1 | 80.7 | 77.4 |

| R06/06C/06B | Zone B | 21.8 | 91.1 | 79.4 | 8.9 | 20.6 | 69.6 | 37.4 | 68.0 | 29.1 | 58.8 | 69.8 |

| R07/07C/07B | Zone C | 20.0 | 89.7 | 84.5 | 10.3 | 15.5 | 90.3 | 80.9 | 90.8 | 79.6 | 85.9 | 88.1 |

| R08/08C/08B | Zone D | 12.0 | 90.3 | 81.3 | 9.7 | 18.7 | 83.2 | 79.8 | 82.7 | 77.8 | 87.0 | 88.4 |

| R09/09C/09B | Zone E | 26.4 | 88.4 | 88.5 | 11.6 | 11.5 | 71.6 | 71.6 | 70.4 | 69.7 | 80.6 | 86.3 |

| R10/10C/10B | High Cu/Zn | 27.2 | 82.8 | 86.8 | 17.2 | 13.2 | 85.9 | 82.2 | 84.2 | 80.3 | 93.9 | 94.7 |

The rougher concentrates of MC3 and MC4 were reground to a P80 of 17 µm and leached with cyanide. The concentrates were pre-aerated for 8 h prior to leaching. The leach tests were performed over a duration of 24 h with 2 g/L sodium cyanide, pH of 11, 30% solids and oxygen sparging. The tests for MC5 and MC 6 were performed at leach densities between 20 and 45% solids. Additional bulk leach tests, BL-1073 CN03M and CN04L, were performed at 45% solids to generate products for downstream testing. The rougher concentrate leach results are presented in Table 13‑16.Table 13‑16

Table 13‑16: Rougher Concentrate Leach Test Results at Varying Solids Content

Test Number	Composite	Leach Density (% solids)	Leach Au Extraction (%)	Leach Ag Extraction (%)

| BL-1073-CN01B | MC3 | 30 | 87.5 | 86.4 |

| BL-1073-CN02B | MC4 | 30 | 89.9 | 94.1 |

| BL-1073-CN03E | MC5 | 20 | 84.6 | 88.0 |

| BL-1073-CN03F | MC5 | 30 | 86.2 | 89.9 |

| BL-1073-CN03G | MC5 | 45 | 88.2 | 90.9 |

| BL-1073-CN03M | MC5 | 45 | 85.8 | 86.4 |

| BL-1073-CN04D | MC6 | 20 | 81.4 | 91.4 |

| BL-1073-CN04E | MC6 | 30 | 81.3 | 91.1 |

| BL-1073-CN04F | MC6 | 45 | 86.3 | 91.6 |

| BL-1073-CN04L | MC6 | 45 | 86.4 | 91.4 |

Optimum conditions for leaching rougher concentrate were determined to be leaching for 24 h at 45% solids with eight hours of pre-aeration.

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Initial bottle roll cyanide leach tests on rougher tailings from MC3 and MC4 were performed for 24 h, with 2 g/L sodium cyanide, pH of 11, slurry density between 40% and 53% solids and oxygen sparging. Additional optimization bottle roll leach testing on MC5 and MC6 was performed to evaluate leach time, solids density and carbon-in-leach (CIL). Two additional bulk tests, BL-1073-CN03N and CN04M, were conducted to generate material for downstream testing using CIL for 24 h. The rougher tailings leach results are presented in Table 13‑17.

Table 13‑17: Rougher Tailings Leach Test Results at Varying Conditions

Test Number	Composite	Leach Density (% solids)	Leach Au Extraction (%)	Leach Ag Extraction (%)	Leach Au Residue Grade (g/t)	Leach Time (hrs)	Test Description

| CN01D | MC3 | 53 | 81.6 | 88.3 | 0.05 | 24 | Leach |

| CN01G | MC3 | 53 | 80.9 | 96.9 | 0.05 | 24 | Leach |

| CN01F | MC3 | 45 | 73.7 | 96.9 | 0.08 | 24 | Leach |

| CN01E | MC3 | 40 | 78.7 | 97.0 | 0.07 | 24 | Leach |

| CN01H | MC3 | 53 | 88.8 | 84.0 | 0.03 | 24 | 6h Pre-ox, Leach |

| CN02D | MC4 | 53 | 86.4 | 85.1 | 0.04 | 24 | Leach |

| CN02G | MC4 | 53 | 85.1 | 97.5 | 0.04 | 24 | Leach |

| CN02F | MC4 | 45 | 83.5 | 97.5 | 0.05 | 24 | Leach |

| CN02E | MC4 | 40 | 86.0 | 97.5 | 0.04 | 24 | Leach |

| CN02H | MC4 | 53 | 88.9 | 83.0 | 0.03 | 24 | 6h Pre-ox, Leach |

| CN03H | MC5 | 40 | 74.1 | 90.3 | 0.08 | 24 | Leach |

| CN03I | MC5 | 40 | 80.6 | 64.3 | 0.06 | 12 | Leach |

| CN03J | MC5 | 40 | 87.4 | 79.9 | 0.04 | 12 | CIL |

| CN03K | MC5 | 40 | 88.4 | 82.0 | 0.03 | 18 | Leach |

| CN03L | MC5 | 40 | 86.2 | 85.8 | 0.05 | 18 | CIL |

| CN03N | MC5 | 53 | 85.4 | 86.2 | 0.04 | 24 | CIL |

| CN04G | MC6 | 40 | 69.4 | 88.1 | 0.06 | 24 | Leach |

| CN04H | MC6 | 40 | 85.8 | 83.9 | 0.03 | 12 | Leach |

| CN04I | MC6 | 40 | 87.8 | 82.3 | 0.03 | 12 | CIL |

| CN04J | MC6 | 40 | 83.5 | 83.9 | 0.03 | 18 | Leach |

| CN04K | MC6 | 40 | 90.2 | 82.7 | 0.02 | 18 | CIL |

| CN04M | MC6 | 53 | 91.0 | 87.4 | 0.02 | 24 | CIL |

With consideration for capital and operating costs, optimum conditions for leaching of rougher tailings were determined to be leaching at 53% solids for 24 h in a leach carbon-in-pulp (CIP) circuit with no pre-aeration.

13.3.6 Merrill Crowe Feed Solution Composition

The pregnant leach solution from the flotation concentrate leach tests was assessed for amenability to the Merrill-Crowe zinc cementation process. Results are shown in Table 13‑18.

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Table 13‑18: Leach Solution Analysis

Solution Species	Unit	Concentrations impair Zn precipitation	Concentration at which Zn precipitation ceases	Testwork result (MC3)	Testwork result (MC4)

| Sulphides | mg/L | >0.6 | 14 | 0.17 | 0.66 |

| Arsenic | mg/L | >0.1 | 17 | <0.4 | <0.4 |

| Antimony | mg/L | >0.1 | 20 | <0.4 | <0.4 |

| Copper Cyanide | mg/L | >0.2 or 25 | 850 | 913 (CNWAD)^1,^^2^ | 831 (CNWAD) |

| Nickel | mg/L | 150-500 | - | 6.7 | 4 |

| Cobalt | mg/L | >5 | - | 0.67 | 0 |

| Ca | mg/L | | | 1382 | 1000 |

| Hg | mg/L | | | 0.18 | 0.47 |

Note:

1.	Copper cyanide concentration not determined

| 2. | CNWAD is weak acid dissociable cyanide which includes Cu, Ni, Zn and Ag cyanide complexes |

Analysis of the pregnant leach solution produced from flotation concentrates shows no species of concern that would negatively impact precipitation efficiency.

13.3.7 Cyanide Detoxification

13.3.7.1 2021-2023 BaseMet Program

Cyanide detoxification tests were performed in Program BL758 on the leach products from the Production Composite sample using the SO2/air process. Rougher concentrate was reground to a target P80 size of 17 µm and leached at 20% solids with four hours of pre-aeration, rougher tailings were leached at 40% solids. The results of the concentrate leach cyanide detoxification tests are summarized below in Table 13‑19, and the results of the tailings leach cyanide detoxification tests are summarized in Table 13‑20.

Table 13‑19: Production Composite Concentrate Cyanide Detoxification Test Results

Test	Objective	Retention Time (min)	Discharge Chemistry (Solution)					Pulp Volume Treated (L)	Reagent Addition

| Feed | - | - | - | 751 | 659 | 38.3 | 32.9 | - | - | - | - |

| CND-C2 | <10 mg/L CNWAD | 150 | 8.6 | 0.9 | 0.7 | 0.5 | 0.05 | 2.0 | 8.0 | 7.4 | 108 |

| CND-C3 | <10 mg/L CNWAD | 90 | 8.4 | 1.2 | 1.0 | 0.35 | 0.05 | 2.7 | 8.0 | 6.1 | 108 |

| CND-C4 | <10 mg/L CNWAD | 60 | 8.4 | 1.0 | 0.9 | 0.99 | 0.05 | 2.1 | 8.0 | 8.5 | 108 |

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Test		Retention Time (min)	Discharge Chemistry (Solution)					Pulp Volume Treated (L)	Reagent Addition

| CND-C5 | <10 mg/L CNWAD | 60 | 8.3 | 0.9 | 0.7 | 0.63 | 0.05 | 2.2 | 5.0 | 4.9 | 108 |

| CND-C6 | <10 mg/L CNWAD | 60 | 8.2 | 12.7 | 1.7 | 0.13 | 3.94 | 2.4 | 5.0 | 8.4 | 50 |

| CND-C7 | <10 mg/L CNWAD | 45 | 8.3 | 33.8 | 2.5 | 0.18 | 11.2 | 3.3 | 5.0 | 5.8 | 30 |

Table 13‑20: Production Composite Tailings Cyanide Detoxification Test Results

Test	Objective	Retention Time (min)	Discharge Chemistry (Solution)					Pulp Volume Treated (L)	Reagent Addition

| FEED | - | - | - | 384 | 353 | 3.1 | 11.1 | - | - | - | - |

| CND-C1 | <10 mg/L CNWAD | 30 | 8.5 | 0.4 | 0.3 | 0.10 | <1 | 7.6 | 5.0 | 3.5 | 25 |

| CND-C2 | < 10 mg/L CNWAD | 30 | 8.5 | 0.8 | 0.7 | 0.23 | <1 | 13.4 | 3.0 | 1.6 | 25 |

| CND-C3 | < 10 mg/L CNWAD | 20 | 8.5 | 0.9 | 0.7 | 0.48 | <1 | 10.0 | 3.0 | 2.4 | 25 |

Two flotation tests were conducted on the Production Composite using solution containing 0.5 and 1 mg/L CNWAD from the cyanide detoxification process to assess the impact of cyanide recycle in process water on flotation. No adverse effects were seen on gold or sulphur recovery to concentrate.

In Program BL1073, MC3 and MC4 products, both leach concentrate and tailings, were advanced to continuous cyanide detoxification testing. Rougher concentrate was reground to a target P80 size of 17 µm and leached at 45% solids with 8 hours of pre-aeration, rougher tailings were leached at 53% solids. The pregnant leach solution (PLS) produced from the concentrate leach was filtered and treated by Merrill-Crowe. The barren PLS post-Merrill-Crowe as well as the leached solids containing residual cyanide were combined before being advanced to cyanide detoxification. The MC3 concentrate cyanide detoxification test results are summarized in Table 13‑21, and the MC4 concentrate cyanide detoxification test results are summarized in Table 13‑22.

The rougher tailings were leached at 53% solids with carbon in a CIL circuit, then the cyanide slurry was advanced directly to cyanide detoxification. The cyanide concentrations targeted for both tailings and concentrate streams were less than 5 mg/L CNWAD. The MC3 tailings cyanide detoxification test results are summarized in Table 13‑23, and the MC4 tailings cyanide detoxification test results are summarized in Table 13‑24.

Table 13‑21: MC3 Concentrate Cyanide Detoxification Test Results

Test	Retention Time (min)	Discharge Chemistry (Solution)					Pulp Volume Treated (L)	Reagent Addition

| Feed | - | - | 237 | 197 | 12 | 14 | - | - | - | - |

| CND-C1 | 60 | 8.0 | 3.0 | 1.5 | 0.14 | 0.54 | 1.13 | 5.0 | 9.9 | 50 |

| CND-C2 | 30 | 8.1 | 3.1 | 1.5 | 0.16 | 0.54 | 1.32 | 5.0 | 4.2 | 50 |

| CND-C3 | 30 | 8.0 | 7.4 | 1.9 | 0.12 | 1.95 | 1.36 | 5.0 | 4.1 | 25 |

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Table 13‑22: MC4 Concentrate Cyanide Detoxification Test Results

Test	Retention Time (min)	Discharge Chemistry (Solution					Pulp Volume Treated (L)<br> <br>pH	Reagent Addition

| Feed | - | - | 178 | 153 | 11.8 | 8.99 | - | - | - | - |

| CND-C1 | 30 | 8.0 | 8.6 | 1.4 | 0.15 | 2.6 | 1.36 | 5.0 | 4.4 | 25 |

| CND-C2 | 30 | 8.0 | 10.3 | 1.7 | 0.36 | 3.1 | 1.39 | 3.5 | 4.3 | 25 |

| CND-C3 | 30 | 8.2 | 2.3 | 1.1 | 0.14 | 0.42 | 1.39 | 3.5 | 5.3 | 25 |

| CND-C4 | 30 | 8.0 | 4.9 | 1.1 | 0.14 | 1.35 | 1.41 | 3.5 | 5.3 | 15 |

Table 13‑23: MC3 Tailings Cyanide Detoxification Test Results

Test	Retention Time (min)	Discharge Chemistry (Solution					Pulp Volume Treated (L)<br> <br>pH	Reagent Addition

| Feed | - | - | 445 | 440 | 8.1 | 1.9 | - | - | - | - |

| CND-C1 | 30 | 8.0 | 2.8 | 2.4 | 0.94 | 0.16 | 5.8 | 5.0 | 5.7 | 25 |

| CND-C2 | 30 | 8.1 | 13.9 | 12.6 | 10 | 0.46 | 3.9 | 5.0 | 6.4 | 25 |

| CND-C3 | 30 | 7.9 | 1.7 | 1.7 | 0.53 | 0.01 | 2.5 | 10.0 | 9.9 | 75 |

| CND-C4 | 60 | 8.0 | 1.4 | 1.2 | 0.43 | 0.05 | 2.5 | 10.0 | 14.9 | 75 |

| CND-C5 | 60 | 8.1 | 1.0 | 1.0 | 0.56 | 0.01 | 2.4 | 10.0 | 12.9 | 150 |

Table 13‑24: MC4 Tailings Cyanide Detoxification Test Results

Test	Retention Time (min)	Discharge Chemistry (Solution					Pulp Volume Treated (L)<br> <br>pH	Reagent Addition

| Feed | - | - | 402 | 387 | 4.7 | 5.3 | - | - | - | - |

| CND-C1 | 60 | 8.0 | 2.4 | 2.4 | 1.2 | 0.01 | 2.5 | 5.0 | 14.7 | 125 |

| CND-C2 | 60 | 8.0 | 1.9 | 1.7 | 0.88 | 0.04 | 1.6 | 5.0 | 13.0 | 125 |

| CND-C3 | 120 | 8.1 | 1.0 | 0.9 | 0.29 | 0.03 | 2.2 | 7.5 | 28.4 | 125 |

| CND-C4 | 120 | 8.1 | 0.5 | 0.5 | 0.32 | 0.01 | 2.2 | 10.0 | 20.6 | 125 |

The optimum conditions for concentrate were determined to be a 5:1 ratio of SO2 to CNWAD, 25 mg/L Cu and 30 minutes retention time.

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Rougher tailings for MC3 achieved the target CNWAD concentration with a 5:1 ratio of SO2 to CNWAD, 25 mg/L Cu and 30 minutes retention time. MC4 achieved the target CNWAD concentration, however a shorter time and lower copper sulphate addition was not tested.

13.3.8 Tailings Dewatering

13.3.8.1 2021 FLSmidth Program

A tailings sample generated from the Production Composite in BaseMet program BL758, with a primary grind P80 size of 144 µm was studied to establish settling and filtration characteristics. Tests were conducted at FLSmidth’s laboratory in Salt Lake City.

The dynamic settling testwork achieved an underflow solids slurry density of 70% by weight using flocculant AN920 dosed at 16 g/t, with a settling rate of 0.04 m^2^/t/d.

Vacuum filtration achieved a moisture content of 18.5% with a filtration rate of 875 kg/m^2^/h.

Pressure filtration tests were conducted using a bench-scale filtration testing unit which can simulate FLSmidth’s recessed chamber and membrane squeeze filter characteristics. The test apparatus allows for variations in fill pressures, cake thicknesses, fill times, air blow times as well as air blow pressures. Pressure filtration test results are summarized in Table 13‑25.

Table 13‑25: Summary of Tailings Pressure Filtration Test Results

Measurement	Tailings

| Test ID | 1 | 2 | 3 | 4 |

| Feed Density (% solids by weight) | 70 | | | |

| Form Pressure (kPa) | 690 | | | |

| Cake Blow Pressure (kPa) | 970 | | | |

| Membrane Squeeze Pressure (kPa) | - | - | 970 | 970 |

| Cake Thickness (mm) | 32 | 50 | 32 | 50 |

| Formation Time (minutes) | 1 | 1 | 1 | 1 |

| Air Blow Time (minutes) | 2 | 1.5 | 6 | 8 |

| Membrane Squeeze Time (minutes) | - | - | 6 | 8 |

| Final Cake Moisture (% by weight) | 12 | 13 | 8.5 | 9 |

| Filtration Rate (kg/m^2^/hr) | 222 | 374 | 147 | 196 |

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The pressure filtration tests achieved a final cake moisture content of 9–12% by weight.

13.3.8.2 2021-2023 BaseMet Program

Solid-liquid separation tests were carried out in Program BL758 on the Variability Composites (rougher tailings leach residue) to test for differences in settling and filtration properties as a function of zone and lithology. The results of these tests are summarized below in Table 13‑26.

Table 13‑26: Summary of Static Settling Test Results

Composite	Flocculant		pH	Density (% solids by weight)		Free Settling Velocity (m/h)

| HG MS and TUFF | MF155 | 40 | 8.4 | 13.7 | 53.4 | 18.0 |

| LG MS and TUFF | MF155 | 40 | 10.8 | 13.7 | 65.7 | 30.3 |

| LG Breccia | MF10 | 40 | 10.8 | 13.7 | 50.6 | 28.4 |

| HG Breccia | MF10 | 40 | 10.8 | 13.7 | 48.2 | 13.3 |

| HG Trachyte | MF10 | 40 | 10.8 | 13.7 | 54.2 | 26.5 |

| LG Trachyte | MF10 | 40 | 10.8 | 13.7 | 47.4 | 11.5 |

| Zone A | MF10 | 40 | 10.8 | 13.7 | 66.3 | 19.8 |

| Zone B | MF10 | 40 | 10.8 | 13.7 | 60.9 | 15.9 |

| Zone C | MF10 | 40 | 10.8 | 13.7 | 61.0 | 19.3 |

| Zone D | MF10 | 40 | 10.8 | 13.7 | 56.3 | 15.1 |

| Zone E | MF10 | 40 | 10.8 | 13.7 | 62.2 | 14.6 |

The final density results ranged from 47.4% to 66.3% solids by weight, but among the more blended (mixed lithology) Zone Composites the range was tighter between 56.3% and 66.3% by weight.

The filtration test results are summarized in Table 13‑27.

Table 13‑27: Summary of Pressure Filtration Test Results

Composite	Pressure (kPa)	Blow Time (sec)		Cake Thickness (mm)	% Moisture	Filter Rate (kg/m ^2^ /hr)

| Zone A | 550 | 180 | 11 | 23 | 7.8 | 670 |

| Zone B | 550 | 180 | 7 | 26 | 7.8 | 687 |

| Zone C | 550 | 180 | 5 | 23 | 8.8 | 674 |

| Zone D | 550 | 180 | 11 | 23 | 10.6 | 660 |

| Zone E | 550 | 180 | 10 | 22 | 10.8 | 665 |

| HG MS and Tuff | 1240 | 5 | 23 | 58.7 | 85.2 | 3625 |

| LG MS and Tuff | 1240 | 8 | 24 | 59.1 | 85.7 | 2202 |

| LG Breccia | 550 | 180 | 4 | 24 | 9.0 | 668 |

| HG Breccia | 550 | 180 | 10 | 23 | 11.6 | 650 |

| HG Trachyte | 550 | 180 | 7 | 24 | 10.1 | 672 |

| LG Trachyte | 550 | 180 | 7 | 23 | 9.1 | 680 |

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The test results showed similar cake moistures and filter rates except for the MS and Tuff Composites. However, these two tests most likely had breakthroughs and were not considered reliable.

13.3.9 Concentrate Dewatering

Concentrate dewatering testing was performed to assess static and dynamic thickening. The flocculant AN913SH was found to be optimal based on testing on MC3 rougher concentrate. The dynamic thickening test results are summarized in Table 13‑28.

Table 13‑28: Dynamic Settling Test Results for Rougher Concentrate

Composite	Test	Loading Rate (t/m ^2^ /hr)	Flocculant Dosage (g/t)	Underflow Density (% solids)

| T01 Rougher Conc MC3 | D1A | 0.3 | 40 | 50.2 |

| | D1B | 0.5 | 60 | 49.0 |

| | D1C | 0.5 | 60 | 51.8 |

| | D1D | 0.5 | 40 | 51.3 |

| T02 Rougher Conc MC4 | D2A | 0.5 | 60 | 55.0 |

| | D2B | 0.7 | 60 | 52.7 |

| | D2C | 1.0 | 60 | 48.9 |

The settling rates and underflow densities showed good settling characteristics.

The unsheared and high shear viscosities of MC3 and MC4 rougher concentrates were measured using a Brookfield DV2T viscometer with a vane spindle. Slurry viscosity was 0.051 to 0.052 Pa .s at the shear rate of 130 s^-1^ which is acceptable for pumping applications. Viscosity was 0.53 to 0.64 Pa,s at the shear rate of 4.5 s^-1^ which is acceptable for mixing and screening applications.

Both bench scale vacuum and pressure filtration tests were conducted on the rougher concentrates from MC3 and MC4.

The vacuum filtration tests at different filter cake thicknesses resulted in final moistures ranging between 30 and 35%.

For pressure filtration, the concentrate was first thickened to between 52 and 55% solids. Pressure filtration was conducted at a constant pressure of 550 kPa using a Micronics pressure filter unit with a filter area of 0.0016 m^2^. A moisture content of 10% was achieved with five minutes air blow and cake thickness of 11 mm with a filtration rate of 183 kg/m^2^/h for MC3. A moisture content of 16% was achieved with five minutes air blow and cake thickness of 23 mm with a filtration rate of 381 kg/m^2^/h for MC4.

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13.4 Metallurgical Variability

Comminution tests were conducted on material from a total of 14 drill holes and 37 samples were tested for Bond Ball Mill Work Index (BWi), 21 for Bond Abrasion Index (Ai) and 23 for SAG Power Index (SPI) during the 2020 SGS and 2021-2023 BaseMet programs. Flotation and leach variability tests were conducted on 24 samples from 10 drill holes in the 2021-2023 BaseMet program and on 12 samples from 4 drill holes in the 2020 SGS program.

For the 2021-2023 BaseMet program, 10 drill holes were selected. These drill holes are supplemental to the drill holes used for metallurgical testwork in the 2020 SGS program. The 2021-2023 BaseMet program plan determined a requirement for approximately 1,200 kg of sample, and the number of drill holes was selected in collaboration with First Mining to provide the required mass of material for testing, as well as to meet the requirements for variability testing. Variability composites (10 kg each) were created as follows:

·	Zone composites to investigate spatial variability.
·	Grade composites to investigate effect of feed grade variability.
·	Lithology composites to investigate variability due to lithology or rock group.

The locations of all drill holes and intervals used for metallurgical testing for the 2020 SGS and 2021-2023 BaseMet programs are shown in Figure 13‑11.

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Figure 13‑11: Drill Hole and Interval Locations for Samples in the 2020 SGS and 2021-2023 BaseMet Programs

Source: First Mining, 2022

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13.5 Deleterious Elements

Head assays confirm the presence of mercury at concentrations that warrant mercury control. Some areas of elevated copper and zinc concentrations are seen which will result in higher than average cyanide consumption. No other deleterious elements have been identified.

13.6 Recovery Estimates

Estimation of recovery is based on the 2020 SGS and 2021-2023 BaseMet program data. Results from the 2020 SGS test numbers F13, F23 and F24 were excluded from the evaluation due to high mass recovery to flotation concentrates (>40%) caused by foaming compound, and due to a coarse primary grind P80 size of 221µm.

Gold recovery is estimated from testwork data as a function of feed sulphur and gold grades as follows:

·	for >0.7g/t Au: plant recovery =-2.1853 * (Feed S/Feed Au)+ 91.716
·	where Feed S is expressed in %, and Feed Au in g/t
·	for <0.7g/t Au and S/Au>4: plant recovery =79%

Silver recovery is estimated as a function of silver feed grade as follows:

·	Plant recovery = 7.7014 * ln(Feed Ag) + 73.314
·	where Feed Ag is expressed in g/t

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Figure 13‑12: Plant Recovery for Gold (includes 1.7% deduction for plant losses)

Source: Ausenco, 2025

Figure 13‑13: Plant Recovery for Silver (includes 2.2% deduction for plant losses)

Source: Ausenco, 2025

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Plant losses for gold and silver were calculated based on the sources and concentrations in Table 13‑29.

Table 13‑29: Parameters Used for Estimation of Plant Losses

Description	Value - Au	Value - Ag	Units

| Tailings CIP solution losses | 0.010 | 0.05 | mg/L |

| Fine carbon losses | 30 | 30 | g/t |

| Gold loading on fine carbon | 100 | 200 | g/t |

| Other losses (scats) | 0.1 | 0.1 | % |

| Conc CCD solution losses | 0.04 | 0.5 | mg/L |

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14 MINERAL RESOURCE ESTIMATES

14.1 Introduction

The mineral resource statement presented herein represents the sixth mineral resource evaluation that has been prepared for the Springpole Gold Project in accordance with NI 43-101 and the current CIM Definition Standards, 2014.

There is a total of 853 drill holes in the Springpole database provided to the QP, 212 drilled by First Mining and 641 drilled by the previous owners of the Property. The current mineral resource model prepared by the QP utilizes results from 499 core boreholes. The resource estimation work was completed by Dr. Gilles Arseneau, P.Geo. (APEGBC No. 23474), an independent Qualified Person as this term is defined in NI 43-101. The effective date of the resource statement is September 30, 2025.

This section describes the resource estimation methodology and summarizes the key assumptions considered by the QP. In the opinion of the QP, the resource evaluation reported herein is a reasonable representation of the global gold and silver resources found in the Springpole Gold Project at the current level of sampling. The mineral resources were estimated in conformity with the CIM Best Practices Guidelines and are reported in accordance with the CIM Definition Standards, 2014.

The database used to estimate the mineral resources was audited by the QP. The QP is of the opinion that the current drilling information is sufficiently reliable to interpret with confidence the boundaries for porphyry gold mineralization and that the assay dataset is sufficiently reliable to support mineral resource estimation.

GEMS (6.8.4) software was used to construct the geological solids, prepare assay data for geostatistical analysis, construct the block model, estimate metal grades, and tabulate mineral resources. The geostatistical software SAGE2001 was used for variography.

14.2 Resource Estimation Procedures

The resource evaluation methodology involved the following procedures:

·	Database compilation and verification.
·	Construction of wireframe models for the boundaries of the Springpole gold mineralization.
·	Definition of resource domains.
·	Data compositing and capping for geostatistical analysis and variography.
·	Block modelling and grade interpolation.
·	Resource classification and validation.
·	Assessment of “reasonable prospects for economic extraction” and selection of appropriate cut off grades (COGs).
·	Preparation of the mineral resource statement.

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14.3 Drill Hole Database

The Project currently consists of three separate mineralized zones: East Extension, Camp, and Portage. The Portage zone is by far the largest of the three and represents more than 90% of the stated resource.

The entire Springpole database provided to the QP consists of 853 drill holes totalling 206,672 m. Of these, 356 drill holes were not used in the estimation because they didn’t intersect the mineralized zone. Included in the excluded holes are two metallurgical holes and two geotechnical holes that intersected the mineralized zones but were not assayed.

14.4 Core Recovery

Drill core recovery for both the East Extension and Camp zones was generally very good with average recovery recorded as approximately 97%. For Portage, with areas of intense argillic and potassic alteration, core recovery is a much more significant issue, primarily affecting near surface intervals and intervals that appear to intersect a narrow zone of intense biotitic alteration.

The QP conducted studies to determine if there was any significant gold bias indicated, either high or low, as a function of core recovery. To a certain extent it was anticipated that more intense zones of alteration could also often reflect more intense mineralization.

Core recovery was generally recorded in 3 m intervals, with some data recorded in 1.5 m intervals. Consequently, for this analysis, it was decided to composite the core recovery values to the 3 m sample lengths and compare them with assay grades. The comparison indicates that the gold grade is generally lower with the increased recoveries (Figure 14‑1). For this reason, the QP decided to separately model areas of low core recoveries by creating a domain for rock with <40% recovery and treat those volumes as having hard boundaries during grade interpolation.

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Figure 14‑1: Gold Grade versus Core Recovery Relationship

Source: SRK, 2019

14.5 Geological Domains

The Portage zone is by far the most significant domain, extending from beneath the southern extent of the Camp zone and striking for more than 1,500 m to the southeast. Other than location, the Portage zone exhibits fewer similarities with the other two domains. Also, in contrast with East Extension and Camp, Portage has significant silver mineralization closely associated with gold. Drill-tested mineralization is extremely continuous with very little evidence of isolated erratic higher-grade intervals. As drilled, Portage represents a zone of largely disseminated mineralization striking 135° and extending from the surface to a depth of over 400 m, on average approximately 150 m in width and over 1,500 m in length.

The Camp zone is located to the north portion of the mineralization area and, where the two domains overlap, above the Portage zone. The Camp zone strikes approximately 120° and part of the zone is very similar in character to the East Extension with higher variability in gold and silver grades with limited spatial organization.

The East Extension zone is located to the east of the Camp zone and northeast or the Portage Zone, interpreted as more distally portion of the mineralization system which could account for some portions of grade variability. The East Extension zone strikes approximately 135° and is similar in character to the Camp zone.

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Geological domain boundaries were defined on sections spaced at 50 m intervals using a cut-off grade of 0.2 g/t Au. Figure 14‑2 shows the Springpole drill plan with the three geological domain areas and Figure 14‑3 shows the domain boundaries on a typical section.

Figure 14‑2: Geological Domains for Springpole Gold Project

Source: First Mining, 2025

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Figure 14‑3: Cross Section 1100NW Looking NW Showing Portage and East Extension Domains

Source: SRK, 2022. Note: Grid is 200 by 200 m. Green drill hole traces are > 0.2 g/t and red traces are > 0.3 g/t Au

14.6 Surface Topography

Topography was provided in the form of a Drawing Interchange Format file containing data from a 2011 light detection and ranging (LiDAR) survey with vertical precision of 1 m. The topographic surface beneath the small portion of the lake overlying the Portage zone was established by ground penetrating radar, echo sounder and sub-bottom profiling surveys conducted by Terrasond Ltd. of Palmer, Alaska, from the frozen surface (March 2011) and water lake surface (June 2011). These multiple surfaces were then merged to create a continuous surface to constrain the top of the block model. Overburden surface was modelled by extracting the base of the overburden from all available drill hole logs and generating a surface by simple triangulation of drill hole points.

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14.7 Compositing

An analysis of the sample lengths within the mineralized domains shows that sample lengths are variable ranging from a low of 0.1 m to a maximum of 21 m; however, the majority of the samples are between 0.5 and 3 m in length with the largest proportion of the samples at 1 m in length (Figure 14‑4). Most samples, 99%, are less than 3 m in length and for this reason the QP decided to composite all assays to a 3 m length within the mineralized envelopes. Compositing was generated from the drill collars and compositing was interrupted at domain boundaries. The compositing process generated 25,380 composites. A total of 275 composites with length less than 1.5 m were discarded from the database prior to resource estimation.

Figure 14‑4: Histogram of Sample Lengths within Mineralized Domains

Source: SRK, 2022

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14.8 Grade Capping

The primary goal of grade capping is to identify and restrict the influence of suspected “outlier” grades in an estimate.

Grade capping for the Springpole Gold Project was carried out in two stages. First the assay dataset was investigated to determine if sample length could bias the average grade. An analysis of gold grade against sample length seems to indicate that sample length of less than 1 m has a significantly higher average grade than other sample lengths, indicating these samples were taken over a specific geological domain, perhaps quartz veins or narrow siliceous zones with visible gold (Figure 14‑5). Most short sample lengths seem to have been taken from the Camp and East Extension zones; for this reason, the QP decided to treat these short sample lengths as a separate statistical population and capped these short assays prior to compositing. The QP capped all gold assays for sample lengths less than 1 m to 100 g/t Au prior to compositing.

All assays were then composited to 3 m lengths and all 3 m composites were evaluated for outliers by examining their distribution on cumulative probability plots and capped as outlined in Table 14‑1.

Figure 14‑5: Comparison of Sample Length and Average Gold Grade

Source: SRK, 2022

Table 14‑1: Capping Levels for Springpole

Element	3 m Composite Capping Level

| Au | 25 g/t |

| Ag | 200 g/t |

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14.9 Statistical Analysis and Variography

Statistical analyses were carried out on both the raw assay data and on the 3 m composited data. There is a total of 131,033 entries in the drill hole assay table for the Springpole Gold Project. Of these, 56,072 are within the interpreted wireframes representing the three mineralized domains. Of the data accepted, 16,782 samples are missing silver assays because the historical holes were only selectively assayed for silver; 9,147 from the Portage zone and 7,635 from the Camp and East Extension. The missing silver values were ignored so as to not overly underestimate the silver content of the mineralized zone. The assay data is also missing 414 gold assays from the Portage zone due to poor core recovery; these were also removed from the database prior to estimation. Table 14‑2 presents the statistical data for the assay data within the mineralized zones and Table 14‑3 shows the information for the 3 m composited data.

Table 14‑2: Basic Univariate Statistical Information for Raw Uncapped Assay Data

Zone	Max (g/t)	Min	Mean (g/t)	Std. Dev.	CoV	Count

| East Extension Gold | 1568 | 0.005 | 2.05 | 29.88 | 14.58 | 7028 |

| Camp Gold | 459 | 0.005 | 1.13 | 10.04 | 8.88 | 4184 |

| Portage Gold | 207 | 0.005 | 0.76 | 2.34 | 3.08 | 44786 |

| Portage Silver | 1000 | 0.02 | 4.12 | 12.22 | 2.96 | 35713 |

Note: 9,147 samples have no silver assays in the Portage Zone; the missing data was excluded from the above table

Table 14‑3: Basic Univariate Statistical Information for 3 m Composites

Zone	Max (g/t)	Number capped	Mean (g/t)	Std. Dev.	CoV	Count

| East Extension Gold | 327.89 | - | 1.07 | 7.34 | 6.86 | 2801 |

| East Extension Capped Gold | 25.00 | 17 | 0.85 | 2.70 | 3.18 | 2801 |

| Camp Gold | 42.05 | - | 0.71 | 2.40 | 3.38 | 1635 |

| Camp Capped Gold | 25.00 | 2 | 0.68 | 2.08 | 3.06 | 1635 |

| Portage Gold | 87.05 | - | 0.75 | 1.71 | 2.28 | 20255 |

| Portage Capped Gold | 25.00 | 13 | 0.73 | 1.32 | 1.81 | 20255 |

| Portage Silver | 369.00 | - | 4.41 | 10.84 | 2.46 | 16497 |

| Portage Capped Silver | 200.00 | 7 | 4.37 | 9.80 | 2.24 | 16497 |

Note: There are 414 composites with no gold values and 3,758 with missing silver assays in the Portage zone, these were excluded from the table above and were not used during grade interpolation

Spatial continuity of gold and silver was evaluated with correlograms developed using SAGE 2001 version 1.08 software. The correlogram measures the correlation between data values as a function of their separation distance and direction. The distance at which the correlogram is close to zero is called the “range of correlation” or simply the range. The range of the correlogram corresponds roughly to the more qualitative notion of the “range of influence” of a sample or composite.

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Variographic analysis was completed for gold in the Portage, Camp, and East Extension zones and for silver in the Portage zone. Directional correlograms were generated for composited data at 30° increments along horizontal azimuths. For each azimuth, correlograms were calculated at dips of 0°, 30°, and 60°.

A vertical correlogram was also calculated. Using information from these 37 correlograms, SAGE determines the best fit model using the least square fit method. The correlogram model is described by the nugget (C0) and two nested structure variance contributions (C1, C2) with ranges of the variance contributions and the model type (spherical or exponential). After fitting the variance parameters, the algorithm then fits an ellipsoid to the 37 ranges from the directional models for each structure. The final models of anisotropy are given by the lengths and orientations of the axes of the ellipsoids.

The correlogram models applied in the resource estimates in each domain are presented in Table 14‑4.

Table 14‑4: Gold and Silver Spherical Correlogram Parameters by Domain

Domain	Metal	Nugget	Sill	Gemcom Rotations (RRR rule)			Ranges a1, a2

| Camp | Au | 0.3 | 0.67 | -27 | 57 | 52 | 26 | 8 | 5 |

| | | | 0.03 | -27 | 57 | 52 | 61 | 57 | 180 |

| East Extension | Au | 0.3 | 0.48 | -6 | -67 | -72 | 7 | 11 | 15 |

| | | | 0.22 | -6 | -67 | -72 | 20 | 49 | 150 |

| Portage | Au | 0.19 | 0.56 | 31 | 8 | 34 | 20 | 40 | 20 |

| | | | 0.25 | 31 | 8 | 34 | 60 | 138 | 168 |

| Portage | Ag | 0.1 | 0.61 | -48 | 30 | 27 | 22 | 9 | 18 |

| | | | 0.29 | -48 | 30 | 27 | 100 | 76 | 174 |

14.10 Block Model and Grade Estimation

Block modelling was carried out in GEMS (6.4) software by Dr. Gilles Arseneau, P.Geo., an Associate Consultant with SRK. Block estimates were carried out in 10 by 10 by 6 m blocks using a percent model to weight partial blocks situated at zone boundaries. Block model parameters are defined in Table 14‑5.

Table 14‑5: Block Model Setup Parameters

	Model Origin<br> <br>(UTM - WGS 84)	Block Size<br> <br>(m)	No. of Blocks

| Easting | 548,500 | 10 | 220 |

| Northing | 5,692,400 | 10 | 210 |

| Elevation | 418 | 6 | 90 |

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14.10.1 Gold and Silver Grade Models

Grades were estimated by Ordinary Kriging with a minimum of 4 and a maximum of 15 composites with no more than three composites permitted from a single drill hole. Grade interpolations were carried out in three passes with each successive pass using a larger search radius than the preceding pass and only estimating the blocks that had not been interpolated by the previous pass. Table 14‑6 summarizes the search parameters for each interpolation pass.

Table 14‑6: Search Parameters by Zone and Metal

Metal	Zone	Pass	Rotation			Search Ellipse Size			Number of Composites		Max. Samples per Drill Hole

| | | | Z | Y | Z | X (m) | Y (m) | Z (m) | Min. | Max. | |

| Au | Camp | 1 | -84 | 7 | -32 | 20 | 30 | 20 | 4 | 15 | 3 |

| Au | Camp | 2 | -84 | 7 | -32 | 40 | 60 | 60 | 4 | 15 | 3 |

| Au | Camp | 3 | -84 | 7 | -32 | 60 | 138 | 168 | 4 | 15 | 3 |

| Au | East Ext | 1 | -84 | 7 | -32 | 20 | 30 | 20 | 4 | 15 | 3 |

| Au | East Ext | 2 | -84 | 7 | -32 | 40 | 60 | 60 | 4 | 15 | 3 |

| Au | East Ext | 3 | -84 | 7 | -32 | 60 | 138 | 168 | 4 | 15 | 3 |

| Au | Portage | 1 | -84 | 7 | -32 | 20 | 30 | 20 | 4 | 15 | 3 |

| Au | Portage | 2 | -84 | 7 | -32 | 40 | 60 | 60 | 4 | 15 | 3 |

| Au | Portage | 3 | -84 | 7 | -32 | 60 | 138 | 168 | 4 | 15 | 3 |

| Ag | Portage | 1 | -48 | 30 | 27 | 20 | 30 | 20 | 4 | 15 | 3 |

| Ag | Portage | 2 | -48 | 30 | 27 | 40 | 60 | 60 | 4 | 15 | 3 |

| Ag | Portage | 3 | -48 | 30 | 27 | 100 | 76 | 100 | 4 | 15 | 3 |

Uncapped gold was also estimated for all three domains for comparison against the capped results. The capped estimates were used for use in resource reporting and classification.

14.10.2 Bulk Density Model

There are 626 bulk density entries in the Springpole database with an average of 2.60 t/m^3^. The QP is of the opinion that these are sufficient to estimate a Mineral Resource. The data also included an additional 140 specific gravities (“SG”) from the 2012 drill program which were determined by pycnometer on pulverized material (which does not take into account void space); these values were notably higher than the bulk density readings and were ignored so as to not over-estimate the average bulk density.

The QP decided to estimate the bulk density by inverse distance squared where a dataset was nearby or assign an average density to un-estimated blocks, as presented in Table 14‑7.

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Table 14‑7: Bulk Density of Un-estimated Blocks in the Model

Zone	Average Density of Un-Estimated Blocks (t/m3)

| Camp | 2.91 |

| East Extension | 2.73 |

| Portage | 2.52 |

| Waste rock | 2.74 |

| Overburden | 1.9 |

14.10.3 Total Sulphur and Arsenic Models

In addition to gold and silver grades, total sulphur and arsenic were estimated in the block model to facilitate the determination of acid base accounting (ABA) of the mineralized and waste rock.

Sulphur was determined from 21,907 data points derived from the historical ICP assay database and new analytical data collected by First Mining from mainly historical core in the winter of 2019/2020. Samples collected during the 2019/2020 winter program were analyzed by SGS using a LECO sulphur analyzer with high temperature combustion.

From the samples collected in 2019-2020, 53 intervals have corresponding historical ICP assay values. The QP examined the duplicate samples and concluded that while the ICP assays seemed to be slightly lower than the total sulphur data, the bias was not significant enough to reject the historical data (Figure 14‑6). The QP also noted that the historical ICP sulphur data had an upper detection limit set at 5% sulphur and that 193 samples collected during the 2010 drilling program were not re-assayed to determine their exact sulphur content, effectively capping these samples at 5% sulphur.

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Figure 14‑6: Comparison of Historical S% by ICP with Re-assayed Total S%

Source: SRK, 2021

The QP decided to evaluate the sulphur data by mineralized zones and rock type to determine if sulphur content correlated with mineralization or rock type. The analysis showed that sulphur content is very similar inside and outside of the Portage zone and that the mineralized zone contact is not a hard boundary with respect to sulphur content (Figure 14‑7).

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Figure 14‑7: Contact Profile for Sulphur for the Portage Mineralized Zone

Source: SRK, 2022

Sulphur data was also compared with the updated lithology model from 2021 (Table 14‑8). Except for the Sand Zone, all rock types display similar sulphur content, so the QP used a soft boundary between all rock units. Sulphur was not estimated within the overburden.

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Table 14‑8: Comparison of Lithology and Average Sulphur Contents

Rock type	Count	Average S (%)

| Overburden^1^ (9) | 112 | 0.58 |

| Meta-sedimentary (50) | 498 | 2.69 |

| Volcanic (81) | 5,372 | 2.43 |

| Intrusive (500) | 5,789 | 1.95 |

| Sand Zone^2^ (301) | 2,720 | 3.09 |

| Basement (999) | 7,416 | 1.30 |

| Total | 21,907 | 2.00 |

Note:

1.	Not estimated.

| 2. | The Sand Zone is a structural domain that is central to the Portage zone and is part of the Volcanic rock domain. |

Both sulphur and arsenic were estimated in the model using inverse distance squared interpolation in four consecutive passes. Sulphur was estimated with 21,907 data points from the ICP assay database and arsenic was estimated from 13,835 data points using soft boundaries for all rock types.

14.10.4 Acid Base Accounting Model

In addition to the Total Sulphur and arsenic models, the QP estimated parameters to facilitate the determination of acid generation potential of the rock within the open pit. The model utilized acid potential (AP) and neutralization potential (NP) inputs to determine if the rock was PAG or NAG based on its neutralization potential ratio (NPR, NP/AP). At the time the model was prepared, AP inputs comprised data points based on ICP sulphur and Leco sulphur assays. NP inputs comprised 866 data points based on Sobek NP analyses from previous ML/ARD studies.

Wood plc (now WSP Canada Ltd.) developed an Acid Base Accounting (ABA) Analogue as part of ongoing geochemical studies for the Project. The ABA Analogue utilizes relationships between static testing data and the Project exploration ICP database, specifically AP (based on Leco sulphur), ICP sulphur and NP on ICP calcium, to support the assessment of mine rock ARD potential at greater spatial resolution than possible by traditional static testing approaches. Use of these relationships could allow for integration of an additional 21,040 AP and 12,968 NP data into the block model and improve the identification and delineation of PAG from NAG rock for Project planning purposes (Wood, 2022).

The QP utilized the AP and NP data provided by Wood to interpolate and determine the ABA of the material within the resource pit shell. The data provided contained 21,907 AP data points and 13,834 NP data points. Of these, 869 NP and 866 AP were measured and the remainder were analogue. Table 14‑9 summarizes the AP and NP data by rock types.

Table 14‑9: Average ABA data by Rock types

Code	Rock Type	Average AP	AP Count	Average NP	NP Count

| 9 | Overburden^1^ | 18.20 | 112 | 108.77 | 103 |

| 50 | Meta-sedimentary | 84.18 | 498 | 105.29 | 251 |

| 81 | Volcanic | 76.00 | 5,372 | 72.18 | 2,614 |

| 500 | Intrusive | 61.07 | 5,789 | 55.76 | 3,596 |

| 301 | Sand Zone^2^ | 96.68 | 2,720 | 18.73 | 1,484 |

| 999 | Basement | 40.50 | 7,416 | 99.99 | 5,786 |

| Total | | 62.49 | 21,907 | 74.68 | 13,834 |

Note:

1.	Not estimated.

| 2. | The Sand Zone is a structural domain that is central to the Portage zone and is part of the Volcanic rock domain. |

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The QP decided to estimate the AP and NP values using Ordinary Kriging in four separate passes and using soft boundaries for lithological types (Table 14‑10). All blocks were then coded as PAG if the NPR was less or equal to 2, and coded as NAG if the NPR was greater than 2.

Table 14‑10: Search Parameters for ABA Parameters

Item	Pass	Rotation			Search Ellipse Size			Number of Composites		Max. Samples per DDH

| | | Z | Y | Z | X (m) | Y (m) | Z (m) | Min. | Max. | |

| AP | 1 | -14 | 37 | 32 | 26 | 90 | 47 | 4 | 20 | 3 |

| AP | 2 | -14 | 37 | 32 | 168 | 204 | 350 | 4 | 20 | 3 |

| AP | 3 | -14 | 37 | 32 | 168 | 204 | 350 | 1 | 20 | 3 |

| AP | 4 | -14 | 3 | 32 | 350 | 350 | 350 | 1 | 20 | 3 |

| NP | 1 | 33 | -31 | 31 | 100 | 100 | 100 | 4 | 20 | 3 |

| NP | 2 | 33 | -31 | 31 | 160 | 200 | 160 | 4 | 20 | 3 |

| NP | 3 | 33 | -31 | 31 | 160 | 200 | 160 | 1 | 20 | 3 |

| NP | 4 | 33 | -31 | 31 | 350 | 350 | 350 | 1 | 20 | 3 |

14.11 Model Validation

The Springpole resource block model was validated by completing a series of visual inspections. It was additionally validated by comparing local “well-informed” block grades with composites contained within those blocks, and by comparing average assay grades with average block estimates along different directions – swath plots.

Figure 14‑8 shows a comparison of estimated gold block grades with borehole composite assay data contained within those blocks within the mineralized domains. Figure 14‑9 illustrates the same comparison for the silver grades. On average, the estimated blocks are similar to the composite data, although there is a large scatter of points around the x = y line. This scatter is typical of smoothed block estimates compared to the more variable assay data used to estimate those blocks. The thick white line that runs through the middle of the cloud is the result of a piece-wise linear regression smoother.

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Figure 14‑8: Comparison of Gold Grades for Well-Informed Blocks

Source: SRK, 2022

Figure 14‑9: Comparison of Silver Grades for Well-Informed Blocks

Source: SRK, 2022

Note that there are relatively few data points for silver for the East Extension (domain 100) and Camp zone (domain 200). This is due to the fact that only the Gold Canyon and First Mining drill holes had silver assay data for these two mineralized zones.

As a final check, average composite grades and average block estimates were compared along different directions. This involved calculating de-clustered average composite grades and comparing them with average block estimates along east-west, north-south, and horizontal swaths.

Figure 14‑10 shows the swath plots in the three mineralized zones, and Figure 14‑11 shows the swath plot for silver within the Portage zone. The average composite grades and the average estimated block grades are quite similar in all directions. Similar behaviour was documented for all other mineralized zones. Overall, the validation shows that the current resource estimate is a good reflection of drill hole composited data for the Springpole Gold Project.

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Figure 14‑10: Swath Plots for Gold re (a) the East Extension, (b) the Camp, and (c) the Portage Zone

(a)

| (b) |

| (c) |

Source: SRK, 2022

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Figure 14‑11: Swath Plot for Silver within the Portage Zone

Source: SRK, 2022

14.12 Mineral Resource Classification

Block model quantities and grade estimates for the Springpole Gold Project were classified by Dr. Gilles Arseneau, Ph.D., P.Geo. (APEGBC #23474), an independent Qualified Person for the purposes of NI 43-101. The classification was completed according to the current CIM Definition Standards, 2014.

Mineral resource classification is typically a subjective concept; industry best practices suggest that mineral resource classification should consider the confidence in the geological continuity of the mineralized structures, the quality and quantity of exploration data supporting the estimates, and the geostatistical confidence in the tonnage and grade estimates. Appropriate classification criteria should aim at integrating these concepts to delineate regular areas at similar resource classification.

The QP is satisfied the geological modelling honours the current geological information and knowledge. The location of the samples and the assay dataset is sufficiently reliable to support resource evaluation. The sampling information was acquired primarily by core drilling on sections spaced at 50 m.

The mineral resources were classified according to the following rules:

·	For the East Extension zone, any blocks estimated during Pass 1 or Pass 2 with at least two drill holes and six composites were classified as Indicated mineral resources. All other estimated blocks were classified as Inferred mineral resources.
·	The Portage and Camp classification was based solely on the gold estimate. Silver, as a minor by-product, carries the classification associated with the gold. Any blocks that were estimated during Pass 1 or Pass 2 with at least two drill holes and six composites were classified as Indicated mineral resources. All other interpolated blocks were classified as Inferred mineral resources.

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Most of the Indicated mineral resource is supported with at least two drill holes within an average distance of less than 50 m, and the Inferred resource is supported with a minimum of two holes with an average distance of less than 100 m (Figure 14‑12). Most of the Inferred resource is supported with at least five drill holes within 100 m and the Indicated resource is supported with two to five holes within a 50 m distance (Figure 14‑13). The nearest drill hole for most of the Indicated resource is less than 50 m away while the nearest drill hole for the Inferred resources is 50 to 75 m away (Figure 14‑14).

Figure 14‑12: Average Distance of Drill Holes by Resource Class

Source: SRK, 2025

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Figure 14‑13: Number of Drill Holes by Resource Class

Source: SRK, 2025

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Figure 14‑14: Distance of the Nearest Drill Hole by Resource Class

Source: SRK, 2025

14.13 Mineral Resource Statement

CIM Definition Standards for mineral resources and mineral reserves (May 2014) defines a mineral resource as:

“A Mineral Resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction. The location, quantity, grade or quality, continuity and other geological characteristics of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling”.

The “material of economic interest” refers to diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals.

The “reasonable prospects for economic extraction” requirement generally implies that the quantity and grade estimates meet certain economic thresholds and that the mineral resources are reported at an appropriate cut-off grade taking into account extraction scenarios and processing recoveries. In order to meet this requirement, the QP considers that major portions of the Project are amenable for open pit extraction.

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To determine the quantities of material offering “reasonable prospects for economic extraction” by an open pit, the QP used a pit optimizer and reasonable mining assumptions to evaluate the proportions of the block model (Indicated and Inferred blocks) that could be “reasonably expected” to be mined from an open pit.

The optimization parameters were selected based on a combination of derived project data and cost estimates, experience and benchmarking against similar projects (Table 14‑11). The reader is cautioned that the results from the pit optimization are used solely for the purpose of testing the “reasonable prospects for economic extraction” by an open pit and do not represent an attempt to estimate mineral reserves. The results are used as a guide to assist in the preparation of a Mineral Resource statement and to select an appropriate resource reporting cut-off grade.

The mineral resource statement for the Project is presented in Table 14‑12.

Table 14‑11: Assumptions Considered for Conceptual Open Pit Resource Optimization

Parameter	Units	Value

| Au Price | $/oz | 2450 |

| Ag Price | $/oz | 27.50 |

| Exchange Rate | $US/$CDN | 0.74 |

| Mining Cost | $/t mined | 2.30 (ore) |

| Processing | $/t of feed | 14.50 |

| General and Administrative | $/t of feed | 0.90 |

| Overall Pit Slope | Degrees | 35 to 45 based on domains |

| Au Process Recovery | Percent | 87.2 |

| Ag Process Recovery | Percent | 85.5 |

| In Situ COG | g/t | 0.20 Au |

The QP considers that the blocks located within the conceptual pit envelope show “reasonable prospects for economic extraction” and can be reported as a Mineral Resource (Table 14‑12).

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Table 14‑12: Mineral Resource Statement Inclusive of Mineral Reserves (effective September 30, 2025)

Category	Quantity<br> <br>(Mt)	Grade		Metal

| | | Au | Ag | Au | Ag |

| | | (g/t) | (g/t) | (Moz) | (Moz) |

| Indicated | 191 | 0.78 | 4.6 | 4.8 | 28 |

| Inferred | 64 | 0.38 | 3.1 | 0.8 | 6.5 |

Note:

1.	Mineral resources are reported in relation to a conceptual pit shell. Mineral resources are not mineral reserves and do not have demonstrated economic viability. All figures are rounded to reflect the relative accuracy of the estimate. All composites have been capped where appropriate

| 3. | Preliminary mining cost assumptions of C$2.60/tonne mined of waste, C$2.30/tonne mined of ore, and C$2.00/tonne mined of overburden, with an incremental mining cost of C$0.02/tonne/6m mined |

| 5. | Overall pit shell slope angles ranged from 20 - 45° |

This resource model includes mineralized material in the Camp, East Extension and Portage zones spanning 1,860 m in the southeast direction along the axis of the Portage zone and 900 m in the northeast direction perpendicular to the long axis of the Portage zone. Resource modelling includes mineralized material generally ranging from 340 m to 440 m below surface.

14.13.1 Note on Inferred Resources

An Inferred mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity. An Inferred mineral resource has a lower level of confidence than that applying to an Indicated mineral resource and must not be converted to a mineral reserve. It is reasonably expected that the majority of Inferred mineral resources could be upgraded to Indicated mineral resources with continued exploration.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.

The estimate of mineral resources may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues. The quantity and grade of reported Inferred mineral resources in this estimation are uncertain in nature and there has been insufficient exploration to potentially convert some or all of these Inferred mineral resources as an Indicated or Measured mineral resources and it is uncertain if further exploration will result in upgrading them to the Indicated or Measured mineral resource category. The QP is of the opinion that further attempts to convert the remaining Inferred material to indicated would be of questionable value. The current proportion of the resource classified as Inferred is about 25% of total tonnes, and 14% of contained gold. The mineral resources in this statement were estimated using the current CIM Definition Standards, 2014.

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14.14 Grade Sensitivity Analysis

The Mineral Resources of the Springpole Gold Project are variable depending upon the selected COG. To illustrate this sensitivity, the global block model quantities and grade estimates within the conceptual pit used to constrain the mineral resources are presented at different cut-off grades in Table 14‑13 for the Indicated mineral resource and in Table 14‑14 for the Inferred mineral resource. The reader is cautioned that the figures presented in this table should not be misconstrued with a mineral resource statement. The figures are only presented to show the sensitivity of the block model estimates to the selection of COG. Figure 14‑15 presents this sensitivity as grade tonnage curves for the Indicated mineral resource and Figure 14‑16 displays the same sensitivity curve for the Inferred mineral resource.

Table 14‑13: Indicated Block Model Quantities and Grade Estimates at Cut-off Grades

COG	Quantity	Grade	Grade

| 0.10 | 218 | 0.70 | 4.2 |

| 0.20 | 191 | 0.78 | 4.6 |

| 0.25 | 176 | 0.83 | 4.8 |

| 0.30 | 160 | 0.88 | 5.0 |

| 0.35 | 145 | 0.94 | 5.2 |

| 0.40 | 131 | 1.00 | 5.4 |

| 0.50 | 107 | 1.12 | 5.9 |

| 0.60 | 87 | 1.25 | 6.2 |

| 0.70 | 70 | 1.39 | 6.6 |

| 0.80 | 58 | 1.54 | 6.9 |

Note:

1.	The reader is cautioned that the figures in this table should not be misconstrued with a mineral resource statement. The figures are only presented to show the sensitivity of the block model estimates to the selection of COG. Mineral resource base case is highlighted in grey.

Table 14‑14: Inferred Block Model Quantities and Grade Estimates at Cut-off Grades

COG	Quantity	Grade	Grade

| 0.10 | 80 | 0.34 | 2.9 |

| 0.20 | 64 | 0.38 | 3.1 |

| 0.25 | 53 | 0.42 | 3.3 |

| 0.30 | 41 | 0.46 | 3.6 |

| 0.35 | 31 | 0.51 | 3.9 |

| 0.40 | 22 | 0.56 | 4.1 |

| 0.50 | 11 | 0.66 | 4.1 |

| 0.60 | 6 | 0.78 | 4.3 |

| 0.70 | 3 | 0.90 | 4.6 |

| 0.80 | 2 | 1.03 | 5.2 |

Note:

1.	The reader is cautioned that the figures in this table should not be misconstrued with a mineral resource statement. The figures are only presented to show the sensitivity of the block model estimates to the selection of COG. Mineral resource base case is highlighted in grey.

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Figure 14‑15: Grade-Tonnage Curves for the Springpole Indicated Mineral Resource

Source: SRK, 2025

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Figure 14‑16: Grade-Tonnage Curves for the Springpole Inferred Mineral Resource

Source: SRK, 2025

14.15 Previous Mineral Resource Estimates

Mineral resources for the Springpole Gold Project were previously estimated and reported in a technical report filed on February 26, 2021 (AGP, 2021). The effective date of the mineral resources in that technical report was July 30, 2020. The 2020 mineral resource was reported at a lower cut-off (0.30 g/t as opposed to 0.20 g/t for the current estimate). The 2020 estimate used different optimization parameters and cut-off to restrict the mineral resource within an open pit resource shell. The mineral resources were reported in accordance with NI 43-101 and are summarized in Table 14‑15. These mineral resources are no longer current and are now replaced by the mineral resources presented in Table 14‑13 and Table 14‑14 of this report.

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Table 14‑15: Previous Mineral Resource Statement of July 30, 2020 (at 0.30 g/t Cut-off)

Classification	Tonnage (Mt)	Au<br> <br>(g/t)	Ag<br> <br>(g/t)	Au Contained<br> <br>(Moz)	Ag Contained<br> <br>(Moz)

| Indicated | 151 | 0.94 | 5.0 | 4.6 | 24.3 |

| Inferred | 16 | 0.54 | 2.8 | 0.3 | 1.4 |

To better compare the current mineral resource with the previous resource estimate, the QP reported the current mineral resource at the same cut-off (0.30 g/t) used in 2020 (Table 14.16). As can be seen, at this cut-off the current resource is similar to the previous mineral resource statement with less than a 6% increase in the Indicated tonnes and a 2% increase in the contained Indicated ounces within the resource shell. The Inferred tonnes have increased by about 156% due mainly to the expansion of the Portage zone as a result of the inclusion of data from subsequent drill programs, which resulted in an increase of about 100% of the contained Inferred gold ounces.

Table 14‑16: Current (2025) Mineral Resource at 0.30 g/t Cut-off

Classification	Tonnage (Mt)	Au<br> <br>(g/t)	Ag<br> <br>(g/t)	Au Contained<br> <br>(Moz)	Ag Contained<br> <br>(Moz)

| Indicated | 160 | 0.88 | 5.0 | 4.7 | 25.7 |

| Inferred | 41 | 0.46 | 3.6 | 0.6 | 4.7 |

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15 MINERAL RESERVE ESTIMATES

15.1 Summary

The mineral reserve estimate for the Project is based on the conversion of the Indicated mineral resources within the current technical report mine plan. Indicated mineral resources in the mine plan were converted directly to Probable mineral reserves. There are currently no Measured mineral resource estimates for the Project and therefore there are no Proven mineral reserves. The estimated mineral reserves for the Project are shown in Table 15‑1.

Table 15‑1: Springpole Mineral Reserve Estimate – November 13, 2025

Category	Cut off Grade<br> <br>(g/t Au)	Tonnes<br> <br>(Mt)	Grade Au<br> <br>(g/t)	Grade Ag<br> <br>(g/t)	Contained Ounces Au<br> <br>(M oz)	Contained Ounces Ag<br> <br>(M oz)

| Proven | 0.27 | - | - | - | - | - |

| Probable | 0.27 | 102.0 | 0.94 | 4.90 | 3.1 | 16.1 |

| Total | 0.27 | 102.0 | 0.94 | 4.90 | 3.1 | 16.1 |

Note:

1.	This Mineral Reserve estimate is as of November 13, 2025, and is based on the new mineral resource estimate dated September 30,2025

| 2. | The Mineral Reserve estimation was completed under the supervision of Gordon Zurowski, P.Eng of AGP Mining Consultants Inc., who is a Qualified Person as defined under NI 43-101 |

| 3. | Mineral Reserves are stated within the ultimate design pit based on: |

o	US$2100/ounce gold price, US$24/ounce silver price

| o | Pit Limit corresponds to a pit shell with a revenue factor of 0.60, corresponding to a US$1,260 /ounce gold price and US$14.40/oz silver |

| o | A cut-off grade of 0.27 g/t Au for all pit phases. |

| o | Preliminary mining cost assumptions of C$2.60/tonne mined of waste, C$2.30/tonne mined of ore, and C$2.00/tonne mined of overburden, with an incremental mining cost of C$0.02/tonne/6 m mined. |

| o | Preliminary processing cost assumptions of C$14.50/tonne processed, general and administration assumption of C$0.90/tonne processed, stockpile cost assumption of C$0.75/tonne processed, and incremental ore mining cost of C$0.56/tonne processed |

| o | Preliminary process recovery assumptions of 87.2% for gold and 85.5% for silver |

| o | An exchange rate of C$1.35 equal to US$1.00 |

| o | The preliminary economic, cost and recovery assumptions used at the time of mine planning and reserve estimation may not necessarily conform to those stated in the economic model |

4.	Pit slope inter-ramp slope angle assumptions ranged from 22 - 54°

At the time of preparing the Mineral Reserve Estimate, the QP has not identified any known legal, political, environmental, or other risks that would materially affect the potential development of the Mineral Reserves. The risk of not being able to secure the necessary permits from the government for development and operation of the Project exists but the QP is not aware of any issues that would prevent those permits from being approved per the normal permitting process.

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15.2 Introduction

NI 43-101 defines the terms “mineral reserve”, “Probable mineral reserve” and “Proven mineral reserve” have the meanings ascribed to those terms by the Canadian Institute of Mining, Metallurgy, and Petroleum, as the CIM Definition Standards, 2014 and the CIM Best Practice Guidelines, 2019.

A mineral reserve is the economically mineable part of a measured and/or Indicated mineral resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at Pre-Feasibility or Feasibility level as appropriate that include application of modifying factors. Such studies demonstrate that, at the time of reporting, extraction could reasonably be justified.

Modifying factors are considerations used to convert mineral resources to mineral reserves. These include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social, and governmental factors.

A Probable mineral reserve is the economically mineable part of an Indicated, and in some circumstances, a Measured mineral resource. The confidence in the modifying factors applying to a Probable mineral reserve is lower than that applying to a Proven mineral reserve. A Proven mineral reserve is the economically mineable part of a Measured mineral resource. A Proven mineral reserve implies a high degree of confidence in the modifying factors. Application of the Proven mineral reserve category implies that the qualified person has the highest degree of confidence in the estimate with the consequent expectation in the minds of the readers of the report.

15.3 Mineral Reserves Statement

The mineral reserves for the Springpole Gold project are based on the conversion of the Indicated mineral resources within the current technical report mine plan. Indicated mineral resources in the mine plan were converted directly to Probable mineral reserves. There are currently no Measured mineral resource estimates and therefore there are no Proven mineral reserves.

Table 15‑2 presents the mineral reserves inside the design pit. Note numbers have been rounded.

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Table 15‑2: Proven and Probable Mineral Reserves – Springpole Project (November 13, 2025)

Category	Cut off Grade<br> <br>(g/t Au)	Tonnes<br> <br>(Mt)	Grade Au<br> <br>(g/t)	Grade Ag<br> <br>(g/t)	Contained Ounces Au<br> <br>(M oz)	Contained Ounces Ag<br> <br>(M oz)

| Proven | 0.27 | - | - | - | - | - |

| Probable | 0.27 | 102.0 | 0.94 | 4.90 | 3.1 | 16.1 |

| Total | 0.27 | 102.0 | 0.94 | 4.90 | 3.1 | 16.1 |

Note:

1.	This mineral reserve estimate is as of November 13, 2025, and is based on the new mineral resource estimate dated September 30,2025

| 2. | The mineral reserve estimation was completed under the supervision of Gordon Zurowski, P.Eng of AGP Mining Consultants Inc., who is a qualified person as defined under NI 43-101 |

| 3. | Mineral reserves are stated within the ultimate design pit based on: |

o	US$2100/ounce gold price, US$24/ounce silver price

| o | Pit Limit corresponds to a pit shell with a revenue factor of 0.60, corresponding to a US$1,260 /ounce gold price and US$14.40/oz silver |

| o | A cut-off grade of 0.27 g/t Au for all pit phases. |

| o | Preliminary mining cost assumptions of C$2.60/tonne mined of waste, C$2.30/tonne mined of ore, and C$2.00/tonne mined of overburden, with an incremental mining cost of C$0.02/tonne/6m mined. |

| o | Preliminary processing cost assumptions of C$14.50/tonne processed, general & administration assumption of C$0.90/tonne processed, stockpile cost assumption of C$0.75/tonne processed, and incremental ore mining cost of C$0.56/tonne processed |

| o | Preliminary process recovery assumptions of 87.2% for gold and 85.5% for silver |

| o | An exchange rate of C$1.35 equal to US$1.00 |

| o | The preliminary economic, cost and recovery assumptions used at the time of mine planning and reserve estimation may not necessarily conform to those stated in the economic model |

4.	Pit slope inter-ramp slope angle assumptions ranged from 22 - 54°

15.4 Estimation Procedure

The following summarizes the general procedure followed to estimate the mineral reserve estimate. The details of the steps are further described in Section 16.

·	Mining block model was created by incorporating dilution and loss assumptions into the resource block model
·	Preliminary cost and recovery assumptions were developed to initiate cut-off grade estimations
·	Pit limit analysis was undertaken using a prescribed set of parameters to estimate economic pit limits for the deposit
·	Detailed pit and phase designs were completed incorporating bench slope designs and haul ramps
·	LOM scheduling was completed with the objective of providing 30,000 tonne/ day or 10,950,000 tonnes/year to the process plant
·	Estimation of mine equipment fleet and workforce requirements
·	Estimation of open pit mining operating and capital costs based on LOM mine plan

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15.5 Factors that May Affect the Mineral Reserve Estimates

Factors that may affect the mineral reserve estimate include:

·	Commodity prices and currency exchange rates
·	Interpretations of mineralization geometry and continuity of mineralization zones
·	Geomechanical and hydrogeological assumptions
·	Ability of the mining operation to meet the annual production rate
·	Operating cost assumptions
·	Process plant recoveries
·	Mining loss and dilution
·	Ability to meet and maintain permitting and environmental license conditions

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16 MINING METHODS

16.1 Summary

The PFS mine plan is based on open pit mining. With current metal pricing levels, knowledge of the mineralization, grade tenor, grade distribution and proximity to surface, open pit mining offers the most reasonable approach for development.

A single pit will provide the open pit feed material necessary to maintain the process plant feed rate at 30,000 t/d while operational. The pit is proposed as a three-phase design using 12 m benches which provides 102.0 Mt of mill feed grading 0.94 g/t Au, and 4.90 g/t Ag. Waste from this pit will total 309.5 Mt for a strip ratio of 3.0 (waste:mill feed).

The mill feed cut-off used is 0.25 g/t AuEQ, accounting for the contribution of both gold and silver. This is equivalent to a gold only cut-off of 0.27 g/t Au. The factors contributing to the calculation of Au equivalent grade AuEQ are discussed in Section 16.5.2.1 Cut-off Grade. During the mine operation material would be stockpiled to optimize the plant feed grade and defer lower-grade material until later in the mine schedule. Three grade bins are used for the stockpiles including: low grade (0.25 - 0.50 g/t AuEQ), medium grade (0.50 – 0.75 g/t AuEQ) and high grade (+0.75 g/t AuEQ).

The phases are scheduled to provide 30,000 t/d of feed to the mill over a 9.4 year mining life after one year of pre-production stripping. The last 1.4 years of mine life are stockpile reclaim. The pits are sequenced to minimize initial stripping and provide higher feed grades in the early years of the mine life which the stockpiling strategy accomplishes.

The main fleet will consist of up to four 251 mm rotary drills and two 140 mm drills, two 37 m^3^ electric hydraulic shovels and three 23 m^3^ front end loaders. The truck fleet will total 25 – 240-tonne trucks at the peak of mining. This is due to the long hauls from the pit to the CDF. The usual assortment of dozers, graders, small backhoes, and other support equipment is considered in the equipment cost estimate. A smaller front-end loader (13 m^3^) will be stationed at the primary crusher.

Waste material from the pit will be stored in the CDF. NAG/non-metal leaching material will be used for the outer Embankments while PAG material will be co-disposed with process plant tailings. The majority of the NAG waste rock material from the open pit will be contained within the CDF (134.6 Mt), but a small portion of NAG/non-metal leaching material will be backfilled into Phase 2 near the end of the mine life. This reduces the overall haul length and helps in reclamation of the pit. An estimated total of 15.4 Mt will be backfilled into the pit.

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In addition to the open pit, two quarries will be established during the pre-production period and at closure. These quarries will be used to provide rock material for various mine infrastructure including haul roads, dikes, CDF Embankments, and to meet site fill requirements for other infrastructure. Approximately 37.4 Mt of material is planned to be excavated from the CDF quarry in the preproduction period, while 26.2 Mt of material is planned from the fish habitat development area at the end of the mine life.

When the open pit is complete, the larger mining fleet will move to complete the fish habitat development area. Material will be used as cover for the CDF cells and will be dumped into the open pit. This serves to cover the slopes in the pit for reclamation purposes and upon closure the quarry will become additional lake area which will be contoured to provide suitable fish habitats.

16.2 Geotechnical and Hydrogeological Considerations used in Mine Planning

16.2.1 Geotechnical Considerations

A Feasibility Study (FS) level field investigation was designed and implemented between February and April 2022 at the Springpole Project. The drill hole design was based on the PFS pit shape and understanding of the rock mass condition, as illustrated in Figure 16‑1.

The 2022 drilling and characterization program discovered a southern extension of sandy material which was previously only known in the ‘centre’ of the mineralization towards the north-central part of the PFS pit shape. The material within this extended ‘Sand-zone’, which is in the most altered part of the Low-RQD domain, can be described, using the sample collected from drill hole SG22-029 (Figure 16‑2), as being:

·	well graded sand with silt and gravel (SW-SM) using the Particle Size Distribution of the “combined sample” (Figure 16‑3);
·	the material is medium to fine-grained with non-plastic fines, with the fine content being 13% from the sample analysed;
·	the shear strength tests (undrained triaxial tests) on reconstituted samples yielded an estimated friction angle of φ ≈ 32 to 34°, with a potential upper bound of φ ≈ 38°; and
·	the material genesis (rock exposed to thermo-chemical processes) may support the presence of angular sands at depth.

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Figure 16‑1: Oblique View Looking Eastward showing PFS Pit and Modelled Sand-zone

Source: SRK, 2025

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Figure 16‑2: Material Sampled for 2022 Characterization of Sand-zone

Source: SRK, 2025

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Figure 16‑3: (SW-SM) Well Graded Sand with Silt and Gravel

Source: SRK, 2025

It has been recognized for more than a decade that there are very weak pockets of ‘mush’ at Springpole and that these would have an impact on the slope designs, especially on the central and southern side of the pit. Very low RQD (Rock Quality Designation) material was considered synonymous with the ‘mush’ during the Scoping studies and Preliminary Economic Assessment (PEA). This very low RQD material, and those drill runs which have zero Total Core Recovery (TCR), are sometimes referred to as an “Unconsolidated Granular Material” (UGM) within the context of drilling and core recovery.

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Early attempts to characterize the granular material were difficult because of the variability in the way in which the site geologists logged TCR and RQD, and the challenge of understanding the interconnectedness of the ‘pockets’ of lost core. To better understand the amount of weak material, a preliminary Low RQD model was built by SRK in 2021, which was used for drill hole planning. Verification of the extent of the low RQD volume was attempted in 2022, and it was soon realized that there is a ‘core’ of sand and granular material in the middle of the Low RQD rock mass volume that could be mapped in 3D. Core box photo re-logging was done to confirm the presence of the UGM, and a new 3D model was built. These zones of granular material below the proposed open pit shape have been labelled as the “Sand-zone”, because of their grain size.

The Sand-zone is modelled on a combination of parameters: RQD = 0, TCR < 60, and ‘sand’ interpreted in photo logging. The Sand-zone model has two possible sub-domains:

·	A steep zone through the main body of the deposit, probably associated with protolith fabric.
·	A shallow northward-dipping zone, or ‘sheet’, to the south of the pit, possibly associated with a property-wide basal fault zone.

Differentiation between these two ‘types’ of sand has not been attempted yet.

Two rock geotechnical drilling programs were implemented on the Springpole property by SRK, one in 2012/2013 and a second in 2022. Oriented core data acquisition has focused on rock mass characterization. Triple-tube oriented core joint data have been verified using an independently acquired and interpreted Acoustic Televiewer (ATV) data source.

Full waveform sonic (FWS) data have not yet been acquired to verify the Sand-zone character because of hole stability challenges. Some discing has been noted below the Sand-zone along the main axis of the PFS pit shape, but this doesn’t appear to be influencing the rock mass model’s lower boundary. Intact rock strength data has been acquired, and the joint conditions have been described at an FS level of confidence.

Conditions within the Sand-zone, however, have not been adequately studied yet – and the interpreted zone of increased geotechnical risk associated with this. The material within the Sand-zone, which is in the most altered part of the Low-RQD domain was characterized at a Scoping to PFS level of confidence.

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Figure 16‑4: Vertical Section of Drill Hole Locations and Zone of Increased Geotechnical Risk

Source: SRK, 2025

The material within the Sand-zone was observed to be loose even at depths below 150 m. Small bulk samples were collected by First Mining from drill hole SG22-029 at a depth interval between 158 m and 169 m. The sub-samples were combined into a composite sample for laboratory testing.

Physical characterization (particle-size distribution test) shows a composition of 30% gravel, 57% sand, and 13% non-plastic fines (% passing sieve #200). The sampled material was described as well-graded sand with silt and gravel, using the Unified Soil Classification System (USCS). The sand is medium to fine-grained with few coarse particles. The gravel particles of the grab samples have sub-angular shape and a maximum particle size between 25 and 54 mm.

The presence of interpreted loose (medium to fine) sands at depths below 150 m (and at the expected in-situ stresses) is not consistent with typical transported sand. A different material genesis for this unit, such as rock weathered through thermo-chemical processes, may support the presence of the loose angular sands at such a large depth below surface.

Mechanical characterization included three triaxial consolidated undrained (CU) shear strength tests using the sand fraction of the sample (material smaller than 4.75 mm). Samples were re-constituted loose (void ratio values between 0.87 and 0.93), compressed under three different stresses (93 kPa, 192 kPa, 378 kPa) and saturated (B-value 0.96 to 0.99). The testing stress range is considerably smaller than the stresses expected at the depth from which the material was retrieved, but it is limited by the laboratory’s equipment capabilities.

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The triaxial (CU) test results are indicative of a contractive loose sample that generates pore water pressure during shear. The photos of the sand sample taken at the end of the test show a ‘barrel compression shape’, in other words, a middle zone bulged with the ends constrained by the platens and no distinct shear plane – which supports the interpretation of a loose contractive sample. The void ratio values reported by the lab at the end of the compression phase (the start of the shear phase), ranging between 0.60 and 0.54, are inconsistent with this interpretation of loose sand. Future work should focus on the determination of the void ratio at the end of the compression phase of testing.

Interpretation of the triaxial (CU) test results indicated a critical friction angle of lab = 38°. This friction angle is representative of the laboratory-scale specimen and considered to be too high for a medium-fine sand (sub-angular or rounded). The laboratory value is probably more representative of an angular material. The validation of this critical friction angle for all the Sand-zone, and correlation with the physical properties, requires further testing and interpretation as part of the future study work.

Based on this interim characterization of the Sand-zone, a friction angle of = 34° was used in the interim slope stability design and stability assessment for mining optimization studies. Table 16‑1 tabulates the parameters considered for bench design and overall slope angles, while Figure 16‑5 shows the geotechnical sectors for the Project.

Table 16‑1: Overall Slope Angle Estimation

Slope Design		Preshear Bench		Catch-width	Stack Height	Geotechnical<br> <br>Berm	Maximum

| 1 | FW Rock | 70 | 2 x 12 | 10 | 120 | 20 | 52 | 40 |

| 2 | HW Rock – Upper | 75 | 2 x 12 | 11 | 120 | 20 | 54 | 45 |

| 2 | HW Rock – Lower | 75 | 1 x 12 | 7 | 120 | 20 | 50 | 45 |

| 3 | HW Rock | 65 | 1 x 12 | 7 | 60 | 20 | 44 | 40 |

| 3 | FW Rock | 65 | 1 x 12 | 7 | 60 | 20 | 44 | 40 |

| All | Low – RQD | 65 | 1 x 12 | 10 | 60 | 20 | 38 | 35 |

| All | Sand Zone | ~34 | 1 x 6 | 6 | 30 | 12 | 22 | 20 |

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Figure 16‑5: Geotechnical Sectors

Source: SRK, 2022

16.2.2 Hydrogeological Considerations

No additional testwork or information related to open pit stability is available since the 2021 Pre-Feasbility Study. The following subsection summarizes the hydrogeological considerations considered for the detailed pit design, taken from the 2021 Pre-Feasibilty Study.

Per SRK (2019), during operations, inflow rates will be a function of pit-development shape, volume, and rate of excavation, as well as hydraulic conductivity of the bedrock and the hydraulic gradient between the mine and surrounding surface water sources (lakes). Estimates of potential inflow rates to the pit were made at an “order of magnitude” accuracy level using analytical methods (Dupuit, 1863), and conceptual 2D and 3D groundwater flow simulations.

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Hydraulic conductivity (K) values were derived from preliminary testwork. Given the location of the pit relative to the surrounding lakes, and the limited information on geology and structure currently available, a conservative approach was taken in estimating the pit-inflow. Inflow rates were estimated to reach a maximum of approximately 17,000 m^3^/d. The potential for geological features to connect to the lakes was flagged as a risk to the Project in the form of unanticipated water management and environmental concerns.

16.3 Geologic Model Importation

The resource model for the 2025 PFS was provided by SRK. That model has an effective date of September 30, 2025 and is described in detail in Section 14. This model was imported to MinePlan and Deswik for mine planning work.

SRK provided AGP with resource models in CSV format for open pit mine planning. The type of block model was a single mineralization percentage model. The grades in each block of the resource model is considered to be an undiluted grade.

Framework details of the different open pit block models are provided in Table 16‑2. Resource model item descriptions are shown in Table 16‑3 while the open pit mine model items are displayed in Table 16‑4. The mining model created by AGP includes additional items for mine planning purposes. Hexagon’s MinePlan® software was used for the mining portion of the PFS, and Deswik was used for their pit limit analysis routine. The pit, WRSF design and mine scheduling tools of Hexagon were used for the PFS.

The mine plan is based on Indicated mineral resources, as no Measured mineral resources are contained in the resource model. The block SG values provided in the resource model were estimated based on provided density data, with waste blocks without values receiving a default value of 2.74 t/m^3^.

LiDAR contours and updated bathymetry data were imported into MinePlan and then merged to create an original ground topography surface for the PFS.

Table 16‑2: Open Pit Model Framework

Framework Description	Resource Model Value	Mine Planning Model Value (MinePlan)	Mine Planning Model Value (Deswik)

| Block Model Name | 2025_Model.csv | - | M116.dm |

| MineSight^®^ file 10 (control file) | - | M11610.dat | - |

| MineSight^®^ file 15 (model file) | - | M11615.mp3 | - |

| X origin (m) | 548,500 | 548,500 | 548,500 |

| Y origin (m) | 5,692,400 | 5,692,400 | 5,692,400 |

| Z origin (m) (max) | -122 | -122 | -122 |

| Rotation (degrees clockwise) | 0 | 0 | 0 |

| Number of blocks in X direction | 220 | 220 | 220 |

| Number of blocks in Y direction | 210 | 210 | 210 |

| Number of blocks in Z direction | 90 | 90 | 90 |

| X block size (m) | 10 | 10 | 10 |

| Y block size (m) | 10 | 10 | 10 |

| Z block size (m) | 6 | 6 | 6 |

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Table 16‑3: Resource Model Item Descriptions

Field Name	Min	Max	Precision	Units	Comments

| DENSITY | 0 | 3.267 | 0.001 | t/m^3^ | Rock density |

| PERCENT | 0 | 100 | 0.001 | % | Percent of the block in mineralized zone (0 to 100) |

| AU | -1.066 | 40.063 | 0.001 | g/t | Gold grade in g/t uncapped |

| AUCAP | -1.066 | 40.063 | 0.001 | g/t | Capped gold in g/t (used for resource estimate) |

| AG | 0 | 144.7 | 0.001 | g/t | Silver grade in g/t |

| AGCAP | 0 | 144.7 | 0.001 | g/t | Capped silver grade in g/t (used for resource estimate) |

| CLASS_FINAL | 0 | 3 | 1 | - | Resource Class 1= Measured; 2=Indicated; 3= Inferred |

| ZONE | 0 | 301 | 1 | - | Mineralized zone |

Table 16‑4: Open Pit Model Item Descriptions

Field Name	Min	Max	Precision	Units	Comments

| DENS | 0 | 3.267 | 0.001 | t/m^3^ | Rock density |

| AUDIL | 0 | 39.760 | 0.001 | g/t | Diluted Gold grade in g/t |

| AGDIL | 0 | 144.702 | 0.001 | g/t | Diluted Silver grade in g/t |

| CLASS | 0 | 3 | 1 | - | Resource Class 1= Measured; 2=Indicated; 3= Inferred |

| ZONE | 0 | 301 | 1 | - | Mineralized zone 0=Air; 1=Water; 9=Overburden; 99=Waste; 100=East Extension Zone; 200=Camp Zone; 300=Portage Zone; 301=Portage High Grade |

| AUEQ | 0 | 34.859 | 0.001 | g/t | Gold Equivalent Grade |

| SULF | 0 | 16.606 | 0.001 | - | Sulphur as sulphide % |

| MLARD | 0 | 20 | 1 | - | Metal Leaching Acid Rock Drainage Classification 0=Type 0; 11=Type 1A; 12=Type 1B; 13=Type 1C; 20=Type 2 |

| RTYPE | 9 | 999 | 1 | - | Rock Type (9=overburden,50=Metasediment,81=Volcanic, 500 = Intrusive, 999 = Basement) |

| ROYT | 0 | 3 | 1 | - | Royalty Label 0=None; 1=Legacy; 2=Jubilee; 3=Springpole Group |

| PMF | 0 | 1 | 1 | - | Potential Mill Feed Label 1= Zone 100, 200, 300, 301 (Mineralized Zones) & Class = Indicated 0 = Any Zone & Class = 0, Inferred |

| MAT | - | - | - | - | Material Label combines PMF and Zone |

| SLOPE | 0 | 7 | 1 | - | Pit Slope Sector Label 1=Sector 1; 2=Sector 2 Upper; 3=Sector 2 Lower; 4=Sector 3 HW; 5=Sector 3 FW; 6=Low RQD; 7=Sand Zone |

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The ML/ARD Classification and sulphur attribute values were imported into the mining model based on values provided by WSP in the model file “In_pit_only_blocks_WSP DRAFT_27March2023.csv”. Rock type attribute values were assigned based on solids created from previous model since there was no change to the interpretations. Royalty attribute values were assigned based on solids provided by First Mining.

16.4 Mining Loss and Dilution

The mineral resources are based on a 10 x 10 x 6 m resource model block size. The resource model is a partial percent model, whereby the model captures the proportional make-up of each block—how much of the block is mineralized.

To create the mining model, which is assumed to incorporate mining dilution and loss, the resource model was regularized across the parent block. The mining unit size of 10 x 10 x 6 m, which accounts for planned open pit mine operating conditions, was selected to align with mining equipment, selectivity, and pit configuration. Mining would be completed on 12 m lifts for waste and 6 m lifts for mill feed material if required and the equipment selected is capable of mining in that manner.

In regularizing the resource model to a diluted mining model, the following approach was taken:

·	Grade attributes (Au, Ag) were estimated by calculating the metal content
·	Au Contained Metal (g) = Block Tonnage (t) x Percent Attribute (%) x Au Grade (g/t)
·	Ag Contained Metal (g) = Block Tonnage (t) x Percent Attribute (%) x Ag Grade (g/t)
·	The Contained Metal was divided by the Block Tonnage to calculate the diluted block grade
·	Au Diluted (g/t) = Au Contained Metal (g)/Block Tonnage (t)
·	Ag Diluted (g/t) = Ag Contained Metal (g)/Block Tonnage (t)
·	To quantify the impact of the regularization, the following approach was taken:
·	A solid was created on the Indicated resource blocks of mineralization at the AuEQ cut-off grade of 0.22 g/t from the mining model
·	A preliminary pit design and project topography surface were used as boundary constraints for final reporting
·	The mining model solid was interrogated against the resource model to estimate the amount of dilution and mining recovery incorporated into the mining model as a result of regularizing using the following formulas:
·

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·
·

This results in approximately 2.4% at 0 g/t Au of dilution and 2.2% mining loss.

16.5 Pit Limit Analysis

In the fall of 2022, AGP was retained by First Mining to prepare several internal mine planning scenarios, that examined combinations of phase designs, ultimate pit designs, and mining and milling production rates. Ultimate Pit and Phase designs prepared during this period were carried forward into this current study. The previous work was verified with the updated resource model.

16.5.1 Methodology

In accordance with the guidelines of the NI 43-101 on Standards of Disclosure for Mineral Projects, and the Canadian Institute of Mine, Metallurgy and Petroleum Definition Standards for Mineral Resources and Mineral Reserves, only those mineral blocks classified in the Measured and Indicated resource categories are allowed to drive the pit optimizer for a Pre-Feasibility or Feasibility Study. No economic value is attributed to Inferred blocks and, as such, these blocks are treated as waste blocks by the pit optimization routine.

Pit limit analysis was carried out using the Pseudoflow 3D algorithm in the Deswik mine planning software. The pit optimization algorithm is used to produce pit shells that are physical representations of an economical pit shell, assuming a given set of parameters and 3D block model. Using a variety of input parameters such as mining costs, processing costs, process recovery values and pit slopes, the algorithm outputs the pit shell that maximizes the undiscounted value of the deposit. These shells are devoid of geotechnical and operational features such as ramps, proper benching arrangements, and minimum mining width considerations. The pit shell’s purpose is to be used as:

·	a limit for reporting mineral resources
·	a basis for establishing pit limits and guide for the design of an engineered open pit

No capital expenses, such as those required for initial equipment purchase or waste pile construction, are considered by the pit optimization tool.

A series of pit shells are produced using a range of revenue factors (reduction factors on selling price) from 10 % to 100 % to produce the industry standard pit-by-pit graph. The revenue factor is used to measure the sensitivity of the pit optimizations to changes in mineral selling prices, as well as to evaluate the effect of the pit size and stripping ratios on the project present value (PV). The analysis produces a series of nested pit shells that prioritizes the mining of the most economic material and progressively increase in size, while less profitable material is mined as the revenue factor increases. The pit optimizations results are subsequently compared based on the estimated PV and calculated undiscounted value and tonnes of potential mill feed material and waste material. From these results, a final pit shell that meets project requirements and maximizes project PV is selected.

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16.5.2 Pit Limit Analysis Inputs and Parameters

The major costs and other parameters used for the pit limit analysis runs are detailed in Table 16‑5.

Table 16‑5: Pit Limit Analysis Parameters

Parameter	Unit	Value	Source

| Net Price Calculation – Au | CA$/g FOBMine | 87.76 | -- |

| Net Price Calculation – Ag | CA$/g FOBMine | 0.97 | -- |

| Currency | $ | CA dollars | -- |

| Ore Mining | $/t mined | 2.30 | AGP |

| Waste Mining | $/t mined | 2.60 | AGP |

| Overburden Mining | $/t mined | 2.00 | AGP |

| Incremental Mining Cost | $/t mined per 6m vertical | 0.020 | AGP |

| G&A | $/t milled | 0.90 | Ausenco |

| Stockpile Rehandle | $/t milled | 0.75 | AGP |

| Incremental Ore Mining Costs | $/t milled | 0.56 | AGP |

| Au Payable | % | 99.5 | First Mining |

| Ag Payable | % | 98.0 | First Mining |

| Royalties – Au | % | 3 | First Mining |

| Royalties - Ag | % | 3 | First Mining |

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Parameter	Unit	Value	Source

| Metallurgical Recovery - Au | % | 87.2 | Ausenco |

| Metallurgical Recovery - Ag | % | 85.5 | Ausenco |

| NSR | $/tonne milled | 16.71 | Calculated |

| AuEQ | g/t | 0.22 | Calculated |

| Lease Area Boundary | - | N/A | - |

| Physical Constraints on pit expansion | - | N/A | - |

16.5.2.1 Cut-off Grade

To classify the material contained within the open pit limits as material for processing or material for waste, the milling cut-off grade is used. This break-even cut-off grade is calculated to cover the costs of processing, general and administrative costs, and selling costs using the economic and technical parameters listed in Table 16‑5. Mineral resource material contained within the pit shell and above the cut-off grade is classified as potential mill feed (PMF), while resource material below the cut-off grade is classified as waste.

The cut-off grade for the pit limit analysis takes into account the value contribution of both gold and silver.

The cut-off grade or value can be represented by a single grade, a NSR type value, or by a metal equivalent grade value. The break-even Cut-off Value, based on the parameters in Table 16‑5 is:

·	NSR Value represented by the Processing Cost of $16.71/tonne milled
·	Au Equivalent, AuEQ grade of 0.22 g/t AuEQ

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Figure 16‑6: Cut-off Grade Analysis

Source: AGP, 2025

The calculation for the Au Equivalent grade is based on the following values and equation shown in Table 16‑6.

Table 16‑6: Parameters for Au Equivalent Grade

Metal	Metal Price	Metallurgical	Recovered	Gross Revenue

| Gold | 2729.65 | 87.2% | 0.87 | 2380.26 |

| Silver | 30.14 | 85.5% | 0.86 | 25.77 |

All values are in US Dollars.

For mine design and planning purposes, an elevated cut-off grade of 0.25 g/t AuEQ was selected. This AuEQ cut-off grade is approximately equal to a 0.27 g/t Au cut-off grade.

The basis for the elevated cut-off grade selection was the 2022/2023 internal mine planning work AGP undertook for First Mining. Figure 16‑7 illustrates the grade-tonnage curve, for the mining block model and the preferred ultimate pit design from the 2022/2023 mine planning work.

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Figure 16‑7: Grade-Tonnage Curve, Mining Block Model

Source: AGP, 2025

16.5.3 Pit Limit Analysis Results

The pit limit analysis process results in a series of nested pit shells, each corresponding to a Revenue Factor (RF). The revenue factor scales the metal prices only, and no costs are factored by the RF. The RF 1 corresponds to the selling prices listed in Table 16‑5.

Table 16‑7 and Figure 16‑8 summarize the nested pit shell results for the Springpole deposit at a selection of revenue factors.

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Three lines or categories of the discounted cash flow (DCF) are represented on the graph and table:

·	The Best Case DCF consists of mining out nested Pit Shell 1, the smallest pit, and then mining out each subsequent pit shell from the top down, before starting the next pit shell. This schedule is seldom feasible because the pushbacks are usually too narrow. Its usefulness lies in setting an upper limit to the Present Value (PV).
·	The Worst-Case DCF consists of mining each bench completely before starting on the next bench. This schedule’s usefulness lies in setting a lower limit to the Present Value.
·	The Pit Shell that best approximates the current Ultimate Pit Design is Pit Shell with Revenue Factor 0.60. This is considered a risk adverse pit shell.

Table 16‑7: Nested Pit Shell Results

Revenue Factor	Best Case DCF	Worst Case DCF	Mill Feed Tonnage<br> <br>(Mt)	Grade<br> <br>Au g/t	Grade<br> <br>Ag g/t	Waste Tonnage<br> <br>(Mt)	Total Tonnage<br> <br>(Mt)	Strip<br> <br>Ratio

| 0.1 | 35 | 35 | 0.1 | 3.95 | 2.03 | 0.2 | 0.3 | 1.3 |

| 0.2 | 380 | 378 | 3.0 | 1.89 | 6.78 | 9.5 | 12.5 | 3.1 |

| 0.3 | 1,463 | 1,430 | 21.2 | 1.22 | 5.60 | 46.0 | 67.3 | 2.2 |

| 0.4 | 2,752 | 2,604 | 57.5 | 1.02 | 4.73 | 112.0 | 169.0 | 2.0 |

| 0.5 | 3,864 | 3,495 | 111.3 | 0.92 | 4.95 | 290.0 | 401.4 | 2.6 |

| 0.6 | 3,992 | 3,549 | 122.6 | 0.90 | 4.93 | 336.1 | 458.7 | 2.7 |

| 0.7 | 4,140 | 3,518 | 144.8 | 0.86 | 5.08 | 463.2 | 608.0 | 3.2 |

| 0.8 | 4,213 | 3,468 | 162.0 | 0.83 | 5.01 | 557.5 | 719.5 | 3.4 |

| 0.9 | 4,233 | 3,416 | 172.3 | 0.82 | 4.93 | 621.6 | 793.9 | 3.6 |

| 1 | 4,235 | 3,399 | 174.4 | 0.81 | 4.91 | 637.4 | 811.9 | 3.7 |

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Figure 16‑8: Pit-by-Pit Graph

Source: AGP, 2025

16.6 Pit Design Parameters

16.6.1 Bench Design

The terminology for the pit design slope is described in Figure 16‑9.

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Figure 16‑9: Pit Slope Design Terminology

Source: AGP, 2025

The benching parameters were established following the geotechnical parameter recommendations presented in Table 16‑1. Working benches were designed for 35 to 40 m minimum mining width on pushbacks.

16.6.2 Haul Ramp Design

The haul ramp design is based on the largest truck planned for project. For the level study, the largest haul truck planned is a 240-tonne rigid frame truck. A Liebherr T264 is the example model.

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Figure 16‑10 and Figure 16‑11 illustrate the typical haul ramp profile. Table 16‑8 summarizes the haul ramp width calculation that has been used in the pit and phase designs.

Figure 16‑10: Haul Ramp Design Section, Double Lane

Source: AGP, 2025

Figure 16‑11: Haul Ramp Design Section, Single Lane

Source: AGP, 2025

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Table 16‑8: Haul Ramp Width Calculation

Haul Truck Parameters	Units	Value

| Model | | T264 |

| | | Liebherr |

| | | Rigid Frame |

| Payload (T, Heaped 2:1)) | | 240 tonne |

| Operating Width, W | m | 8.6 |

| Width Factor (of Truck Width) | | |

| Double Lane | | 3x |

| Single Lane | | 2x |

| Running Surface Double Lane | m | 25.9 |

| Running Surface Single-Lane | m | 17.2 |

| Tire Type | | 46/90R57 |

| Tire Overall Diameter | m | 3.6 |

| Factor (of Tire Size) | | 0.75x |

| Berm Height (Calculated) | m | 2.7 |

| Slope | degrees | 37 |

| Berm Width | m | 7.1 |

| Additional Allowance | m | 0.0 |

| Road Berm Allowance | m | 7.1 |

| Drainage Ditch Depth | m | 0.5 |

| Drainage Ditch Bottom Width | m | 0.5 |

| Slope (H:V) | | 1.5:1 |

| Ditch Width | m | 2.0 |

| Allowance from Toe | m | 0.5 |

| Road Drainage Allowance | m | 2.5 |

| Total Ramp Width Double Lane | m | 35.5 |

| Ramp Gradient | % | 10 |

16.6.3 Pit and Phase Selection

Pit designs were developed for the main pit as well as the small satellite pit immediately to the northeast using a 12 m bench height. The initial pit phase design included only the small satellite pit. This small pit is also above the Springpole Lake level so is convenient from an early mining perspective. The main pit has been divided into phases 2 and 3, with phase 3 being the ultimate pit. Pit limit analysis shells had been used to guide the ultimate pit were also used to outline areas of higher value for targeted early mining and phase development.

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Geotechnical parameters presented in 16.2.1 were applied to pit designs. It should be noted that the lowest RQD values were observed in the SE-n and SW design sectors, which is why these slopes are shallower than the other design sectors.

16.6.4 Pit Design

The Ultimate Pit Design for the Springpole Project is shown in Figure 16‑12 and Figure 16‑13.

Figure 16‑12: Ultimate Pit Design, Plan View

Source: AGP, 2025

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Figure 16‑13: Ultimate Pit Design, Section Views

Source: AGP, 2025

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Based on the above parameters, Table 16‑9 shows the ultimate pit design for the Springpole deposit volumetrics and dimensions.

Table 16‑9: General Pit Statistics

Item	Units	Value

| Pit Top Elevation | m | 411 |

| Pit Bottom Elevation | m | 88 |

| Pit Depth | m | 322 |

| Volume of Pit | m^3^ | 159,555,900 |

| Area of Pit Top | m^2^ | 1,303,825 |

| Perimeter at the Top of the Pit | m | 5,300 |

| Length from NE to SW | m | 1,085 |

| Length from NW to SE | m | 1,745 |

The contents within the ultimate pit design are shown in Table 16‑10. The numbers are reported from the mining block model and thus deemed to include mining dilution and loss.

Table 16‑10: Pit Inventory

Material	AuEQ Grade Bin<br> <br>g/t	Tonnage<br> <br>k tonne		Grade<br> <br>Au g/t	Grade<br> <br>Ag g/t		Grade<br> <br>AuEQ g/t

| East Extension Zone | 0.25 | | 956 | 0.39 | 0.64 | 0.35 | |

| | 0.5 | | 345 | 0.70 | 0.77 | 0.61 | |

| | 0.75 | | 1,114 | 2.45 | 1.27 | 2.14 | |

| Subtotal | | | 2,415 | 1.38 | 0.95 | 1.22 | |

| Camp Zone | 0.25 | | 2,079 | 0.39 | 1.30 | 0.35 | |

| | 0.5 | | 811 | 0.68 | 1.65 | 0.61 | |

| | 0.75 | | 1,507 | 1.87 | 1.90 | 1.65 | |

| Subtotal | | | 4,396 | 0.95 | 1.57 | 0.85 | |

| Portage Zone | 0.25 | | 35,677 | 0.39 | 2.66 | 0.36 | |

| | 0.5 | | 22,135 | 0.66 | 4.43 | 0.62 | |

| | 0.75 | | 36,442 | 1.61 | 7.96 | 1.48 | |

| Subtotal | | | 94,254 | 0.92 | 5.12 | 0.85 | |

| Portage High Grade | 0.25 | | 263 | 0.36 | 5.96 | 0.37 | |

| | 0.5 | | 172 | 0.64 | 7.21 | 0.62 | |

| | 0.75 | | 516 | 1.78 | 9.02 | 1.63 | |

| Subtotal | | | 951 | 1.18 | 7.85 | 1.10 | |

| Total | | | 102,015 | 0.94 | 4.90 | 0.86 | |

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Material	AuEQ Grade Bin<br> <br>g/t	Tonnage<br> <br>k tonne	Grade<br> <br>Au g/t	Grade<br> <br>Ag g/t	Grade<br> <br>AuEQ g/t

| Overburden | 0 | 19,200 | - | - | - |

| Other | 0 | 1,542 | - | - | - |

| Waste | 0 | 199,034 | - | - | - |

| East Extension Zone | 0 | 4,849 | 0.12 | 0.33 | 0.11 |

| Camp Zone | 0 | 3,298 | 0.13 | 0.74 | 0.12 |

| Portage Zone | 0 | 81,074 | 0.19 | 1.00 | 0.17 |

| Portage High Grade | 0 | 485 | 0.33 | 3.65 | 0.33 |

| Total | | 309,482 | 0.05 | 0.28 | 0.05 |

| Grand Total | | 411,497 | - | - | - |

16.6.5 Phases

Three phases have been proposed for the study. Table 16‑11, Table 16‑12, and Table 16‑13 tabulate the incremental tonnages and grades of the phases. Figure 16‑14 through Figure 16‑16 illustrate graphically the phase design.

Table 16‑11: Pit Inventory, Phase 1

Material	AuEQ Grade Bin<br> <br>g/t	Tonnage<br> <br>k tonne	Grade<br> <br>Au g/t	Grade<br> <br>Ag g/t	Grade<br> <br>AuEQ g/t

| East Extension Zone | 0.25 | 689 | 0.40 | 0.72 | 0.36 |

| | 0.5 | 318 | 0.70 | 0.80 | 0.62 |

| | 0.75 | 1,114 | 2.45 | 1.27 | 2.15 |

| Subtotal | - | 2,121 | 1.52 | 1.02 | 1.34 |

| Camp Zone | 0.25 | - | - | - | - |

| | 0.5 | - | - | - | - |

| | 0.75 | - | - | - | - |

| Subtotal | - | - | - | - | - |

| Portage Zone | 0.25 | 40 | 0.07 | 0.27 | 0.07 |

| | 0.5 | 21 | 0.44 | 0.91 | 0.39 |

| | 0.75 | - | 0.66 | 1.32 | 0.58 |

| Subtotal | - | 60 | 0.20 | 0.49 | 0.18 |

| Portage High Grade | 0.25 | - | - | - | - |

| | 0.5 | - | - | - | - |

| | 0.75 | - | - | - | - |

| Subtotal | - | - | - | - | - |

| Total | - | 2,181 | 0.22 | 0.47 | 0.19 |

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Material	AuEQ Grade Bin<br> <br>g/t	Tonnage<br> <br>k tonne	Grade<br> <br>Au g/t	Grade<br> <br>Ag g/t	Grade<br> <br>AuEQ g/t

| Overburden | 0 | 405 | - | - | - |

| Other | 0 | 296 | - | - | - |

| Waste | 0 | 4,385 | - | - | - |

| East Extension Zone | 0 | 1,826 | 0.12 | 0.46 | 0.11 |

| Camp Zone | 0 | - | - | - | - |

| Portage Zone | 0 | 6 | 0.07 | 0.27 | 0.07 |

| Portage High Grade | 0 | - | - | - | - |

| Total | - | 6,918 | 0.03 | 0.12 | 0.03 |

| Grand Total | - | 9,098 | - | - | - |

| Strip Ratio | - | 3.2 | - | - | - |

Table 16‑12: Pit Inventory, Phase 2

Material	AuEQ Grade Bin<br> <br>g/t	Tonnage<br> <br>k tonne	Grade<br> <br>Au g/t	Grade<br> <br>Ag g/t	Grade<br> <br>AuEQ g/t

| East Extension Zone | 0.25 | 62 | 0.36 | 0.57 | 0.32 |

| | 0.5 | 0 | 0.60 | 0.44 | 0.53 |

| | 0.75 | - | - | - | - |

| Subtotal | | 63 | 0.36 | 0.57 | 0.32 |

| Camp Zone | 0.25 | 1,823 | 0.39 | 1.30 | 0.35 |

| | 0.5 | 726 | 0.68 | 1.67 | 0.61 |

| | 0.75 | 1,335 | 1.89 | 1.99 | 1.67 |

| Subtotal | | 3,884 | 0.96 | 1.61 | 0.85 |

| Portage Zone | 0.25 | 19,770 | 0.13 | 0.93 | 0.12 |

| | 0.5 | 13,843 | 0.39 | 2.75 | 0.37 |

| | 0.75 | 24,586 | 0.66 | 4.45 | 0.62 |

| Subtotal | | 58,199 | 0.42 | 2.85 | 0.39 |

| Portage High Grade | 0.25 | 147 | 0.34 | 7.17 | 0.36 |

| | 0.5 | 83 | 0.61 | 10.65 | 0.63 |

| | 0.75 | 415 | 1.93 | 9.83 | 1.77 |

| Subtotal | | 645 | 1.40 | 9.33 | 1.31 |

| Total | | 62,791 | 0.45 | 2.80 | 0.42 |

| Overburden | 0 | 10,918 | - | - | - |

| Other | 0 | 573 | - | - | - |

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Material	AuEQ Grade Bin<br> <br>g/t	Tonnage<br> <br>k tonne	Grade<br> <br>Au g/t	Grade<br> <br>Ag g/t	Grade<br> <br>AuEQ g/t

| Waste | 0 | 68,419 | - | - | - |

| East Extension Zone | 0 | 482 | 0.18 | 0.61 | 0.16 |

| Camp Zone | 0 | 2,590 | 0.14 | 0.73 | 0.12 |

| Portage Zone | 0 | 23,355 | 0.13 | 0.92 | 0.12 |

| Portage High Grade | 0 | 165 | 0.74 | 5.16 | 0.69 |

| Total | | 106,504 | 0.03 | 0.23 | 0.03 |

| Grand Total | | 169,294 | - | - | - |

| Strip Ratio | - | 1.7 | - | - | - |

Table 16‑13: Pit Inventory, Phase 3

Material	AuEQ Grade Bin<br> <br>g/t	Tonnage<br> <br>k tonne	Grade<br> <br>Au g/t	Grade<br> <br>Ag g/t	Grade<br> <br>AuEQ g/t

| East Extension Zone | 0.25 | 205 | 0.38 | 0.40 | 0.34 |

| | 0.5 | 26 | 0.62 | 0.43 | 0.54 |

| | 0.75 | 0 | 1.09 | 0.83 | 0.96 |

| Subtotal | - | 231 | 0.41 | 0.40 | 0.36 |

| Camp Zone | 0.25 | 255 | 0.39 | 1.32 | 0.35 |

| | 0.5 | 85 | 0.68 | 1.48 | 0.61 |

| | 0.75 | 172 | 1.72 | 1.17 | 1.51 |

| Subtotal | - | 512 | 0.88 | 1.29 | 0.78 |

| Portage Zone | 0.25 | 15,867 | 0.15 | 0.93 | 0.14 |

| | 0.5 | 8,272 | 0.38 | 2.57 | 0.36 |

| | 0.75 | 11,856 | 0.66 | 4.39 | 0.61 |

| Subtotal | - | 35,995 | 0.37 | 2.44 | 0.34 |

| Portage High Grade | 0.25 | 116 | 0.39 | 4.43 | 0.38 |

| | 0.5 | 90 | 0.66 | 4.05 | 0.62 |

| | 0.75 | 100 | 1.14 | 5.66 | 1.05 |

| Subtotal | - | 305 | 0.72 | 4.72 | 0.67 |

| Total | - | 37,044 | 0.40 | 2.37 | 0.37 |

| Overburden | 0 | 7,878 | - | - | - |

| Other | 0 | 672 | - | - | - |

| Waste | 0 | 126,230 | - | - | - |

| East Extension Zone | 0 | 2,541 | 0.10 | 0.18 | 0.09 |

| Camp Zone | 0 | 707 | 0.12 | 0.75 | 0.11 |

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Material	AuEQ Grade Bin<br> <br>g/t	Tonnage<br> <br>k tonne	Grade<br> <br>Au g/t	Grade<br> <br>Ag g/t	Grade<br> <br>AuEQ g/t

| Portage Zone | 0 | 57,712 | 0.21 | 1.03 | 0.20 |

| Portage High Grade | 0 | 321 | 0.13 | 2.88 | 0.14 |

| Total | - | 196,061 | 0.06 | 0.31 | 0.06 |

| Grand Total | - | 233,105 | - | - | - |

| Strip Ratio | - | 5.3 | - | - | - |

Figure 16‑14: Phase Design Section Views

Source: AGP, 2025

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Figure 16‑15: Phase 1 Design, Plan View

Source: AGP 2025

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Figure 16‑16: Phase 2 Design, Plan View

Source: AGP, 2025

Phase 1 is the first phase mined in the schedule and comprises the small satellite pit to the northeast. The phase will be mined down to the 328 metres above sea level (“masl”) elevation. All waste and mill feed access will be on the west side of the pit in a slot configuration. The slot is the start of the ramp system for the entire pit design.

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Phase 2 is the second phase mined in the schedule and targets the upper and northern portion of the deposit. This phase will be mined down to the 136 masl elevation. Mine access is provided by the main ramp which is a continuation of the slot portion of the prior phase ramp. The ramp continues from 376 masl elevation on the southeast side of phase 2, clockwise to the bottom of the phase.

Phase 3 is the third and final phase mined in the schedule. This phase expands the pit to southwest targeting the lower portion of the deposit. Phase 3 is mined down to the 88 masl elevation. Access to this phase is provide by a ramp that branches off the prior ramp on the southeast side of the phase from the 376 masl elevation.

16.7 Waste Rock Storage Facility

Waste rock material generated from the open pit will be stockpiled in the following locations:

·	Overburden material suitable for reclamation purposes will be stockpiled separately.
·	Remaining waste rock material from the open pit will be stored in the CDF.

Waste rock and process plant tailings will be stored in the CDF that will be located to the northwest of the open pit. A small portion of the NAG waste rock will also be stored within the northern portion of the pit towards the end of the mine life once the CDF embankments have reached their design elevations.

The CDF is comprised of the following elements:

·	North Embankment
·	South Embankment
·	North Cell
·	South Cell

Most of the NAG material will be used to construct the embankments of the CDF. PAG waste rock and process plant tailings will be placed within cells inside of the embankments. A swell factor of 22% (considering some compaction) is used for waste rock and overburden material to estimate volume requirements. A placed density of 1.3 t/m^3^ was assumed for process plant tailings.

Access to the structure will be provided by a ramp located on the southeast corner of the CDF.

PAG waste rock and NAG process plant tailings are planned to be routed to the North Cell of the CDF, while PAG process plant tailings will be routed to the South Cell of the CDF. It has been assumed that 20% of the process plant tailings will be PAG and 80% will be NAG.

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Table 16‑14: Waste Material Storage Requirements

Waste Material	Tonnage	Volume

| From Open Pit | Mt | LCM |

| NAG Overburden | 17.6 | 9,070,200 |

| NAG Waste Rock | 103.8 | 47,187,800 |

| NAG Low RQD Rock | 46.2 | 20,995,500 |

| PAG Waste Rock | 141.8 | 64,471,200 |

| Total | 309.5 | 141,724,700 |

16.8 Life-of-Mine (LOM) Schedule

The mine production schedule consists of 102.0 Mt of mill feed grading 0.94 g/t Au and 4.90 g/t Ag. This provides a mine life of 9.4 years at a processing rate of 10.95 Mt/a. Mine overburden and rock waste tonnage totals 309.5 Mt and will be placed in the CDF, and a small in-pit facility. The overall pit strip ratio will be 3.0:1.

The mining production schedule includes eight years of mining, with one year of pre-production, and approximately one-and-a-half years of stockpile reclaim. The mill feed will consist of material coming from the three pit phases and three stockpiles. In Year 8, tonnage in the pit will be exhausted so mill feed material will be sourced from the remaining stockpiled material until midway through Year 10.

The LOM production schedule for the open pit area has been prepared using the Mine Plan Schedule Optimizer (MPSO) tool in the HexagonTM MinePlan 3D software. Provided with economic input parameters and operational constraints such as phase sequencing, maximum bench sink rates, and mining and milling capacities, the software determines the optimal mining sequence and ore stockpiling strategy, which maximizes the value of the mine production plan.

The overall objective of the mine scheduling and planning process is to maximize Project value while achieving the processing plant objectives and targets.

A mine production schedule was developed based on the following assumptions and criteria:

·	30,000 t/d Mill Feed Throughput.
·	10,950,000 t/a Mill Feed Throughput.
·	The mine plan considers the following ramp-up shown in Table 16‑15.

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Table 16‑15: Mine Plan Ramp-up Schedule

Year	Month	% of<br> <br>Steady State	Mill Feed Throughput<br> <br>(tonnes)

| 1 | 1 | 65% | 604,500 |

| 1 | 2 | 70% | 609,000 |

| 1 | 3 | 75% | 697,500 |

| 1 | 4 | 80% | 720,000 |

| 1 | 5 | 85% | 790,500 |

| 1 | 6 | 90% | 810,000 |

| 1 | 7 | 100% | 930,000 |

| 1 | 8 | 100% | 930,000 |

| 1 | 9 | 100% | 900,000 |

| 1 | 10 | 100% | 930,000 |

| 1 | 11 | 100% | 900,000 |

| 1 | 12 | 100% | 930,000 |

·	A maximum sink rate of eight benches (96 m) per year was implemented.
·	First two years of mining scheduled on monthly basis, followed by annual basis afterward.

Ore stockpiles were considered with a soft constraint of 20,000,000 t capacity for the Low Grade (LG) stockpile. Stockpile designations are:

·	High Grade (HG) Stockpile >= 0.75 g/t AuEQ
·	Mid Grade (MG) Stockpile < 0.75 g/t and >= 0.5 g/t AuEQ
·	Low Grade (LG) Stockpile < 0.5 g/t and >= 0.25 g/t AuEQ
·	Schedule constraint order of priority:
·	Mill Feed tonnage
·	Material Mined Tonnage
·	S grade to Mill, 2.9% S upper limit
·	Au Grade to Mill, Upper limit 1.0 g/t, after initial years
·	Ore stockpiling balance

The mine plan has been developed in order to meet plant feed requirements according to general best open pit mine practices such as equipment fleet smoothing and maximizing NPV.

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A pre-production stripping period at the Springpole was considered. Table 16‑16 tabulates the LOM production plan and Figure 16‑17 to Figure 16‑20 graphically show the LOM production plan.

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Table 16‑16: LOM Production Schedule

Description	Unit	Total	Year -1	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	Year 10

| Mill Feed | tonne | 102,015,282 | 0 | 9,751,500 | 10,950,000 | 10,950,000 | 10,950,000 | 10,950,000 | 10,950,000 | 10,950,000 | 10,950,000 | 10,950,000 | 4,663,782 |

| Au Grade | g/t | 0.94 | 0.00 | 1.24 | 1.09 | 1.00 | 1.00 | 1.00 | 1.00 | 1.00 | 0.99 | 0.39 | 0.39 |

| Ag Grade | g/t | 4.90 | 0.00 | 4.38 | 6.55 | 7.98 | 5.81 | 3.42 | 4.04 | 4.89 | 6.23 | 1.96 | 1.96 |

| ROM to Mill | tonne | 68,497,277 | 0 | 7,277,211 | 10,765,800 | 10,950,000 | 8,782,274 | 2,800,597 | 6,021,396 | 10,950,000 | 10,950,000 | 0 | 0 |

| Au Grade | g/t | 1.00 | 0.00 | 1.22 | 1.10 | 1.00 | 1.15 | 0.49 | 0.60 | 1.00 | 0.99 | 0.00 | 0.00 |

| Ag Grade | g/t | 6.02 | 0.00 | 5.31 | 6.59 | 7.98 | 6.76 | 2.89 | 4.27 | 4.89 | 6.23 | 0.00 | 0.00 |

| ROM to Stockpile | tonne | 33,518,005 | 2,294,612 | 4,886,307 | 7,406,957 | 12,650,456 | 732,445 | 129,727 | 1,748,484 | 2,705,512 | 963,504 | 0 | 0 |

| Au Grade | g/t | 0.8 | 1.1 | 0.6 | 0.6 | 1.1 | 1.6 | 1.7 | 0.3 | 0.4 | 0.4 | 0.0 | 0.0 |

| Ag Grade | g/t | 2.61 | 1.3 | 2.1 | 2.4 | 3.4 | 3.3 | 1.0 | 1.7 | 1.8 | 2.8 | 0.0 | 0.0 |

| Stockpile to Mill | tonne | 33,518,005 | 0 | 2,474,289 | 184,200 | 0 | 2,167,726 | 8,149,403 | 4,928,604 | 0 | 0 | 10,950,000 | 4,663,782 |

| Au Grade | g/t | 0.8 | 0.0 | 1.3 | 0.7 | 0.0 | 0.4 | 1.2 | 1.5 | 0.0 | 0.0 | 0.4 | 0.4 |

| Ag Grade | g/t | 2.6 | 0.0 | 1.7 | 4.2 | 0.0 | 2.0 | 3.6 | 3.8 | 0.0 | 0.0 | 2.0 | 2.0 |

| Stockpile Balance | t | | 2,294,612 | 4,706,630 | 11,929,387 | 24,579,843 | 23,144,561 | 15,124,886 | 11,944,766 | 14,650,278 | 15,613,782 | 4,663,782 | 0 |

| Stockpile Balance (Au) | units | | 2,570,301 | 2,035,001 | 6,710,554 | 20,433,845 | 20,764,506 | 11,391,634 | 4,640,862 | 5,715,177 | 6,060,691 | 1,810,307 | 0 |

| Stockpile Balance (Au) | g/t | | 1.1 | 0.4 | 0.6 | 0.8 | 0.9 | 0.8 | 0.4 | 0.4 | 0.4 | 0.4 | 0.0 |

| Overburden | tonne | 19,199,995 | 4,322,243 | 8,311,048 | 887,534 | 761,559 | 4,848,367 | 69,244 | 0 | 0 | 0 | 0 | 0 |

| Waste Rock | tonne | 290,281,962 | 13,387,144 | 33,558,222 | 30,939,709 | 30,637,985 | 45,636,915 | 47,000,432 | 48,211,273 | 31,824,379 | 9,085,903 | 0 | 0 |

| Total Material Mined | tonne | 411,497,239 | 20,004,000 | 54,032,788 | 50,000,000 | 55,000,000 | 60,000,000 | 50,000,000 | 55,981,153 | 45,479,891 | 20,999,407 | 0 | 0 |

| Total Material Moved | tonne | 445,015,244 | 20,004,000 | 56,507,077 | 50,184,200 | 55,000,000 | 62,167,726 | 58,149,403 | 60,909,757 | 45,479,891 | 20,999,407 | 10,950,000 | 4,663,782 |

| Stripping Ratio | | 3.0 | 7.7 | 3.4 | 1.8 | 1.3 | 5.3 | 16.1 | 6.2 | 2.3 | 0.8 | 0.0 | 0.0 |

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Figure 16‑17: Mill Feed Tonnage and Au Grade by Year Graph

Source: AGP, 2025

Figure 16‑18: Total Material Mined and Mining Production Rate by Year Graph

Source: AGP, 2025

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The annual mining rate starts at 20.3 Mt/a in the pre-production year and reaches a peak of 60.0 Mt/a in Year 6. A maximum descent rate of eight benches/year/phase is applied to ensure that reasonable mining operations and ore control will occur.

Figure 16‑19: Material Mined Production Graph

Source: AGP, 2025

Figure 16‑20: Material Mined by Phase by Year Graph

Source: AGP, 2025

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16.9 LOM Plan Sequence

The open pit progression by year is illustrated in Figure 16‑21 through Figure 16‑29.

Figure 16‑21: End-of-Year Progression Plan, Year -1

Source: AGP, 2025

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Year -1: Mining is initiated in Phase 1 and 2 of the open pit with some stripping from Phase 3 area. In this period, a total of 17.7 Mt of waste material is moved as the Project ramps up. As the processing plant is not yet operational, 2.3 Mt of material grading 1.12 g/t Au and 1.31 g/t Ag is stockpiled. Phase 1 and 2 of the open pit descend to the 376 masl elevations. NAG/non=metal leaching waste material is routed to construction of the CDF facility and PAG material is placed in the North Cell.

Figure 16‑22: End-of-Year Progression Plan, Year 1

Source: AGP, 2025

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Year 1: Mining continues in Phase 2. Phase 1 is completed to the 328 masl elevation. Phase 2 descends to the 340 masl elevations. The processing plant is operational, and 9.75 Mt grading 1.24 g/t Au and 4.39 g/t Ag are fed to it. A total of 41.9 Mt of waste is mined. NAG/non-metal leaching waste material is routed to the construction of the North and South Embankments. PAG waste and NAG process plant tailings material are routed to North Cell while the PAG process plant tailings are routed to the South Cell.

Figure 16‑23: End-of-Year Progression Plan, Year 2

Source: AGP, 2025

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Year 2: Mining continues in Phase 2 and descends to the 292 masl elevations. Feed of 10.95 Mt grading 1.22 g/t Au and 6.81 g/t Ag is processed, and 32.6 Mt of waste is mined.

Figure 16‑24: End-of-Year Progression Plan, Year 3

Source: AGP, 2025

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Year 3: Mining continues in Phase 2 and descends to the 196 masl elevations. Feed of 10.95 Mt grading 1.00 g/t Au and 8.37 g/t Ag is processed, and 25.2 Mt of waste is mined.

Figure 16‑25: End-of-Year Progression Plan, Year 4

Source: AGP, 2025

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Year 4: Mining continues to the bottom of Phase 2 at the 136 masl elevation. Phase 3 re-commences and descends to the 352 masl elevation. Feed of 10.95 Mt grading 1.00 g/t Au and 5.50 g/t Ag is processed. 3.9Mt of material is rehandled from the ROM Stockpiles and 47.4 Mt of waste is mined.

Figure 16‑26: End-of-Year Progression Plan, Year 5

Source: AGP, 2025

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Year 5: Mining continues in Phase 3 and descends to the 304 masl elevation. Feed of 10.95 Mt grading 1.00 g/t Au and 3.19 g/t Ag is processed. 8.3Mt of material is rehandled from the ROM Stockpiles. 52.2 Mt of waste is mined.

Figure 16‑27: End-of-Year Progression Plan, Year 6

Source: AGP, 2025

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Year 6: Mining continues in Phase 3 and descends to the 232 masl elevation. Feed of 10.95 Mt grading 0.87 g/t Au and 3.85 g/t Ag is processed. Head grades are trending lower as the pit progresses into a lower grade zone of the pit and reclaiming from the LG stockpile begins. 4.2Mt of material is rehandled from the ROM Stockpiles. 51.8 Mt of waste is mined.

Figure 16‑28: End-of-Year Progression Plan, Year 7

Source: AGP, 2025

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Year 7: Mining proceeds in Phase 3 descending to the 172 masl elevation. Feed of 10.95 Mt grading 1.00 g/t Au and 4.91 g/t Ag is processed, and 31.7 Mt of waste is mined.

Figure 16‑29: End-of-Year Progression Plan, Year 8 (End of Mining)

Source: AGP, 2025

Year 8: Mining continues to the bottom of Phase 3 to the 88 masl elevation. Feed of 10.95 Mt grading 0.99 g/t Au and 6.25 g/t Ag is processed, and 9.0 Mt of waste is mined. 1.3 Mm^3^ of the NAG waste is stored in pit at the north end.

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Year 9: LG Stockpile reclaim of 10.95 Mt grading 0.39 g/t Au and 2.0 g/t Ag is processed.

Year 10: Stockpiles are exhausted during this year with 4.7 Mt grading 0.39 g/t Au and 2.00 g/t Ag is processed.

16.9.1 Blasting and Explosives

Blast patterns have been estimated for the purposes of generating a PFS level equipment plan and cost estimate. Drill and blast patterns have been estimated at 7.7 m x 6.7 m (spacing x burden). Holes will be 12 m long plus an additional 1.3 m for sub-drill for a total length of 13.3 m using a 251 mm bit.

Powder factors are estimated at 0.29 kg/t using bulk emulsion explosives. It is recommended to use emulsion due to the expected wet conditions.

Additional drilling capability will be available with smaller 140 mm drills. They will supplement the drilling and provide the pre-shear drilling. When used in regular blast patterns, mill feed blast patterns will be 4.6 x 4.0 m (spacing x burden) and for mill feed will be 4.8 x 4.2 m (spacing x burden). The holes will be 12 m plus an additional 0.8 m sub-drill totaling 12.8 m. The powder factor for the smaller drill will be 0.26 kg/t for mill feed and 0.29 kg/t for waste.

It has been assumed that wall control drill and blast methods will be required for final walls. For purposes of this study, the Pre-shear method has been assumed. Pre-shear holes will be 12 m deep, spaced 1.65 m apart and be separated from production blasts by 2 m. The powder charge will be 62 kg per hole in a decoupled manner.

16.10 Mining Equipment

The following section discusses the fleet requirements that were estimated to carry out the open pit mine production plan. This section includes indicative parameters for drilling, blasting, loading, and hauling. The objective of equipment selection for this level of study is to produce an estimate of costs suitable for a PFS level study and not necessarily to design an optimized equipment fleet.

For purposes of estimating typical fleet requirements, all equipment is assumed to be owned, operated, and maintained by First Mining. The mining equipment has been planned as a combination of both electric-powered and diesel-powered.

Open pit mine operations are based on 365 days per year and correspond to operations running two 12-hour shifts per day, seven days per week, with the assumption that five operating days will be lost on average due to inclement weather and two to non-production days.

Drilling will be completed with down the hole hammer (DTH) drills with a 251 mm bit. This provides the capability to drill 12 m bench heights in a single pass. The supplemental drill will also be a down the hole hammer drill with a 140 mm bit. This is capable of drilling 12 m benches but must add steel from its carousal.

The primary loading units will be 37 m^3^ hydraulic shovels. Additional loading will be completed by 23 m^3^ front-end loaders. It is expected that one of the loaders will be at the primary crusher for the majority of its operating time. The haulage trucks will be conventional 240-tonne rigid body trucks.

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The support equipment fleet will be responsible for the usual road, pit, and dump maintenance requirements, but due to the climatic conditions expected, the support equipment will have a larger role in snow removal and water management. Snowplows and additional graders have been included in the fleet. In addition, smaller road maintenance equipment is included to keep drainage ditches open and sedimentation ponds functional.

Figure 16‑30: Major Production Equipment, by type, by year

Source: AGP, 2025

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Figure 16‑31: Support Equipment, by type, by year

Source: AGP, 2025

16.11 Grade Control

Grade control will be completed with a separate fleet of reverse circulation (RC) drill rigs. They will drill the deposit off on a 10 x 5 m pattern in areas of known mineralization, taking samples each metre. The holes will be inclined at 60°.

In areas of low-grade mineralization or waste, the pattern spacing will be increased to 20 x 10 m with sampling over 12 m. These holes will be used to find undiscovered veinlets or pockets of mineralization.

These grade control holes serve to define the mill feed grade and mineralization contacts.

Samples collected will be sent to the assay laboratory and assayed for use in the short-range mining model.

Blasthole sampling will also be part of the grade control program initially to determine the best method for the planned Springpole operations.

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16.12 Quarry Area

Two quarry areas have been considered for the project. The CDF Quarry, located within the footprint of the CDF facility, and the fish habitat development area, is situated adjacent to the pit in the southeast corner. This area contains a large peat zone which will be used for later reclamation purposes.

The intent of mining in these areas is to provide rock for various infrastructure projects during the early works and pre-production period as well as at closure. This includes:

Haul roads to the:

·	pit
·	CDF
·	dikes
·	fill material for the dikes
·	fill material for CDF Embankments
·	fill material for construction purposes

The CDF Quarry is planned to be worked in Year -3 and Year -2 of the overall schedule with a smaller equipment fleet. During Year -2 and Year -1, a small fleet of larger mining equipment will also be used to excavate and haul the necessary quantities of material.

A small fleet of equipment will be used for pre-construction activities including road building, CDF foundation preparation and dike construction. The fleet will be comprised of 91-tonne rigid body trucks and 6 m^3^ excavator and 13 m^3^ front-end loader. They will be initially matched with the smaller drills.

The fish habitat development area quarry can be activated at the completion of the open pit with the larger mine fleet. The intention is to lower the area generally to a depth 3+ m below the proposed final lake level after mining is complete. There will also be a deeper basin area incorporated that will be approximately 30 m in depth. Material in this final stage of excavation will be used for cover material at the CDF cells as well as be dumped into the pit to flatten the slopes of the pit before flooding at the end of the mine life.

Total material planned to be excavated from the CDF Quarry is 46.2 Mt, including overburden removal. Total material planned to be mined from the fish habitat development area quarry is approximately 26.2 Mt

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17 RECOVERY METHODS

17.1 Overview

The metallurgical testwork was thoroughly analyzed and several options of process routes were assessed in the initial stages of the 2025 PFS. The processing route was selected based on the testwork analysis, the project requirement for split tailings streams for flotation concentrate and flotation tailings to manage potential acid generating (PAG) rock, and economic evaluation of processing alternatives. The unit operations selected are typical for this industry. The material is found to be highly variable, both in terms of ore hardness and metallurgical response. The flowsheet selected has catered for a variable feed through selection of appropriate design criteria. The project will utilize a capital cost-effective mill design, one stage of rougher flotation, flotation concentrate regrind as well as flotation concentrate leaching, flotation tailings leaching and adsorption, carbon elution, concentrate and tailings cyanide detoxification, and metal recovery via Merrill Crowe for both concentrate and tailings.

Key process design criteria are listed below:

·	Throughput of 30.0 kt/d or 10.95 Mt/a.
·	Crushing availability of 75%.
·	Plant availability of 92% for grinding, flotation, leach, CCD and cyanide detoxification.

17.2 Process Flowsheet

The process design considers the following unit operations and circuits:

·	Primary crushing of run-of-mine (ROM) ore.
·	Grinding circuit comprising of a SAG mill followed by a ball mill with cyclone classification and trash screen.
·	Rougher flotation.
·	Flotation concentrate regrind, thickening and leaching.
·	Concentrate CCD circuit.
·	Flotation tailings thickening, leaching and CIP adsorption.
·	Acid washing of loaded carbon and Anglo-American Research Laboratory (AARL) type elution.

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·	Carbon regeneration.
·	Concentrate and tailings Merrill-Crowe precipitation circuit.
·	Concentrate and tailings cyanide detoxification.
·	Gold room.

An overall process flow diagram showing the unit operations in the process plant is presented in Figure 17 1 and is described in the sections below.

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Figure 17 1: Process Flowsheet

Source: Ausenco, 2025

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17.3 Plant Design

The key process design criteria for the plant are listed in Table 17 1.

Table 17 1: Process Design Criteria

Design Parameter	Units	Value

| Plant throughput | t/d | 30,000 |

| Design head grades | | |

| sulphur, maximum | % | 2.9 |

| sulphur, minimum | % | 1.8 |

| Concentrate leach feed gold | g/t | 5.7 |

| Concentrate leach feed silver | g/t | 42.5 |

| Tailings leach feed gold | g/t | 0.4 |

| Tailings leach feed silver | g/t | 3.2 |

| Crushing plant availability | % | 75 |

| Mill availability | % | 92 |

| Bond Crusher Work Index (CWi) | kWh/t | 10.7 |

| Bond Ball Mill Work Index (BWi) | kWh/t | 15.9 |

| Bond Abrasion Index (Ai) | g | 0.157 |

| SPI | min | 74.7 |

| JK Dropweight Parameter Axb | - | 65.2 |

| Crushing circuit throughput | t/h | 1667 |

| Milling and flotation throughput | t/h | 1359 |

| Concentrate regrind and leach throughput, design | t/h | 198 |

| Tailings leach throughput, design | t/h | 1238 |

| Grinding circuit product size, P80 | µm | 150 |

| Concentrate regrind product size (P80) | µm | 17 |

| Rougher flotation residence time - design | minutes | 24 |

| Mass recovery to concentrate | % | 14-Sep |

| Concentrate leach pre-aeration time | h | 8 |

| Concentrate leach residence time | h | 12 |

| CCD soluble loss, gold in solution, design | mg/L | <0.05 |

| CCD wash ratio | m^3^/m^3^ leach feed solution | 3.5 |

| Tailings leach residence time | h | 16 |

| CIP gold soluble tailing, design | mg/L | <0.015 |

| Elution carbon batch size | t | 13 |

| Number of carbon strips per week, max | | 12 |

| Cyanide destruction method | - | SO2 /air |

| Detox WAD cyanide concentration, design | mg/L CNWAD | 2.5 |

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17.3.1 Primary Crushing & Stockpiling

The crushing circuit is designed for an annual operating time of 75% availability. Ore is hauled from the open pit to the primary crusher dump pocket that feeds the primary crusher at 1,667 t/h. The crushed ore is conveyed to a covered stockpile that provides approximately 12 h of live storage. Crushed ore storage capacity will allow routine crusher maintenance to be carried out without interrupting feed to the mill.

The mill feed stockpile is equipped with apron feeders to regulate feed at 1,359 t/h into the SAG mill. Crushed ore is withdrawn from the stockpile by two apron feeders (one operating, one stand-by) and feeds the mill circuit via the SAG mill feed conveyor.

The materials handling and crushing circuit includes the following key equipment:

·	Primary crusher
·	Mill feed apron feeders (equipped with variable speed drives or VSD’s)
·	Crushed ore stockpile
·	Belt conveyors and feeders

17.3.2 Grinding Circuit

The grinding circuit consists of a SAG mill followed by a ball mill in closed circuit with cyclones. The circuit is designed to provide a product size of 80% passing (P80) 150 µm. The SAG mill slurry discharges through a trommel screen where the pebbles are screened out and recycled to the SAG mill via conveyor. Trommel undersize discharges into the cyclone feed pumpbox. The SAG mill is designed for variable speed operation to accommodate variable ore hardness.

The ball mill is fed by the cyclone underflow and discharges through a trommel screen. Trommel screen oversize is discharged to a scats bunker. Trommel screen undersize discharges into the cyclone feed pumpbox.

Both the SAG and ball mill have dual pinion drives.

Water is added to the cyclone feed pumpbox to obtain the required cyclone feed density. Cyclone overflow gravitates to the rougher flotation circuit after passing through a trash screen.

SAG mill grinding media will consist of 125-150 mm diameter forged steel balls that will be loaded by front-end loader from a storage bunker into the SAG mill ball charging hopper and SAG mill ball feeder that feeds the SAG mill feed conveyor. The ball mill grinding media will consist of 50-75 mm diameter forged steel balls that are loaded into the ball mill ball storage bin from the ball mill storage bunker by a front-end loader. Grinding media is added as required to the ball mill using a kibble.

The grinding circuit includes the following key equipment:

·	SAG mill (10.4 m dia. x 5.2 m EGL, 11.2 MW, variable speed)
·	Ball mill (7.9 m dia. x 11.4 m EGL, 14.2 MW)
·	Cyclone feed pumpbox
·	Cyclone cluster, with 10 each 750 mm diameter cyclones

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17.3.3 Flotation Circuit

The flotation circuit is comprised of one stage of rougher flotation. Ground ore from the classification cyclones feed five rougher flotation cells in series. The rougher flotation circuit requires the addition of a collector potassium amyl xanthate (PAX) and frother, methyl isobutyl carbinol (MIBC). Concentrate from the cells is collected and pumped to the regrind circuit where the concentrate will undergo further liberation to expose gold and silver minerals for leaching.

The flotation circuit includes the following key equipment:

·	Rougher flotation cells (5 x 300 m3)
·	Flotation tailings pumpbox
·	Flotation concentrate pumpbox

17.3.4 Flotation Concentrate Regrind Circuit

Rougher flotation concentrate is directed to the regrind cyclone cluster. Cyclone cluster underflow is evenly split between two regrind mills operating in open circuit. The regrind mill product, with a P80 of 17 µm, is combined with the cyclone overflow and sent to the concentrate thickener.

The rougher concentrate regrind circuit includes the following key equipment:

·	Regrind cyclone cluster, with eight each 250 mm diameter cyclones.
·	Regrind mills (2 x M15000 IsaMill, 3.8 MW each)

17.3.5 Flotation Concentrate Dewatering

Regrind mill product and regrind cyclone overflow combine and report to the flotation concentrate thickener (23 m diameter). The thickener will have a flocculant addition rate of 60 g/t and target underflow density of 45% solids by weight before being pumped to the concentrate leach tanks.

17.3.6 Concentrate Leach

The flotation concentrate leach circuit consists of one pre-aeration tank providing eight hours residence time and four leach tanks, providing 12 h residence time.

Oxygen is added to each tank to maintain adequate dissolved oxygen levels for leaching and hydrated lime is added to control the operating pH at 10.5 - 11. Cyanide solution is also added to all leach tanks.

The flotation concentrate leach circuit includes the following key equipment:

·	Concentrate leach pre-aeration tank (2,700 m3)
·	Concentrate leach tanks (4 x 1,000 m3)

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17.3.7 Counter Current Decantation

Concentrate leach slurry will report to the first of six CCD thickeners with the thickened solids being pumped up the CCD train from CCD thickener No. 1 to No. 6. The leached solids will be washed with barren solution from the concentrate Merrill Crowe circuit that is mixed with the slurry feeding the sixth thickener in the CCD train. Barren solution is added at a wash ratio of 3.5:1 of incoming solution from the leach circuit. CCD thickener overflow solution flows by gravity down the CCD train, from CCD thickener No. 6 to No. 1. Solids settling will be facilitated with flocculant addition to each thickener. Flocculant additions range from 60 g/t to 10 g/t, added to the interstage mix tanks. The overflow (pregnant solution) from the first CCD thickener reports to a clarifier.

Clarifier overflow will advance to a pregnant solution tank to feed the concentrate Merrill Crowe circuit. Underflow from the clarifier will be pumped back to the clarifier feed box by the clarifier recirculation pump. Intermittently, the clarifier underflow pump will pump the accumulated settled solids back to the concentrate leaching circuit.

The counter current decantation circuit includes the following key equipment:

·	Count current decantation thickeners (6 x 23 m diameter)
·	Interstage agitated mix tanks
·	Clarifier (29 m diameter)

17.3.8 Flotation Tailings Dewatering

Rougher flotation tailings report to the flotation tailings thickener. The thickener will have a flocculant addition rate of 16 g/t and target underflow density of 53% w/w before being sent to the flotation tailings leach circuit.

The flotation tailings dewatering circuit includes the following key equipment:

·	Flotation tailings trash screens (2)
·	Flotation tailings thickener (40 m diameter)

17.3.9 Flotation Tailings Leach & Adsorption Circuit

The flotation tailings leach circuit consists of four leach tanks and six carbon in pulp (CIP) tanks, providing a total leach/CIP residence time of 16 h. The circuit is fed by the flotation tailings thickener underflow.

Cyanide and oxygen will be added to the leach tanks to cause the gold to go into solution. Hydrated lime is also added to for pH control. Leached slurry flows to the CIP tanks and where it is contacted with activated carbon to adsorb soluble gold and silver. The activated carbon is pumped counter currently to the leached slurry flow. One inter-tank screen in each CIP tank retains the carbon whilst allowing the slurry to flow by gravity to the downstream tanks. Recessed impeller pumps are used to transfer slurry between the CIP tanks and from the lead tank to the loaded carbon screen mounted above the acid wash column in the elution circuit.

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The flotation tailings leach and carbon adsorption circuit include the following key equipment:

·	Tailings leach tanks (4 x 4,200 m3)
·	CIP tanks (6 x 1,600 m3)
·	Inter-tank carbon screens (6)
·	Loaded carbon screen
·	Carbon sizing screen

17.3.10 Cyanide Detoxification

Leach-adsorption tailings feed the tailings cyanide detoxification circuit at 50% solids and underflow from the sixth CCD thickener feeds the concentrate detoxification circuit at 40% solids. The tailings detoxification circuit and the concentrate detoxification circuit both consist of two tanks configured in parallel, each with a residence time of approximately 90 mins to reduce weak acid dissociable cyanide (CNWAD) concentration from 200 mg/L to less than 5.0 mg/L to comply with environmental requirements prior to deposition in the tailings facilities.

Cyanide detoxification is undertaken using the SO2/air process. The reagents required are oxygen, lime, copper sulphate, and sulphur dioxide. The cyanide detoxification tanks are equipped with SO2 addition sparging points and an agitator to ensure that the reagents are thoroughly mixed with the tailings slurry. From the tailings cyanide detoxification tanks, the tailings report to the carbon safety screen. Screen undersize feeds the tailings surge tanks ahead of the co-disposal tailings facility, screen oversize (recovered carbon) is collected in fine carbon bags. From the concentrate detoxification tanks, the tailings report to the PAG tailings deposition cell.

The main equipment in this area includes:

·	Flotation tailings cyanide detoxification tanks (2 x 1,500 m3)
·	Flotation tailings surge tanks (2 x 6,000 m3)
·	Flotation concentrate cyanide detoxification tanks (2 x 300 m3)
·	CIP carbon safety screen

17.3.11 Carbon Acid Wash

Prior to the gold stripping stage, loaded carbon from the tailings CIP circuit is treated with a weak hydrochloric acid solution to remove calcium, magnesium, and other salt deposits that could render the elution less efficient or become baked on in subsequent steps and ultimately foul the carbon.

Loaded carbon from the loaded carbon recovery screen flows by gravity to the acid wash column. Entrained water is drained from the column, and the column is refilled from the bottom up with the hydrochloric acid solution. Once the column is filled with the acid, it is left to soak, after which the spent acid is rinsed from the carbon and discarded to the tailings cyanide detoxification tank.

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The acid-washed carbon is then hydraulically transferred to the elution column for elution of gold and silver.

The main equipment in this area includes:

·	Acid wash carbon column (13 t capacity)
·	Loaded carbon screen

17.3.12 Carbon Stripping (Elution)

The carbon stripping (elution) circuit uses the AARL process.

The elution sequence commences with the addition of a weak NaOH and weak NaCN solution into the bottom of the elution column. Once the prescribed volume has been added, the pre-soak period commences. During the pre-soak, the caustic/cyanide solution is circulated through the column and the elution heater until the working temperature is achieved.

Upon completion of the pre–soak period, additional water is pumped through the recovery heat exchanger and elution heater, then through the elution column to the pregnant solution tank at a rate of 2.0 Bed Volumes (BV)/h. At this stage, the temperature of the strip solution passing through the column is maintained at 120°C and the gold is stripped off the loaded carbon.

Strip solution flows up and out of the top of the column, passing through the recovery heat exchanger via the elution discharge strainers and to the pregnant eluate tank.

Upon completion of the cool down sequence, the carbon is hydraulically transferred to the carbon regeneration kiln feed hopper via a de-watering screen.

The stripping circuit includes the following key equipment:

·	Elution column (13 t capacity)
·	Strip solution heater with heat exchanger
·	Strip solution tank
·	Pregnant eluate tank

17.3.13 Carbon Reactivation

Dewatered barren carbon from the stripping circuit is held in a feed hopper. A screw feeder metres the carbon into the reactivation kiln, where it is heated to 850°C in an atmosphere of superheated steam to restore the activity of the carbon. The kiln will be fitted with mercury scrubbing equipment.

Carbon discharging from the kiln is quenched in water and screened on the carbon sizing screen located on top of the CIP tanks to remove undersized carbon fragments. The undersize fine carbon gravitates to the carbon safety screen, whilst carbon screen oversize is directed to the CIP circuit.

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As carbon is lost by attrition, new carbon is added to the circuit using the carbon quench tank. The new carbon is then transferred along with the regenerated carbon to feed the carbon sizing screen.

The carbon reactivation circuit includes the following key equipment:

·	Carbon dewatering screen
·	Regeneration kiln including feed hopper (1,000 kg/h capacity)
·	Carbon quench tank

17.3.14 Concentrate Merrill Crowe Circuit

Overflow from the clarifier flows into the pregnant solution tank. Solution will be pumped from the pregnant solution tank through pressure leaf clarifiers to further reduce total suspended solids (TSS) and subsequently to a deaeration tower to remove dissolved oxygen. De-aerated solution is then mixed with zinc powder and lead nitrate to precipitate the gold and silver and pumped to concentrate precipitate filters to recover the precious metal rich precipitate. Solids from the filters will be periodically discharged and moved within the refinery for doré production, while filtrate will return to the barren solution tank. Barren solution will then be distributed to the CCD circuit as wash water with a bleed stream to the concentrate detoxification circuit as necessary.

Major equipment in this area will include:

·	Pregnant solution tank
·	Concentrate deaeration tower
·	Pressure leaf clarifier filters (3)
·	Concentrate plate and frame precipitate filters (3)
·	Barren solution tank

17.3.15 Tailings Merrill Crowe Circuit

Solution will be pumped from the pregnant eluate tank directly to the tailings precipitate filters to recover the precious metal rich precipitate. The hot eluate has low dissolved oxygen concentration, so no further deaeration is required. Zinc power will be added prior to the filters to precipitate gold and silver. Like the concentrate Merrill Crowe circuit, solids from the filters will be discharged and moved within the refinery for doré production, while filtrate will advance to the common strip solution tank.

Major equipment in this area will include:

·	Tailings precipitate filters (3)

17.3.16 Gold Room

Zinc precipitate from both Merill Crowe circuits will be transported manually using trays to the mercury retort oven for mercury removal and drying. Mercury collected will be sent off site for third-party processing. Dried zinc precipitate will be mixed with fluxes and loaded into the electric induction furnace for smelting. The fluxes will react with the base metals to form oxides that report to the slag and separate from the silver and gold molten metal. The molten metal will be poured into moulds to form the doré bars. The doré bars will be cleaned, assayed, stamped, and stored in a secure vault ready for regular transfer to off site refiners.

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The refinery will be a vendor package and will include the following major equipment:

·	Flux dosing and mixing system
·	Mecury retort and gas scrubbing systems
·	Melting furnace
·	Doré scale and storage vault
·	Slag handling equipment
·	Dust collection system

17.4 Product/Materials Handling

All process plant unit operations are located in a single facility with a compact layout. Detoxified tailings from the flotation concentrate and tailings circuits are pumped through enclosed pipelines to the PAG tailings deposition cell and the co-disposal tailings facility respectively.

17.5 Energy, Water and Process Materials Requirements

17.5.1 Process Materials Requirements

Reagents and consumables used in the process plant are listed in Table 17 2 and Table 17 3 along with the estimated yearly average consumption for each item. Details on reagent handling and storage systems are described in the following subsections.

Table 17 2: Average Annual Reagents Consumption

Name	Value (tonnes)

| Collector (PAX) | 1,095 |

| Frother (MIBC) | 153 |

| Quick lime | 27,426 |

| Oxygen | 17,819 |

| Sulphur Dioxide | 11,468 |

| PAX | 1,095 |

| Flocculant 1 (AN920 or equivalent) | 154 |

| Flocculant 2 (AN913SH or equivalent) | 329 |

| Coagulant | 289 |

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Name	Value (tonnes)

| Sodium Cyanide | 7,079 |

| Sodium Hydroxide | 962 |

| Hydrochloric Acid | 1,557 |

| Copper Sulphate | 1,132 |

| Activated Carbon | 289 |

| Zinc Powder | 314 |

| Diatomaceous Earth | 690 |

| Lead Nitrate | 23 |

| Fluxes | 66 |

| Antiscalant | 29 |

Table 17 3: Average Annual Processing Consumables

Item	Unit	Consumption

| SAG mill media | tonnes | 2281 |

| Ball mill media | tonnes | 7814 |

| Regrind mill media | tonnes | 588 |

| SAG Mill Liner | set/year | 1 |

| Ball Mill Liner | set/year | 0.8 |

| Primary Crusher Liner | set/year | 6 |

| Regrind Mill Liner | set/year | 2 |

| Crucibles | set/year | 50 |

| Precipitation Filter Cloth | set/year | 2 |

| Clarifier Filter Cloth | set/year | 2 |

| Baghouse Cartridges | set/year | 1 |

In general, each set of compatible reagents mixing and storage systems is located within containment areas to prevent spillage and contamination of the environment and unintended mixing with other reagents. Storage tanks are equipped with level indicators, instrumentation, and alarms to ensure spillage does not occur during normal operation. Appropriate ventilation, fire and safety protection, eyewash stations and safety data sheets are located throughout the facilities. Sumps and sump pumps are provided in the containment areas for spillage control.

17.5.1.1 Quick Lime

Quick lime (pebble size) is delivered in bulk and is pneumatically transferred from the tanker to the pebble lime silo. Pebble lime is extracted from the lime silo and fed into the lime slaking mill. Lime cyclone underflow is returned to the mill for further grinding of coarse particles, cyclone overflow reports to the lime mixing/storage tank. The lime distribution pumps move lime slurry from the lime mixing/storage tank to the flotation concentrate leaching circuit, flotation tailings leaching circuit, concentrate cyanide detoxification, and tailings cyanide detoxification circuits.

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17.5.1.2 Sodium Cyanide (NaCN)

Sodium cyanide briquettes are delivered to site in bulk ISO containers. The containers are connected to the cyanide mixing system. Fresh water is added to the cyanide mixing tank where it is heated and the pH is adjusted with sodium hydroxide. The fresh water from the mixing tank is pumped into the ISO container and circulated to the mixing tank until the cyanide briquettes have dissolved. The ISO container is emptied into the cyanide mixing tank using compressed air, empty containers are returned to the supplier for refilling. A transfer pump moves the cyanide solution to a storage tank.

Sodium cyanide is delivered to the flotation concentrate leach circuit, flotation tailings leach circuit, elution circuit, and both the tailings and concentrate Merrill Crowe circuits with dedicated dosing pumps.

A HCN gas detector and alarm system are included in the sodium cyanide reagent area to alert operators to the presence of toxic hydrogen cyanide gas. The sodium cyanide area is secured and located adjacent or inside the alkaline reagent area for easy access to alkaline pH modifiers.

17.5.1.3 Sodium Hydroxide (NaOH)

Sodium hydroxide (caustic soda) solution is delivered in 1 m3 intermediate bulk containers (IBC) tanks. Dosing pumps are connected directly to the IBC and deliver the reagent to the required locations, the strip solution tank and sodium cyanide mixing tank.

17.5.1.4 Hydrochloric Acid (HCl)

Hydrochloric acid solution is delivered in 1m3 IBC tanks. Hydrochloric acid is delivered to the acid wash circuit using the hydrochloric acid dosing pump.

17.5.1.5 Copper Sulphate (Pentahydrate)

Copper sulphate pentahydrate is delivered in solid crystal form in bulk bags. Process water is added to the agitated copper sulphate mixing tank. Bulk bags are lifted using a frame and hoist, and periodically a single bag is placed on the copper sulphate bag breaker on top of the tank. The solid reagent falls into the tank and is dissolved in water to achieve the required dosing concentration.

Copper sulphate solution is transferred by gravity to the copper sulphate storage tank, which has a stacked arrangement with the mixing tank. Copper sulphate is delivered to both the concentrate and tailings cyanide detoxification circuits using dedicated dosing pumps. An extraction fan is provided over the copper sulphate bag breaker/mixing tank to remove reagent dust that may be generated during reagent addition/mixing.

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17.5.1.6 Sulphur Dioxide (SO2)

The SO₂ storage system will be vendor-supplied and designed to feed both the concentrate and tailings cyanide detoxification circuits. Liquid SO₂ will be delivered to site via tanker truck and stored in dedicated storage tanks. The vendor package will include a control building, storage tanks, an air padding compressor, an air compressor cooling fan, a wet air receiver, a desiccant air dryer, and a dry air receiver.

17.5.1.7 Flocculant

Powdered flocculant is delivered to site in bulk bags and stored in the warehouse. Two self-contained mixing and dosing systems are installed, each including a flocculant storage hopper, flocculant screw feeder, flocculant preparation water heater, flocculant mixing tank, flocculant transfer pump, flocculant storage tank and flocculant dosing pump. Powdered flocculant is loaded into the flocculant storage hopper using the flocculant hoist. Dry flocculant is pneumatically transferred into the wetting head, where it is contacted with water.

Flocculant solution, at 0.50% w/w, is agitated in the flocculant mixing tank for a pre-set period. After a pre-set time, the flocculant is transferred to the flocculant storage tank using the flocculant transfer pumps. One system supplies flocculant to the flotation tailings thickener and the other to the concentrate thickener and CCD thickeners. Dedicated dosing pumps supply flocculant to each dosing point. Flocculant solution is diluted to 0.05% w/w concentration at each dosing point through an inline mixer with process water.

17.5.1.8 Coagulant

Coagulant is delivered to site in IBC tanks and stored in the warehouse. Dosing pumps deliver the reagent to the flotation tailings thickener as required.

17.5.1.9 Collector (PAX)

PAX is delivered in granular powder form in bulk bags. Fresh water is added to the agitated PAX mixing tank. Bags are lifted using a frame and hoist into the PAX bag breaker on top of the tank. The solid reagent falls into the tank and is dissolved in water to achieve the required dosing concentration. PAX solution is transferred by gravity to the PAX storage tank, which has a stacked arrangement with the mixing tank.

The mixing tank is ventilated using the PAX tank fan to remove any carbon disulphide gas. PAX is delivered to the flotation circuit using the PAX dosing pump.

17.5.1.10 Activated Carbon

Activated carbon is delivered on site as a granulated solid in bulk bags. The carbon is supplied to the adsorption tanks.

17.5.1.11 Frother (MIBC)

MIBC is delivered as a liquid in IBC tanks. It is used as received and without dilution. A dosing pump delivers the reagent to the required locations within the flotation circuit.

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17.5.1.12 Lead Nitrate

Lead nitrate is delivered in bags. Fresh water is added to the agitated lead nitrate mixing tank. Bags are into the bag breaker on top of the tank. The solid reagent falls into the tank and is dissolved in water to achieve the required dosing concentration.

The mixing tank is ventilated using the lead nitrate exhaust fan. Lead nitrate is delivered to both the tailings and concentrate precipitation circuits via the lead nitrate distribution pump.

17.5.1.13 Oxygen

An oxygen generation plant (50 t/d) will be leased to supply oxygen to the concentrate and tailings leach and detoxification circuits. The plant will be vendor-supplied and will include all necessary equipment for on-site oxygen production and distribution.

17.5.2 Process Water Requirements

The process plant will require fresh water for use for fire water, gland seal water, mill lubrication systems cooling, reagent mixing, dust suppression, elution and potable water production.

Process water will be used for dilution ahead of grinding and for various other uses in the plant such as spray water, flocculant dilution, flushing and hosing. Process water will be sourced from water recovered from the tailings and concentrate thickeners, the co-disposal tailings facility and the PAG tailings deposition cell. Fresh water and contact water provide any additional process water make-up requirements.

Average daily fresh and process water make-up demand is 4,195 m3/day and 29,048 m3/day respectively.

17.5.3 Power Requirements

The total installed power requirement for the process plant is estimated at 51.7 MW. Energy consumption is based on designed equipment power demand and its corresponding operation hours. Average operating load is estimated at 41 MW, or 327.3 GWh/a of energy used. This is a unit consumption of 29.9 kWh/t.

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18 PROJECT INFRASTRUCTURE

18.1 Introduction

Infrastructure contemplated in this study includes:

·	Roads: main site access road, mine infrastructure access (MIA) road, administration in-plant access roads, dike crest roads, waste rock and tailings dump haul road, explosives light vehicle access road, and ventilation raise and laydown light vehicle access road
·	Open pit mining area
·	CDF complete with co-located contact water management pond
·	West and east dikes
·	Site main gate and guard house
·	Administration and dry building, training, first aid, change house and car park
·	Control room
·	Process plant building
·	Reagent storage building
·	Merril-crowe & refinery (including gold room)
·	Assay laboratory and sample preparation area
·	Truck shop (inc. truck wash bay)
·	Plant workshop and warehouse
·	Fuel facility, fuel storage and dispensing
·	230 kV overland, overhead transmission lines
·	230 kV tie-in to Provincial grid
·	Project site sub-station
·	25 kV power distribution
·	Freshwater intake, supply, and distribution
·	Contact water collection ponds

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·	Contact water treatment plant
·	Raw water tank
·	Explosives magazine
·	Stockpile pads
·	Coarse ore stockpile cover (geodesic dome structure)
·	Permanent camp accommodations

A layout of proposed major infrastructure is included in Figure 18 1 and Figure 18 2.

Figure 18 1: Overall Site Layout Plan

Source: Ausenco, 2025

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Figure 18 2: Process Infrastructure Layout Plan

Source: Ausenco, 2025

18.2 Access

18.2.1 Site Access

Access to the Project area is described in detail in Section 5. This PFS assumes that the main access to the Springpole site will be via the new Springpole Mine Road, entering from the southeast. This road is a final leg extension of the existing Wenesaga Road, a publicly-owned forestry road currently used for most of its length by EACOM timber for forestry activities. First Mining will be responsible for constructing and maintaining the extension from the termination of Wenesaga Road to Springpole site, which will be for First Mining’s exclusive use. The main access road and location of the connection to the proposed road extension is shown above in Figure 18 1.

First Mining believes that the Project can play a meaningful role in encouraging the development of the road network in the area, with potential to connect the communities to the north with an all-season road that will provide access to the Municipality of Sioux Lookout, a major services hub for Northern Ontario. The PFS considers that the community access road may be completed prior to commencement of construction and is not part of the study.

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18.2.2 On-site Roads

Within the Project area, the main access road will provide access to the process plant site, permanent camp, and will pass alongside the gatehouse. It will be consistent with a 40 km/h design and posted speed and design criteria will be as follows: maximum longitudinal grade 8%, two lanes with a lane width of 4.0 m, road cross-slope of 3% and 1 vertical to 3 horizontal fill slope, 1 vertical to 2 horizontal back slope and 1 vertical to 3 horizontal side slopes.

Additionally, two in-plant and three maintenance roads will be constructed as follows:

·	Mine infrastructure in-plant access road - a 100 m road running E-W connecting the process plant pad to the MIA.
·	Administration in-plant access road - a short in-plant road that runs E-W connecting the process plant and Administration area.
·	Runoff pond maintenance road - a short N-S maintenance road connecting the MIA pad to the runoff pond.
·	Stockpile pad maintenance road - a N-S maintenance road, connecting the stockpile pad to the process plant area.
·	Substation maintenance road - a short E-W road, connecting the process plant to the substation pad.

In-plant and maintenance access roads will be consistent with a 30 km/h design and posted speed limit and design criteria will be as follows: maximum longitudinal grade 12%, 3% road cross-slope and 1 vertical to 3 horizontal fill slope and 1 vertical to 2 horizontal back slope and 1 vertical to 3 horizontal side slope. In-plant roads will be dual lane 3.5 m wide each, and maintenance roads will be single lane, 5 m wide.

18.3 Built Infrastructure

The main plant site, as depicted in the layout shown in Figure 18 1, includes multiple buildings. Table 18 1 lists each building along with its type and planned size.

The built infrastructure requirements are summarized by the buildings list in Table 18 1.

Table 18 1: Buildings List

Item	Size (footprint)	Comment

| Surface Mine Buildings | | |

| Mine Explosives Magazine | 10 m × 10 m × 2.7 m (100 m²) | Modular |

| Process Buildings | | |

| Primary Crushing Building | 24.9 m x 19.1 m x 19.5 m (475.6 m^2^ ) | Stick-built |

| Reclaim Tunnel | 33.0 m x 6.0 m x 7.0 m (198.0 m^2^) | Concrete structure |

| Coarse Ore Stockpile Cover | 73.0 m x 73.0 m x 35.0 (5,329.0 m^2^) | Geodesic dome structure |

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Item	Size (footprint)	Comment

| Grinding Building | 48.5 m x 44.0 m x 35.0 m (2,134.0 m^2^) | Pre-engineered |

| Process Plant Building | 88.0 m x 44.0 m x 25.0 m (3,872.0 m^2^) | Pre-engineered |

| Merrill-Crowe & Refinery Building | 54.0 m x 30.5 m x 13.0 m (1,647.0 m^2^) | Pre-engineered |

| Reagents & Services Building | 104.0 m x 30.5 m x 15.0 m (3,172.0 m^2^) | Pre-engineered |

| Laboratory | 21.5 m x 6.0 m x 2.7 m (129.0 m^2^) | Modular |

| Dewatering Barge | N/A | Details to be confirmed |

| Infrastructure/Ancillary Buildings | | |

| Mine Administration/Dry | 40.2 m x 18.3 m x 2.7 m (735.7 m^2^) | Modular |

| Plant Office Complex | 47.5 m x 18.3 m x 2.7 m (869.5 m^2^) | Modular |

| Plant Workshop & Warehouse | 36.4 m x 18.2 m x 12.0 m (662.5 m^2^) | Pre-engineered |

| Plant Truck Shop | 74.5 m x 24.5 m x 19.5 m (1,825.3 m^2^) | Pre-engineered |

| Vehicle Wash Bay | 15.7 m x 24.5 m x 19.5 m (384.7 m^2^) | Pre-engineered |

| Permanent Camp (Common Area) | 21.5 m x 18.5 m x 3.5 m (397.8 m^2^) | Modular |

| Permanent Camp (200 personnels, x5 dorm units) | 68.6 m x 8.8 m x 2.7 m (606.1 m^2^), x5 | Modular |

| Gatehouse | 10.9 m x 4.4 m (47.7 m^2^) | Modular |

18.3.1 Built Infrastructure – Surface Mine Buildings

For the surface mine, the following buildings and infrastructure have been defined to support mining operations.

·	Mine explosives magazine: Dedicated storage for mine explosives conforming to applicable explosives regulations. Dirt berms with sufficient distance to drive around and clear snow will be placed around the magazines for additional security.

18.3.2 Built Infrastructure – Process Buildings

The process buildings include the range of buildings and infrastructure required to support ore processing.

·	Primary crushing building: Stick-built structure equipped with 120 t jib crane and covered dump pocket. It houses the primary crusher and associated systems, forming the initial stage of the ore handling process. The crusher electrical room is in a separate building located close to the primary crushing building.
·	Reclaim tunnel: Concrete structure containing two stockpile reclaim apron feeders. The reclaim tunnel structure is located underneath the coarse ore stockpile.
·	Coarse ore stockpile cover: The coarse ore stockpile will be enclosed within a geodesic dome structure, serving as the receiving point for ore delivered via the overland conveyor from the crushing plant and providing interim storage prior to beneficiation in the grinding circuit.

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·	Grinding building: Pre-engineered, steel frame metal cladded building housing the SAG and ball mills. This structure includes two overhead cranes, a 60-t crane for the SAG mill and a 45-t crane for the ball mill. The grinding electrical room is in a separate building located close to the grinding building.
·	Process plant building: Pre-engineered, steel frame metal cladded structure attached to and forming a continuation of the Grinding Building. The structure is equipped with a 7.5-t overhead crane and serves as the central processing facility, accommodating the flotation and regrind circuits. The concentrate and tailings thickeners are located outside this building along with the concentrate leach and CCD circuit as well as the tailings leach/CIP circuit. The flotation/CIP electrical room is in separate building located close to the tailings leach tanks.
·	Merrill-Crowe and refinery building: Pre-engineered, steel frame metal cladded building housing the Merrill-Crowe circuit, refinery and associated equipment.
·	Reagents and services building: Pre-engineered, steel frame metal cladded building housing the reagent systems and plant services.
·	Laboratory: Modular building for housing the laboratory equipment for assaying, metallurgical, and environmental requirements. Dust-collection equipment will be located external to the laboratory building. The building will be serviced with power, water, air conditioning and heating, communications, air and mercury scrubbers, and fume hoods.
·	Dewatering barge: Stick-built structure for supporting tailing dewatering pumps at the CDF.

18.3.3 Built Infrastructure – Infrastructure Buildings

The infrastructure buildings include the structures necessary for housing and supporting site personnel. These include:

·	Mine Administration/Dry: Modular building complex which includes offices and the mine dry/changehouse areas, will have a peak day-shift occupancy of 90 personnel — 10 in the offices and 80 in the mine dry. Will be serviced with power, water, sewage, air conditioning and heating, communications.
·	Plant Office Complex: Modular building complex which includes offices and the plant dry/changehouse areas, will have a peak day-shift occupancy of 106 personnel — 52 in the offices and 54 in the plant dry (with 90 lockers).
·	Plant Workshop and Warehouse: Pre-engineered, steel frame fabric cladded building for maintenance of process equipment, as well as for the storage of equipment spare parts.
·	Plant Truck Shop: Pre-engineered, steel frame metal cladded building. This structure is separated into sections for warehousing spare parts and tool storage and the other for maintenance workshop. The building is equipped with two 25-t overhead cranes.
·	Vehicle Wash Bay: Pre-engineered, steel frame fabric cladded structure attached to and forming a continuation of the Truck Shop. This structure includes a fluid-collection sump and oil-water separator located adjacent to the truck workshop and warehouse. Wash water will be collected in sump where settling will occur and passed through oil-water separator prior to the water being recirculated back to the wash system.

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·	Permanent Camp (Common Area): Modular building containing kitchen, mess and recreational facilities for 200 personnel. Will be serviced with power, water, sewage, air conditioning and heating, communications.
·	Gatehouse: Modular security building for two personnel.
·	Potable Water Treatment: A structure to supply potable water to support personnel facilities and site operations.
·	Waste Water Treatment Plant: A plant designed to process contact water prior to reuse or discharge.
·	Sewage Treatment Plant: A treatment system supporting domestic wastewater management at the plant site. Treated effluent will be discharged to the stormwater management system.

18.3.4 Accommodation

A 450 to 500-person construction camp is envisaged. This camp will be substantially downsized and reconfigured following construction to serve as the permanent accommodation facility for approximately 200 operations personnel.

18.4 Stockpiles

The coarse ore stockpile will be enclosed within a geodesic dome structure, serving as a receiving point for ore delivered from the overland conveyor. The stockpile provides 12 h of live storage to allow for routine crusher maintenance without interrupting feed to the mill. Beneath the stockpile are two reclaim apron feeders that feed stockpiled material to the mill circuit via the SAG mill feed conveyor.

18.5 Tailings and Mine Rock Co-Disposal Facility

The Springpole Project will extract approximately 101 million tonnes (Mt) of ore, plus PAG mine rock and NAG mine rock from the open pit operation. Ore processing will generate two streams of tailings: NAG flotation tailing and lower volume of PAG sulphide concentrate tailing. Figure 18 3 shows the project site layout with the CDF where NAG tailings will be co-disposed with mine rock in the north cell and PAG tailings will be stored in the south cell. The CDF will cover an area of approximately 380 ha and will be located on land west of the open pit.

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Figure 18 3: Springpole Site Layout with Co-Disposal Facility

Source: WSP, 2024

18.5.1 Background Information

18.5.1.1 Regional Geology

The Project area was glaciated by south-southwest flowing ice during the last major glaciation referred to as the “Late Wisconsinan” glaciation (WSP, 2024). This glaciation appears to have removed all evidence of older glacial and non-glacial material and to have deposited glaciated material (silty sand till) over the bedrock formations. As the last ice sheet retreated from the region, glacial Lake Agassiz formed covering much of the area, including the Springpole Project area. Deposition of clay and silt size materials took place in deep water. As Lake Agassiz retreated from the region, innumerable depressions remained water filled. The deeper of these remain as lakes today in which organic sediments accumulated. The shallow depressions were soon partly infilled with alluvium and with an ever-increasing amount of organic material (e.g., peat). Overburden deposits in this region are relatively thin and primarily consist of glacial till, glaciofluvial deposits, glaciolacustrine sediments and organic deposits. The bedrock at the site includes mainly mafic to intermediate metavolcanic rock at the southwest zone, a narrow zone of mafic to ultramafic rock near the open pit area, and felsic to intermediate metavolcanic rocks at the northeast zone of the Project site.

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18.5.1.2 Seismicity

The Project site is located approximately at Latitude 51° 23' N and Longitude 92° 17' W within a Stable Seismic Zone incorporated in the National Building Code of Canada (Kolaj M. et al., 2020). This is a source zone of relatively low to moderate seismicity with average maximum credible earthquake of magnitude (Mw) 7.0. The Geological Survey of Canada provides ground motion parameters (peak ground acceleration and spectral accelerations for various return periods) for locations within Canada. The latest ground motion parameters corresponding to 2020 NBCC are made available by Earthquakes Canada through an online seismic hazard calculator tool. The peak horizontal ground acceleration (PGA) values obtained from the online calculator are summarized below in Table 18 2. The PGA value corresponding to the 1 in 10,000-year event was obtained by extrapolation.

Table 18 2: Peak Ground Accelerations for Springpole Project Site

Probability	10%/50 Year	5%/50 Year	2%/50 Year	0.5%/50 Year

| Return period (years) | 0.37152778 | 1:1,000 | 1:2,475 | 1:10,000 |

| PGA (g), mean values – Site Class C (Vs-30 = 760 m/s) | 0.01 | 0.02 | 0.04 | 0.131 |

Note:

1.	Extrapolated value. Source: NBCC, 2020.

18.5.1.3 Climate and Hydrology

The climate and hydrology information provided here is based on the detailed Hydrology Baseline Study (WSP, 2024). The average annual precipitation of 677 mm to 777 mm is based on Environment and Climate Change Canada stations in the region, with approximately 516 mm falling as rain and the remainder falling as snow. Most of the rainfall occurs in the summer and fall months, with peak rainfall occurring during June and July. Snow normally starts in late September and ends in early June, with peak snowfall in December. The average annual lake evaporation was estimated to be 460 mm. The coldest month is January with average temperature of 18.3°C and the warmest month is July with average temperature of 18.1 °C. Temperatures fall below freezing from November through March. The 72-hour Probable Maximum Precipitation (PMP) was calculated to be about 400 mm (Knight Piésold, 2021).

18.5.1.4 Mine Waste Production Schedule

The Project mine plan and waste production (including separated PAG, NAG and low Rock Quality Designation (RQD) NAG mine rock tonnages) are summarized in Table 18 3. Approximately 20% of tailings (20.2 Mt) will be PAG tailings and remaining 80% will be NAG tailings.

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Table 18 3: Annual Mine Waste Production Schedule

Year	PAG Mine Rock	NAG Mine Rock	Low RQD NAG Mine Rock	Tailings

| | (Mt) | (Mt) | (Mt) | (Mt) |

| Pre-production Year 1 | 8.8 | 4.2 | 0.4 | 0 |

| Year 1 | 14.9 | 12.6 | 2.9 | 10.95 |

| Year 2 | 14.8 | 12.1 | 3.6 | 10.95 |

| Year 3 | 18.4 | 9.2 | 1.5 | 10.95 |

| Year 4 | 18.4 | 17.6 | 9.3 | 10.95 |

| Year 5 | 20.3 | 15.2 | 12 | 10.95 |

| Year 6 | 18.1 | 11 | 11.5 | 10.95 |

| Year 7 | 23.9 | 11.8 | 9.1 | 10.95 |

| Year 8 | 7.4 | 1.4 | 1.3 | 10.95 |

| Year 9 | 0.6 | 0 | 0 | 10.95 |

| Year 10 | 0 | 0 | 0 | 2.45 |

| Totals | 145.5 | 95.2 | 51.7 | 101 |

18.5.1.5 Tailings – Physical Characteristics

Laboratory test programs (Knight Piesold, 2021 and FLSmidth, 2022) were completed assessing physical and mechanical properties of tailings. The laboratory tests results are summarized in Table 18 4. The tailings are classified as a non-plastic inorganic silt with a low permeability when compacted. The consolidated undrained shear strength is typical for an inorganic silt.

Table 18 4: Physical and Mechanical Tailings Properties

Description	Value

| Specific Gravity | 2.7 – 2.8 |

| Fines Content (< 0.074 mm; %) | 53 – 68 |

| Liquid Limit (%) | 26 |

| Plastic Index (%) | Non-Plastic |

| USCS classification | ML (Inorganic Silt) |

| Hydraulic Conductivity (m/s) | 1x10^-7^ |

| Consolidated Undrained Triaxial – Friction Angle (o) | 31 |

| Standard Proctor Maximum Dry Density | 1.7 – 1.9 |

18.5.1.6 Tailings – Geochemical Characteristics

Geochemical test programs have been conducted to assess the metal leaching and acid rock drainage (ML/ARD) potential of the tailings. Tailings samples included in the program were produced as part of bench scale metallurgical testwork conducted for the Project from 2021 to 2022. The test program included a suite of static and kinetic tests to evaluate short-term static conditions and the long-term potential for acid rock drainage and metal leaching.

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Static testing (acid base accounting, net acid generation testing, elemental content analysis, short term leaching tests, and mineralogical testing) was conducted on 18 synthetic tailings samples (including 9 samples each of NAG and PAG tailings). Kinetic testing (humidity cells, subaqueous column tests) was conducted on four samples of NAG tailings and one sample of PAG tailings. Kinetic testing programs have been ongoing for several years and the PAG tailings test and three of the NAG tailings tests continue to operate. Water quality analysis was conducted for synthetic tailings decant water (supernatant) samples.

Results of static testing indicated that flotation tailings are non-potentially acid generating (NPR>2) with low sulphur contents (~0.15%) and a low solid phase metal content. Humidity cell testing indicated no potential for ARD and a low potential for metal leaching from the NAG tailings samples. Flotation tailings will be stored in the north cell of the CDF, co-disposed with PAG and/or metal leaching (“ML”) mine rock. The PAG mine rock co-disposed within the CDF will be encapsulated with tailings prior to the onset of net acidic conditions in the rock. To maintain stable geochemical conditions over the long term, a suitable thickness of NAG tailings will be required to be placed over the underlying PAG/ML mine rock toward the end of mine life. Geochemical programs and modelling are proposed to further evaluate these design requirements and support deposition planning and closure concepts.

Sulphide concentrate tailing samples were confirmed to be PAG (NPR<1) and metal leaching. Due to their high sulphide content, the tailings are assumed to have a short lag time to ARD and they will require deposition under a water cover immediately at discharge in the south cell of the CDF. The PAG tailings will require permanent isolation from oxygen over the long term to maintain stable geochemical conditions. Geochemical programs and modelling are proposed to further assess the potential performance of closure cover concepts for the PAG tailings.

Water treatment of the tailings discharge from the mill will be used to manage process water quality discharged with the tailings to the CDF during mine operations.

18.5.1.7 Mine Rock – Geochemical Characteristics

Baseline geochemical programs have been conducted for the Project since 2012 to evaluate the ML/ARD potential of Project rock in support of baseline studies and ongoing engineering studies.

Geochemical programs have included a suite of static and kinetic tests to evaluate short-term static conditions and the long-term potential for ML/ARD. The static testing assessment includes on the order of 900 drill core samples and has involved several laboratory analytical methods, including acid-base accounting, net acid generation testing, elemental content analysis, shake flask extraction testing, and mineralogical testwork. Kinetic testing has been ongoing for several years and has included humidity cell tests (16 tests), column tests (13 tests) and field cell tests (10 tests). The field tests continue to be monitored and several humidity cell tests continue to operate.

Results indicate that both PAG (NPR<2) and NAG (NPR>2) mine rock is present at the Project. An NPR threshold of two is being used for current purposes and is supported by available kinetic test results.

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The sulphur contents of the samples were on the order of 1%, typically present as pyrite. Samples showed a wide range of sulphur contents (<0.005% to 10%). The NP content of the rock was variable, with some samples showing low levels of NP (5 to 20 kg/t) and other samples showing relatively higher levels of NP (100 kg/t), primarily present as carbonate minerals. The geochemical characteristics of the samples were not strongly dependent on rock type. Some spatial trends were observed and were attributed to varying mineralization/alteration styles present in the rock. Samples from the Portage zone generally had a higher sulphur content and were more often classified as PAG relative to samples from the East Extension and Camp zones which were often NAG and had a relatively lower sulphur content. Low NP samples were primarily present in the Portage zone, whereas samples with a higher NP were present in all three zones.

Lag times to acid onset for PAG samples ranged from less than one year to several years for samples with a low NP content, to one to several decades for samples with a higher NP content. For current planning purposes, PAG rock with an NP content <20 kg/t is assumed to require prompt management in the CDF to prevent acidification. Additional programs to refine the lag time to ARD for PAG rock are proposed.

The geochemical assessment indicated that NAG mine rock included both arsenic-leaching and non-arsenic leaching NAG rock. The arsenic leaching potential of the NAG rock appeared to be related to the solid phase arsenic content of the samples and varying arsenic host mineralogy (e.g., arsenian pyrite, arsenopyrite, gersdorffite). Other metals did not appear to be of potential risk for neutral leaching.

Test program results were used to define a solid phase arsenic threshold to differentiate between arsenic leaching (≥50 mg/kg) and non-arsenic leaching samples (<50 mg/kg), in support of identifying NAG rock that is geochemically suitable for use in construction from that which may require additional management. One year of monitoring from field test cells validated this threshold and suggested that a higher threshold value for non-arsenic leaching NAG rock may be achievable. The field tests continue to be monitored. Additional programs to validate the arsenic threshold are proposed.

18.5.1.8 Subsurface Conditions at CDF

Geotechnical site investigation programs carried out at the CDF and the surroundings include:

·	2020 geotechnical investigation program by Fracflow (Fracflow, 2020)
·	2021 geotechnical investigation program by Ausenco (Ausenco, 2022)
·	2022 geotechnical investigation program by Knight Piésold (KP, 2022)
·	2022 Supplemental Geotechnical Investigation (WSP, 2023b)

Figure 18 4 shows a location plan of the boreholes, monitoring wells and test pits completed at the CDF area. Based on the site investigation data, overburden at the CDF area generally consists of peat, sand to silt, silt to clay and sand and gravel deposits. Overburden thickness at the CDF dam footprint area is generally 2.5 m and less, except for the limited zone at the southwest corner of the CDF footprint and limited localized low-lying areas. Bedrock at the CDF footprint is generally consistent and composed of mafic to intermediate metavolcanics with average shallow bedrock permeability of 5.7 x 10-8 m/s and deep bedrock permeability of 1.6 x 10-8 m/s. In general, subsurface conditions at the CDF area are favourable with limited thickness of overburden and low permeability bedrock.

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Figure 18‑4: Boreholes, Monitoring Wells and Test Pits Locations

Source: WSP, 2025

18.5.2 Design Basis, Requirements and Criteria

The CDF is designed as a two-cell facility consisting of north and south cells, contained by perimeter dams and separated by an internal dam. PAG slurry tailings will be deposited within the south cell under water cover to limit oxygen ingress and prevent acidification. NAG tailings and PAG/NAG/ML mine rock will be co-disposed within the north cell. Non-metal leaching NAG mine rock (NAG/NML) will be used for CDF perimeter dam construction.

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Adequate elevation difference will be maintained between the north and south cells so that all runoff and tailings water reports to the south cell by gravity. Allowance for the operating pond and containment of the Environmental Design Flood (EDF) will be made within the south cell. Considering the yearly rate of rise of CDF, the spillway will be constructed at closure. During the operational phase, the CDF will be designed to contain the Inflow Design Flood (IDF).

The minimum design offset from Springpole Lake and Birch Lake is to be 120 m in accordance with the Ontario MNR shoreline reservation policy.

The minimum design offset from open pit is also assumed to be 120 m as the open pit will be reclaimed at closure and reconnected to Springpole Lake. An allowance of 60 m is considered for the seepage and runoff collection system and perimeter access road around the CDF. The proposed CDF layout with offset areas is shown in Figure 18‑5.

The proposed CDF layout provides setback ranging from 120 m (minimum offset) to as much as 300 m from waterways providing additional allowance for environmental contingency measures, if required.

The south cell perimeter starter dams and subsequent raises will be designed as downstream raised. The upstream slope will be lined with low permeability liner (e.g., reinforced geosynthetic clay liner; GCL) and anchored to low permeability bedrock and/or overburden. Higher hydraulic conductivity bedrock zones along the GCL anchor will be grouted as required.

The north cell perimeter starter dams will be designed as downstream raised, with subsequent raises as centreline raises. The north cell perimeter dams will be designed as pervious dams. The lack of a permanent pond in the north cell and presence of wide tailings zone upstream of the dams will limit seepage losses.

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Figure 18‑5: CDF Layout with Offset Constraints

Source: WSP, 2025

18.5.2.1 Waste Storage Requirements

Based on the mine production plan, CDF waste storage volume requirements are calculated and summarized below (Table 18‑5). PAG tailings, NAG tailings and PAG mine rock are to be stored and managed within the CDF.

As discussed later, NAG/NML mine rock will be used for construction of CDF dams and other site infrastructures.

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Table 18‑5: CDF Waste Storage Volume Requirements

Year	Annual Tailings (Mt)	PAG Tailings<br> <br>(Mm^3^)		Annual NAG Tailings (Mm^3^)	Annual PAG Mine Rock (Mt)	Annual PAG<br> <br>Mine Rock<br> <br>(Mm^3^)	PAG Mine Rock +<br> <br>NAG Tailings (Mm^3^)

| PP1 | 0 | - | - | - | 8.8 | 4.0 | 4.0 | 4.0 |

| Y1 | 10.95 | 1.7 | 1.7 | 6.7 | 14.9 | 6.8 | 13.5 | 17.5 |

| Y2 | 10.95 | 1.7 | 3.4 | 6.7 | 14.8 | 6.7 | 13.5 | 31.0 |

| Y3 | 10.95 | 1.7 | 5.1 | 6.7 | 18.4 | 8.3 | 15.1 | 46.0 |

| Y4 | 10.95 | 1.7 | 6.7 | 6.7 | 18.4 | 8.3 | 15.1 | 61.1 |

| Y5 | 10.95 | 1.7 | 8.4 | 6.7 | 20.3 | 9.2 | 16.0 | 77.1 |

| Y6 | 10.95 | 1.7 | 10.1 | 6.7 | 18.1 | 8.2 | 15.0 | 92.1 |

| Y7 | 10.95 | 1.7 | 11.8 | 6.7 | 23.9 | 10.9 | 17.6 | 109.7 |

| Y8 | 10.95 | 1.7 | 13.5 | 6.7 | 7.4 | 3.4 | 10.1 | 119.8 |

| Y9 | 10.95 | 1.7 | 15.2 | 6.7 | 0.6 | 0.3 | 7.0 | 126.8 |

| Y10 | 2.45 | 0.4 | 15.5 | 1.5 | 0.0 | 0.0 | 1.5 | 128.3 |

Notes:

1.	Nominal plant throughput: 30,000 tonnes per day (First Mining).

| 2. | PAG tailings is 20% and NAG tailings is 80% of total tailings production. |

| 3. | Specific gravity of tailings solids: 2.8 |

| 4. | Settled dry density and void ratio of deposited tailings: 1.3 tonnes per cubic metre (t/m3) and 1.1 (assumed). |

| 5. | Specific gravity of waste rock: 2.7 (assumed). |

| 6. | Average dry density of Mine Rock (PAG & NAG): 2.2 t/m3 (assumed). |

18.5.2.2 Design Standards

Applicable requirements from the following regulations, guidelines, and standards have been adopted for the CDF design and analyses:

·	Dam Safety Guidelines 2007 (2013 Edition). Canadian Dam Association (CDA 2013).
·	Technical Bulletin: Application of Dam Safety Guidelines to Mining Dams (CDA 2019).
·	Global Industry Standard on Tailings Management (GISTM 2020).
·	Technical Bulletins, Ontario Environment and Energy, Dam Management (Lakes and Rivers Improvement Act).

The CDF Pre-Feasibility design has applied the highest design criteria from the above guidelines and standards.

The key design requirement, criteria and assumptions are summarized below:

·	Production

o	Ore resources: 101 Mt.

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o	Nominal plant throughput and LOM: 30,000 t/d, 9.5 years.
o	Mine rock: 292.5 Mt.
o	Mining ends about two years prior to cessation of milling.

·	Waste Management Requirements

o	Tailings:

■	Potentially acid generating (PAG) tailings (20% of total): 20.2 Mt.
■	Non-acid generating (NAG) tailings (80% of total): 80.8 Mt.
■	Tailings disposal technology: south cell conventional slurry, north cell thickened slurry, centrifugal pumping/gravity discharge.
■	Average in-situ tailings settled dry density: 1.3 t/m3 (calculated, assuming void ratio of 1.1, Gs of 2.8).

o	Mine rock:

■	PAG mine rock: 145.5 Mt.
■	NAG mine rock: 147.0 Mt.
■	Dry density of mine rock (NAG & PAG): 2.2 t/m3 (assumed).

·	Design Criteria and Assumptions

o	Design earthquake: 1 in 10,000-year seismic event.
o	Inflow design flood (IDF) of Probable Maximum Precipitation (PMP).
o	Static and seismic stability criteria for embankment dams:

■	Static Factor of Safety >1.5.
■	Pseudo-static Factor of Safety >1.0 or acceptable deformation.
■	Post-earthquake Factor of Safety >1.2.

o	Allowance for operating pond and containment of IDF within the south cell during operation. Emergency spillway is to be constructed at closure.
o	Minimum design offset from waterways and waterbodies to be 120 m as per environmental commitments.
o	Minimum design offset from open pit is to be 120 m.
o	Allowance for CDF perimeter ditch (for seepage and runoff collection) and perimeter access road is 60 m.

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18.5.3 CDF Ultimate and Starter Dams Sizing

The CDF footprint was established considering the offset requirements and is shown in Figure 18‑5 with offset constraints. Height of the CDF north and south cells ultimate dams were established considering the tailings and mine rock storage requirements and allowances for water management as summarized below:

·	North Cell: Storage volume (NAG tailings and PAG mine rock): 125 Mm3 (excluding PAG mine rock required for internal dam).
·	South Cell: Storage volume (PAG tailings): 15.5 Mm3 and total of 9 m allowance for subaqueous tailings deposition, operating pond and EDF and IDF storage.

An estimated allowance of 2 m was assumed to contain IDF (PMP) within the south cell. The 72-hour PMP is 400 mm (Knight Piésold, 2021), a longer duration PMP will be considered as part of future designs which may require more or less than a 2 m allowance.

The CDF with perimeter and internal dams were sized considering waste storage and water management requirements. North and south cell perimeter dams were established at elevation 485 m and 480 m, which correspond to average dam heights of approximately 76 m and 75 m respectively. The starter dams crest elevations were established at 437 m (for both north and south cells) to accommodate 24 months of production.

18.5.4 Tailings and Mine Rock Management

South Cell: The south cell will be contained by lined perimeter dams and the unlined internal dam. PAG tailings will be subaqueously deposited using a floating pipeline or barge within the south cell to maintain a water cover to prevent acid generation. Following completion of PAG tailings deposition, NAG tailings or other suitable soil cover will be placed to remove excess pond capacity and provide suitable erosion protection cover and vegetative cover over the PAG tailings. These cover materials will isolate the tailings from oxygen.

North Cell: The north cell will be contained by unlined perimeter dams and the internal dam. NAG tailings and PAG/metal leaching NAG mine rock will be co-disposed within the north cell. PAG and NAG/ML mine rock will be trucked from the open pit and placed within the north cell developing a mine rock stockpile generally in the middle of north cell and away from the perimeter dams. NAG tailings will be spigotted from the perimeter dams to develop a low permeability zone of tailings against the perimeter dams and enclose the mine rock with a low permeability NAG tailings zone to limit oxygen ingress. Mine rock placement and tailings discharge need to be actively managed to achieve effective development of wide NAG tailings zone enclosing the PAG mine rock. The PAG mine rock will be encapsulated in NAG tailings prior to acidification of the mine rock.

18.5.5 Water Management

All runoff from the surface of the CDF (both north and south cells) will be directed to and managed within the south cell. The surface of the north cell will be graded to sustain surface drainage from the north cell through the internal dam and to the south cell, as the facility expands and dams are raised throughout the course of operations. Runoff from the outside slopes of the perimeter dams, as well as interflow and seepage through and beneath the dams will be captured by perimeter collection ditches and ponds. The water collected in the perimeter collection ponds will be considered contact water and pumped back into the CDF (see Figure 18‑2).

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Water collected in the south cell pond will be reclaimed to the plant/mill, reducing the demand for freshwater from Birch Lake. Excess water will be pumped to the central water storage pond for monitoring, treatment, and discharge to environment as needed.

Construction of the CDF perimeter dams will be advanced at the rate necessary to maintain sufficient storage above the tailings surface to accommodate an aggregate of the following storages throughout operations (approximately 9.5 years):

·	A minimum water cover required to deposit PAG tailings subaqueously and an operational volume to account for typical seasonal fluctuations.
·	The EDF to prevent direct discharge to the environment.
·	The IDF (PMF) to prevent over-topping of the perimeter dam.
·	Freeboard between the maximum IDF (PMF) water level and the perimeter dam crest.

As the height of the CDF approaches its ultimate configuration in the later years of operation, a permanent overflow spillway will be constructed. The overflow spillway will have sufficient capacity to safely convey the IDF (PMF).

A site wide water balance has been prepared incorporating the CDF (WSP, 2023f). This will be refined at later design stages as additional detail for the mine plan and tailings and mine rock deposition plans are advanced.

18.5.6 Closure Concept

The CDF closure concept involves (a) continue directing run-off from the north cell to the south tailings, (b) maintain a minimum pond (or no pond), placement of a cover to prevent oxygen ingress into the PAG tailings in south cell; and (c) implement an overflow spillway at the south cell to safely convey excess water to environment. Preparing the CDF for closure will involve the following:

·	Constructing the overflow spillway at the south cell perimeter dam to safely manage extreme runoff events.
·	Following completion of PAG mine rock disposal within the north cell, depositing a suitable thickness of NAG tailings over the entire north cell surface to fully cover the PAG mine rock and limit oxygen ingress.
·	Placing a vegetation cover or if necessary, erosion protection, over the north cell surface to direct all runoff to the south cell.
·	Placing a layer of NAG tailings of suitable soil cover, following completion of PAG tailings deposition within the south cell and upon final closure of the CDF, to provide a vegetative and erosion resistant cover over the PAG tailings. The cover will be designed to promote the long-term geochemical stability of the underlying PAG tailings.
·	Decommissioning the perimeter ditch collection system and ponds to allow runoff and seepage water to report to environment once water quality requirements are met.

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18.5.7 CDF Dam Design and Analyses

The CDF perimeter dam and internal dam design configuration, stability and seepage analyses are summarized in the following sections. The CDF layout and typical sections of the perimeter and internal dams are provided in Figure 18‑5 and Figure 18‑6.

The shell of the perimeter dams will be construction from run of mine directly hauled to the dams. The upstream filters and GCL (south cell) would be placed by specialty equipment/workers. This will require year-round (12 month) construction.

18.5.7.1 South Cell Perimeter Dam

The south cell perimeter dam is designed as a downstream raised rockfill dam with an upstream low-permeability liner (e.g., reinforced GCL). The liner is to be anchored to a concrete plinth constructed on bedrock along the upstream toe or anchored into suitable low-permeability overburden. Bedrock grouting is to be implemented as required to reduce bedrock permeability. It is expected that the downstream dam raise construction of coarse to fine rockfill zones can be implemented year around. Construction of filter, GC liner and protection layer can be implemented during the summer months.

The dam is to be founded directly on bedrock or competent overburden, except for the southwest corner. The southwest corner of the CDF contains thick silt to clay overburden which will require excavation or ground improvement. For the purpose of material take-offs, ground improvement is assumed. Additional investigation of this area is planned at the next stage of design to confirm the optimum design solution. The proposed typical section of the south cell perimeter dam is provided in Figure 18‑6.

18.5.7.1.1 Seepage and Stability Analyses

Steady state seepage analysis of the typical section was performed considering the ultimate dam closure configuration and hydraulic conductivity parameters for various material zones based on available site investigation data, and literature. The seepage analyses show that on the order of 90% of seepage would be captured with a perimeter seepage collection system.

Stability analyses were performed for long-term and 1 in 10,000-year design earthquake loading conditions. Material parameters for various zones were assumed based on available site investigation data, and literature.

The stability analyses determined that the required factors of safety criteria are met or exceeded as required by CDA (2013).

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18.5.7.2 North Cell Perimeter Dam

The north cell perimeter dam is designed as a centreline raised rockfill dam with suitable transition and filter zones from the tailings to the rockfill shell. The designed dam is to be founded directly on bedrock or competent overburden. The proposed typical section of the north cell perimeter dam is provided in Figure 18‑7.

18.5.7.2.1 Stability and Seepage Analyses

Steady state seepage analysis of the typical section was performed considering the ultimate dam closure configuration and using hydraulic conductivity parameters for various material zones based on available site investigation data, and literature. The seepage analyses indicate that in the order 90% of seepage would be captured with a perimeter seepage collection system.

Stability analyses were performed for long-term and 1 in 10,000-year design earthquake loading conditions. Material parameters for various zones were based on available site investigation data, literature and design criteria. The stability analyses results indicate that required minimum factors of safety criteria are met or exceeded as required by CDA (2013).

18.5.7.3 Internal Dam

The internal dam between the north and south cells is designed as a centreline raised rockfill dam. Considering the presence of tailings and/or mine rock buttressing on both sides of the dam, stability analyses were not performed for this design update. The internal dam is to allow seepage across from north cell to south cell. It may be required to incorporate a low permeability zone within the internal dam initial lift to contain south cell pond until adequate tailings deposited on either side of the internal dam.

The internal dam may have an internal spillway or pipe culvert to allow surface runoff from north cell to freely flow into the south cell. The internal spillway or pipe culvert will also allow flow back into north cell during EDF and IDF events.

18.5.8 Instrumentation and Monitoring

Instrumentation and monitoring will be required to ensure performance of the CDF for stability and geochemical performance (water quality). An allowance has been made in the capital cost estimate.

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Figure 18‑6: Proposed Ultimate CDF Configuration

Source: WSP, 2024

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Figure 18‑7: Proposed CDF – Profiles and Sections

Source: WSP, 2025

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18.5.9 Construction Rock Mass Balance

NAG/NML rock for construction can be provided from the open pit excavation, CDF quarry and from excavation of the fish habitat development area. If additional rock is required to meet the requirements and/or because mine rock production may not match the construction schedule, additional rock could be sourced from the following areas:

·	Expansion of the fish habitat development area compensation area both during initial construction and closure.
·	Quarry within the CDF footprint would be available during initial construction.
·	Expansion of the open pit would be available during operations and at closure.

For current purposes, 17 Mm^3^ of additional rockfill is assumed to be quarried from the CDF footprint.

18.6 Dikes

18.6.1 Design Basis

During the first year of the site development/construction phase, two dikes (west dike and east dike) will be constructed to isolate the open pit basin within Springpole Lake. Figure 18‑8 shows the general site plan for the Project and the locations of the proposed two dikes. The design of the dikes will address the following requirements:

·	Establish a hydraulic barrier between Springpole Lake and the open pit basin area following controlled dewatering.
·	Provide sufficient freeboard above the lake level for protection during storm events for worker safety, along with protection from ice heaving that may occur in the winter.
·	Maintain stability during and following open pit basin development (meeting or exceeding regulatory factor of safety requirements).
·	Minimize the footprint of the open pit area.
·	Minimize the required area to be isolated from Springpole Lake.
·	Provide for a safe open pit work area.
·	Provide enhanced fish habitat on the downstream side to support fish habitat offset goals.

During closure phase, the open pit basin will be filled with water near final water level, and once the water quality meets all regulatory requirements, lowering of the dikes in a controlled manner and reconnection of the reclaimed basin to Springpole Lake.

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Figure 18‑8: Location of West and East Dikes, and Limits of Bay Dewatering

Source: WSP, 2024.

18.6.2 Dike Design

Figure 18‑9 shows a sketch of a typical dike cross-section. The dikes will have a 28 m crest width to accommodate bulk rockfill placement using the mine fleet for three-point turning and dumping, with upstream and downstream slopes constructed at the natural angle of repose. The dikes include sufficient freeboard above the elevation of Springpole Lake to prevent flooding during storm events (i.e., 5 m above the average lake level elevation of approximately 391 m amsl). In situ monitoring has shown that water levels in Springpole Lake vary by approximately 2.1 m, with a peak water elevation of approximately 392.8 m amsl, and a minimum water level of approximately 390.7 m amsl. This information was used to predict the average lake level and the dike freeboard height in accordance with the Canadian Dam Association Dam Safety Guidelines (CDA 2013, 2019). Ongoing monitoring of water levels in Springpole Lake will continue during all phases of the Project. The information will be used to verify the required dike’s crest elevation.

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The dike embankment will be constructed of three primary material zones, as shown on Figure 18‑9:

·	Zone 1 is a 25 mm minus rock fill material used for construction of the central core zone of the embankment. Zone 1 material is required though the core of the dikes to facilitate installation of slurry wall and minimize potential slurry loss.
·	Zone 2 material is a 100 mm minus rock fill material used for construction of the embankment shell.
·	Riprap will be placed in a 2 m thick layer on the upstream faces of the dikes to provide erosion protection.

Dikes will be provided with two seepage control elements as follows:

·	A seepage cut-off wall (Slurry Wall) will be installed within the embankment Zone 1 crushed rockfill to minimize seepage from Springpole Lake through the structure into the open pit basin.
·	A foundation grout curtain will also be installed to minimize seepage through the shallow bedrock zone and low RQD zone.

The following outline the key steps for dike construction:

Foundations within the dike footprint will be prepared by clearing organics and trees on land and removing soft and/or loose lakebed materials in water through dredging or long-reach excavation. Initial embankment placement will involve Zone 1 and Zone 2 materials to develop a working platform for slurry wall installation. Zone 1 will then be vibro-compacted to improve density and support the trench construction. The slurry wall will be constructed by installing a guide wall, excavating the slurry trench, and backfilling with cement-bentonite or plastic concrete. Bedrock grouting will further reduce seepage. Finally, riprap will be placed upstream and the embankment will be raised to its final crest elevation.

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Figure 18‑9: Dike Typical Section

Source: WSP, 2024

18.6.3 Hazard Classification

The design standards for the dikes are based on the relevant federal and provincial guidelines for construction of mining dams in Canada. The following regulations and guidelines were used to determine the dam hazard classification and suggested minimum target levels for some design criteria, such as the IDF and EDGM:

·	Technical Bulletin – Application of Dam Safety Guidelines to Mining Dams (CDA, 2019).
·	1990 Ontario Lakes and Rivers Improvement Act (LRIA, 2011).

The dikes have been classified as Very High (KP, 2021) under both CDA guidelines and the LRIA. The recommended IDF during operations is defined as 2/3 between the 1/1000-year return period flood and the PMF for a Very High dam classification. A 72-hour event will be considered, given the dikes will not be constructed with spillways (during operations) and will need to provide temporary ‘storage’ as Springpole Lake responds to the flood event.

The recommended design earthquake is characterized as halfway between the 2,475-year and 10,000-year return period seismic events for a Very High dam classification.

18.6.4 Monitoring

Instrumentation and monitoring will be required to assess dike embankment performance. Vibrating wire piezometers will be installed to monitor pore pressure within the foundation materials and slope inclinometers and survey monuments will be installed to monitor slope movement and deformation.

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18.6.5 Environmental

A turbidity barrier will be installed on both the upstream and downstream side of the dike alignment prior to placement of dike embankment fill material. It is expected that the upstream turbidity barrier will remain in place following construction, isolating Springpole Lake from any sediment laden runoff from the embankments.

18.7 Water Management

18.7.1 Water Management Plan

A water management plan for the operations phase of the Project was developed to describe the way site contact water will be collected, contained, treated and discharged to the southeast arm of Springpole Lake (the proposed receiving environment). A comprehensive review of site topography was completed to determine the location of ditches and local collection ponds to minimize the mine contact water footprint and prevent uncontrolled discharge to the environment. The water management plan considers Project-specific design criteria, site limitations and opportunities, as appropriate. The purpose of the Plan is to describe how site contact water will be collected, contained, treated and discharged to the southeast arm of Springpole Lake (the proposed receiving environment). Key water management infrastructure includes:

·	CDF internal pond
·	CWSP
·	Open pit watershed storages

The water management system design uses standard engineering criteria for ditches, water storage ponds, and any necessary emergency spillway. Storage ponds and water management structures are designed to manage the EDF without discharge of untreated water to the environment. An environmental design flood (“EDF”) is a hypothetical flood (peak discharge or hydrograph) adopted as the basis in the engineering design of project components to prevent uncontrolled release of affected water. For the operations phase, the EDF has been defined as a flood event with a 1:100-year return period, which is a typical requirement for mines in Ontario. Durations for this return period included:

·	high-intensity shorter duration 24-hour event;
·	30-day rain on snow; and
·	365-day cumulative rain or water equivalent depth.

To meet the above design criteria, surface water management infrastructure such as ditching, berms and pumps are required to convey contact water to water storage facilities for re-use or for treatment and discharge to the environment in accordance with applicable regulatory requirements. The mine site runoff collection systems will also be designed to a 1:100 year flood event. Typically, the design of mine site water management systems in Northern Ontario is governed by the spring freshet (which is a long duration event, lasting several weeks) or a summer rainstorm (which is a shorter period, ranging from several hours to several days). For the ditch sizing, a short duration storm event will produce the largest peak flow and therefore govern the sizing. Conveyance requirements for the collection ditches were also conservatively developed to convey the peak flow from the 100-year flood event.

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A layout of planned surface water management infrastructure is provided in Figure 18‑10 with the resulting operations phase sub-watersheds.

18.7.2 Construction Water Management

Facilities described in the Operations Water Management will be developed as needed to support the water management during the construction phase. Additional temporary ditching, sumps and ponds not described in that section may also be used to facilitate water collection and transfer during construction. Water management during construction is designed to accommodate appropriate design storms reflective of the shorter duration (two to three years) associated with construction. Controlled dewatering of the open pit basin during construction will initially involve the transfer of lake water out of the isolated basin back into Springpole Lake on the downstream side of the dikes within the north basin. During initial controlled dewatering of the open pit basin, it is anticipated that water quality will not exceed water quality guidelines, such that direct discharge to Springpole Lake can occur.

A detailed water quality monitoring program will be implemented to demonstrate this during the dewatering activity. The initial dewatering will continue until the threshold for TSS in the discharged water is approached and likely to be reached (likely 15 mg/L monthly average or 30 mg/L single grab sample, according to MDMER Schedule 4). Once the TSS threshold is approached/reached, the remaining water in the isolated open pit basin will be directed to a settling pond and/or other form of treatment (e.g., clarification, filtering, flocculation) to reduce TSS prior to discharge to the receiving environment.

18.7.3 Operations Water Management

Contact water arising from precipitation and groundwater is collected in ditches, sumps and ponds and transferred into the integrated site water management system for containment, treatment and discharge to the environment in accordance with applicable regulatory requirements, as needed. The water management strategy is to collect site runoff in local collection ponds within each sub-watershed. The largest ponds are the CDF internal pond, CWSP and ponds located within the open pit sub-watershed. The water collected in these ponds is considered as contact water and requires treatment through the effluent treatment plant (ETP) prior to discharge to the environment. Designs and locations for perimeter ditching and ponds consider distances from nearby infrastructure and natural waterbodies and maintain setbacks from these features. For example, perimeter collection ponds will be strategically located in the topographic low points surrounding the CDF.

The storage requirements for the major water storage is based on the EDF (1:100-year return period). For the pond sizing, the ability to contain various durations of the 1:100-year event is a function of the available storage and pumping capacity. The minimum necessary pump rates were estimated such that the EDF volume is pumped out within one year. These minimum pumping rates and associated storage requirements adopted for the water management plan are summarized in Table 18‑6. Additionally, the design of the dikes includes 5 m of freeboard (height above the lake level). The freeboard provides a reliable buffer to accommodate waves and ice movement, as well as natural year-to-year lake level fluctuations and major precipitation events. Should ongoing lake level monitoring indicate an increasing trend during operations, the crest height of the dikes can be raised to provide additional contingency to safely continue operations.

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The IDF is defined as the largest runoff event that a facility is designed to safely withstand and prevent overtopping of the water containment structures. The IDF for the CDF will be defined as the Probable Maximum Flood (PMF). The 72-hr Probable Maximum Precipitation was calculated to be about 400 mm by Knight Piésold (March, 2021) and has been used as the IDF criteria for the internal pond in the CDF.

The integrated site water management system for the operations phase includes the key water collection locations and infrastructure described below, most of which are shown schematically in Figure 18‑11.

18.7.3.1 Co-Disposal Facility Internal Pond

The CDF internal pond in the south cell collects water from both the north and south cells and from CDF perimeter seepage collection ponds. Water collected in the CDF internal pond will be reclaimed to the plant/mill, reducing the need for freshwater demands from Birch Lake. Excess water will be pumped to the CWSP for monitoring, treatment, and discharge to the environment in accordance with applicable regulatory requirements to environment, as needed.

The CDF internal pond will require approximately 1.4 Mm^3^ of active storage assuming a minimum pumping rate of 100 cubic metres per hour (m^3^/h) (to the CWSP) and reclaim rate of 1,178 m^3^/h (to the process plant). These rates are necessary to manage the 1:100-year event within one year. The 1.4 Mm^3^ storage is required in addition to the following storage that will be considered at a later engineering/design stage:

·	The maintenance of saturated tailings conditions to prevent acid generation.
·	An operational volume to account for typical seasonal fluctuations.
·	Freeboard between the maximum IDF water level and the dam crest.

18.7.3.2 Central Water Storage Pond

The CWSP is the ultimate collection point for contact water and will provide make-up water to the process plant as needed. Excess water will be pumped to the effluent treatment plant (ETP) for treatment, and subsequently discharged to the environment in accordance with applicable regulatory requirements. The storage required to contain the EDF is estimated to be approximately 0.7 Mm^3^, assuming a minimum discharge/treatment rate of 950 m^3^/h required to manage the 1:100-year event. Higher treatment rates may be considered to reduce the storage required and optimize the operating ranges within the pond. Based on bathymetric data for the CWSP and an assumed water surface elevation of 393.0 m amsl (from 2020 LiDAR survey), the CWSP storage capacity is estimated to be approximately 1 Mm^3^.

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18.7.3.3 Open Pit Basin

The open pit basin watershed storage will include temporary ponds to provide storage and house the dewatering pumps. Additional temporary ponds or ditching may also be provided in the open pit basin to help control runoff entering the pit. The combined open pit basin contact water (surface and groundwater) will be pumped from the sumps to the CWSP. A combined storage within the open pit basin of approximately 0.8 Mm^3^ would be required during an EDF event, assuming a minimum pump rate of 500 m^3^/h to manage the 1:100-year event.

18.7.3.4 Ore and Surficial Soil Stockpiles

The high/mid-grade ore stockpile is located just south of the process plant. Runoff from the southern end of high/mid-grade ore stockpile will be collected by ditching, directed to a local collection pond, and transferred to the CWSP as needed. This local collection pond will also capture runoff from the western side of the haul road during operations. Excess water will be pumped to the CWSP if topography does not allow gravity drainage.

The low grade ore stockpile will require collection ponds at surrounding topographic low points to manage the surface water and seepage from the sub-watershed (Figure 18‑10). Contact water collected in these ponds may be partially consolidated before being pumped to the CWSP.

Water from the surficial soil stockpile will be directed to a contact water management pond or a collection ditch and pumped to the CWSP.

18.7.3.5 Plant Site Area and Plant Site Pond

The plant site area will be built up on a pad and graded towards ditches that drain by gravity to the plant site pond. The plant site pond will also capture runoff from the northern portion of the surficial soil stockpile. Runoff from the surficial soil stockpile will be directed by ditching and culverts towards the plant site pond.

The plant site pond will be either pumped to the CWSP or drain by gravity if grading allows.

18.7.3.6 Haul Roads

Surface water management infrastructure such as ditching, berms and pumps are required to convey contact water to water storage facilities for re-use, or for treatment and discharge to the environment in accordance with applicable regulatory requirements. All contact water from the Project mine site development area will be captured and managed by the water management system; this includes all haul roads but excludes the access road and treated effluent pipeline corridor. Ditching and berms will also be used to divert non-contact water from site facilities and haul roads.

18.7.4 Fresh Water Facilities – Operations

Fresh water will be required so that sufficient water is available for processing at all times of the year, and as needed for specialty uses where use of recycled water is not appropriate. These freshwater requirements are expected to include the following:

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·	Gland water for pumps.
·	Make-up water.
·	Elution circuit make-up water.
·	Fire water for use in the sprinkler and hydrant system.
·	Cooling water for mill motors and mill lubrication systems (closed loop).
·	Potable water.

A freshwater intake is proposed for Birch Lake, a very large waterbody located close to the primary fresh water use locations (process plant and accommodations complex). The intake will be located and designed to minimize environmental effects, including potential fish entrainment and impingement. Fresh water will be pumped from Birch Lake to water storage tank(s) until needed. Approximately 2.14 Mm^3^/a of fresh water will be required for the process plant and an additional 0.03 Mm^3^/a for the accommodations complex, on average, over the project life.

A potable water treatment system will be established to treat water intended for human consumption, although bulk bottled water may be used for drinking purposes, particularly during the construction phase.

18.7.5 Effluent Treatment Plant and Discharge

Effluent treatment will be in addition to the cyanide destruction and metal reduction that will occur within the process plant and the natural physical and chemical processes that will occur within the site ditching and ponds.

The ETP will be designed to produce an effluent quality appropriate for discharge to the environment in accordance with applicable regulatory requirements, including the MDMER, and the effluent concentrations required by the MECP to protect the receiving water and aquatic resources. Best available technologies that are economically achievable will be considered for the ETP to meet protection requirements.

The preliminary ETP considered in the PFS was a modular effluent treatment system. Additional engineering has optimized the water treatment concept as follows:

A biological process will be used based on the moving bed bioreactor concept, where plastic carriers with attached biofilm move freely in the water column and remove contaminants present in the wastewater. The moving bed bioreactor will also be used for cyanide destruction in addition to the in-plant destruction of cyanide in tailings using the sulphur dioxide/air(SO2/air) treatment process. The by-products are nitrate, carbon dioxide (CO2) and associated biomass.

The treatment process will continue to the removal of metals. Arsenic removal will be achieved by ferrous sulphate and iron co-precipitation principles. This will be followed by sulphide precipitation for further metals removal with the dosing of sodium sulphide. Adjustment of pH will be controlled by dosing acid and caustic to alkaline conditions of 7.5 to 8 as needed.

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·	The final stage involves flocculation, which includes a mixing tank before feeding to a clarification process. Following clarification, the fully treated effluent will be confirmed to meet all applicable regulatory discharge criteria before being released to the environment at the final discharge location in the southeast arm of Springpole Lake.

The southeast arm of Springpole Lake was selected as the preferred discharge location for treated effluent through the comprehensive alternatives assessment. This channelized section of Springpole Lake was selected as the primary effluent discharge location as it provides enhanced effluent mixing/attenuation, which will be supplemented with the use of a diffuser at the point of discharge. Although part of Springpole Lake, this portion of the lake has a defined current, much like a river. There is the potential that other minor discharge locations could be identified for the Project during construction, which could include aggregate operations if developed below the water table.

Effluent is proposed to be pumped to the discharge location through an HDPE pipeline, a distance of 9.3 km. The pipeline will be situated adjacent to the mine access road for a portion of the distance and then along a section of new access road to the discharge location for ease of construction and maintenance.

18.7.6 Water Balance

A site-wide water balance was completed to estimate the quantity of mine site contact water expected to be managed during the construction, operations and closure phases of the Project. The water balance model considers the precipitation and groundwater gains, and losses such as porewater loss, evaporation and infiltration.

Throughout all phases and scenarios, inflows to the Project are largely driven by site runoff generated by precipitation on the Project site. Water losses (apart from site discharge) are driven by porewater (void) loss in the CDF during operations, and evaporation outside of operations.

Discharge of treated water to the environment is expected to be necessary through all Project phases, with the exception of extreme dry climate conditions during the operations phase, during which only sewage treatment plant (STP) discharge is required. During the operations phase, the average treated water discharge (ETP and STP) is expected to be 6.82 Mm^3^/a. The highest treated water discharge simulated occurs during the construction and active closure phases under a 1:100 wet year climate condition. A treatment and discharge rate of 4.58 and 4.9 Mm^3^/a is required for construction and active closure phases, respectively. During construction, additional discharge related to the open pit basin dewatering is also required.

During the operations phase, water takings from Birch Lake will be required to support the process plant and accommodations complex. The greatest fresh water takings are estimated to occur during the operations phase, under the 1:100 dry year climate scenario, at a rate of 3.96 Mm^3^/a.

Table 18‑7 provides a summary of annual inflows and outflows from the integrated site water management system during the operations phase when mining is occurring, for average conditions. A summary for a 1:100 wet year (simulated in the final year of operations) and 1:100 dry year (simulated in the first year of operations) are also provided.

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Table 18‑6: Summary of Environmental Design Flood Storage and Pumping Requirements

Location	Storage Required (Mm^3^)	Pump Rate (m^3^/h)

| CDF Internal Pond | 1.4 | 100 |

| Open Pit watershed Area | 0.8 | 500 |

| CWSP | 0.7 | 950 |

| Total Project Site | 2.9 | 950 |

Table 18‑7: Average Annual Water Balance

	Water Volumes (Mm^3^/a)

| Water in Ore | 0.58 | 0.58 | 0.58 |

| Freshwater Takings for Process Plant from Birch Lake | 2.14 | 1.91 | 3.93 |

| Freshwater Takings for Potable Water from Birch Lake | 0.03 | 0.03 | 0.03 |

| Groundwater Inflows to Open Pit | 0.77 | 0.95 | 0.27 |

| Total Site Runoff | 4.47 | 6.76 | 2.36 |

| Total Inflows^1^ | 7.99 | 10.23 | 7.17 |

| Loss to Tailings Voids | 6.50 | 6.05 | 6.89 |

| Total Site Evaporation | 0.38 | 0.40 | 0.30 |

| Total Losses | 6.88 | 6.45 | 7.19 |

| Required Discharge (Total) | 1.11 | 3.59 | 0.03 |

Note:

1.	Total inflows are not equal to total losses and discharges due to accumulated or reduction in storage on site.

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Figure 18‑10: Watersheds and Flow Concept

Source: WSP, 2025

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Figure 18‑11: Water Management During Mining

Source: WSP, 2024

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18.8 Power and Electrical

Average electrical demand site-wide is estimated to be 60 MW. The PFS envisages that power supply to the proposed Springpole mine and process plant comes from a connection to the provincial distribution grid. This would be by a new 90 km south-east ward, 230 kV, single circuit power overhead transmission line which would provide tie-in to a new 230 kV line between Dinorwic and Pickle Lake which is currently being constructed by Wataynikaneyap Power. As alternatives to the 230 kV transmission lines, 115 kV and 138 kV have been considered, but the 230 kV option prevails in infrastructure cost and reduced power loss in the study conducted by Nordmin Engineering on behalf of First Mining.

The incoming electrical power from the 230 kV transmission line will be stepped down at the prefabricated substation to 25 kV for in-plant distribution through one 230/25 kV step-down transformer. All required auxiliary services, control room for substation operation, will be housed within the substation perimeter fence. The main substation control and automation system is designed for centralized operation of the substation, with a communication link to the plant-wide process control system (PCS).

The 25 kV distribution switchgear will be situated in the prefabricated main electrical room located next to the main substation. From the 25 kV switchgear, power will be supplied to all electrical rooms within the plant site through cable trays or via underground duct banks as needed. Overhead powerlines will feed distant facilities such as mine area, camp area, and water supply pumps.

There are total of six electrical rooms planned, including the main electrical room and regrind mill VFD room and those in various crushing and processing areas. All rooms will be prefabricated with internal electrical equipment pre-loaded prior to delivery to site.

Variable frequency drives have been allowed where required and will be fed from the main 25 kV switchgear location. All medium-voltage motors or drives will be fed from 4.16 kV switchgears, and starters for low-voltage motors will be grouped in motor control centres (MCC), with incoming breakers. The MCC’s will be in the electrical room and will include intelligent combination starters, with circuit breakers for instantaneous fault protection.

To support early works and construction activities, a temporary tie-in to a 115 kV line located approximately 30 km to the south will be utilized until the permanent transmission infrastructure is commissioned.

18.9 Fuel

The PFS includes a refueling facility which oversees fuel distribution operations for both heavy and light vehicles. It includes operational controls, safety systems, and security monitoring for the adjacent fuel infrastructure. It includes separate fueling stations for heavy equipment (haul trucks). Infrastructure includes a 350,000-L diesel fuel tank (one-week fuel demand), pump skid, control booth, and secondary containment.

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18.10 Water Supply and Management

Fresh water will be required so that sufficient water is available for processing at all times of the year, and as needed for specialty uses where use of recycled water is not appropriate. These fresh water requirements are expected to include the following:

·	Gland water for pumps
·	Make-up water
·	Elution circuit make-up water
·	Fire water for use in the sprinkler and hydrant system
·	Cooling water for mill motors and mill lubrication systems (closed loop)
·	Potable water

A fresh water intake is proposed for Birch Lake, a very large waterbody located close to the primary fresh water use locations (process plant and accommodations complex). The intake will be located and designed to minimize environmental effects, including potential fish entrainment and impingement. Fresh water will be pumped from Birch Lake to water storage tank(s) until needed. Approximately 2.14 Mm^3^/a of fresh water will be required for the process plant and an additional 0.03 Mm^3^/year for the accommodations complex, on average, over the Project life.

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19 MARKET STUDIES AND CONTRACTS

19.1 Market Studies

The gold markets are mature global markets with reputable smelters and refiners located throughout the world.

Gold is a principal metal traded at spot prices for immediate delivery. The market for gold trading typically spans 24 h/day within multiple locations around the world (such as New York, London, Zurich, Sydney, Tokyo, Hong Kong, and Dubai). Daily prices are quoted on the New York spot market and can be found on www.kitco.com.

19.2 Gold and Silver Price

First Mining has not completed any formal marketing studies with regard to gold production that will result from the mining and processing of gold ore from the Springpole Gold Mine into doré bars. Gold production is expected to be sold on the spot market. Terms and conditions included as part of the sales contracts are expected to be typical of similar contracts for the sale of doré throughout the world. There are many markets in the world where gold is bought and sold, and it is not difficult to obtain a market price at any particular time. The gold market is very liquid with a large number of buyers and sellers active at any given time.

The mineral resources were calculated at a gold price of US$2,450/oz and at a silver price of US$27.50/oz. As of November 2025, the median consensus price forecast from 25 investment dealers estimated a gold price of US$3,250/oz and a silver price of US$34.00/oz long term. As at November 17, 2025, the trailing two-year gold price was US$2,794/oz and silver price was US$32.08/oz and the trailing three-year gold price was US$2,504/oz and silver price was US$29.17/oz.

For the purpose of the 2025 PFS, a gold price of US$3,100/oz and silver price of US$35.5/oz were assumed. The exchange rate used in the study is C$1.00:US$0.741.

19.3 Contracts

Doré is shipped from site to major refineries. First Mining will enter into a refining agreement with various refiners around the world when the timing is appropriate. The terms and conditions will be consistent with standard industry practices. Refining charges include treatment and transportation.

19.4 Comments on Market Studies and Contracts

The qualified person has reviewed the relevant reports and analyses and is of the opinion that the marketing and commodity price information is suitable to be used in cashflow analysis to support this Pre-feasibility Study and its Technical Report. There are currently no firm contracts in place for the execution of the project (i.e. equipment, labour, power supply), however this is appropriate for the project in Pre-Feasibility Study phase.

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20 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

20.1 Environmental Considerations

20.1.1 Baseline and Supporting Studies

The Project site is in a remote area of northwestern Ontario, and there are no nearby active industrial/commercial developments. First Mining, and its predecessor Gold Canyon Resources, have been collecting environmental baseline data to support the Project’s EA and permitting since 2011, and data collection is ongoing. These studies are primarily focused on characterizing biological and physical components of the aquatic and terrestrial environments that may be impacted by and may interact with the proposed Project. The dataset compiled to date within these programs exceeds the level of environmental baseline data one would typically have in support of a PFS.

20.1.1.1 Meteorology and Climate

An onsite weather station (Springpole Station) was originally installed approximately 100 m northwest of the exploration camp in September 2011. It was programmed to record hourly measurements of temperature, precipitation, snow depth, humidity, air pressure, windspeed and direction. A new onsite weather station was installed near the existing exploration camp site in June 2020 to represent the meteorological and climate data for the Project. Several instrumentation updates were made to the station over the years, including the addition of an evaporation pan and a heated precipitation gauge. Regional climate stations operated by Environment and Climate Change Canada (ECCC) within 150 km of the site were also considered in summarizing the climatic conditions and included Ear Falls, Red Lake, Sioux Lookout and Pickle Lake stations. Climate data has been collected at these stations for several decades and is summarized as climate normals from 1981 to 2010 for Sioux Lookout A, Red Lake A and Pickle Lake A and as climate normals from 1971 to 2000 for Ear Falls (ECCC, 2021).

Monthly mean precipitation was determined using historical data and data from the onsite weather station and stations surrounding the Project site. The calculated mean annual precipitation for the Project site is 704 mm. In general, the ECCC stations show slightly more precipitation (10% to 22%) than the precipitation recorded at the Springpole station for the period from April 2022 to November 2022. However, all exhibit similar trends, with the greatest precipitation in the spring compared to fall. Climate normals for the Red Lake Airport Station (Red Lake A, Station 6016975, Canadian climate normals 1971 to 2000), which is the closest ECCC weather station to Project site, showed average daily air temperatures ranging between -23.9°C and -12.7°C in January and between 12.4°C and 23.8°C in July.

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20.1.1.2 Air Quality

The Project site is located in a remote area, absent of nearby large urban centres and industrial sources. Baseline air quality is influenced by long-range transport of air contaminants as well as by natural sources such as forest fires and volatile organic compound emissions from vegetation. Data sources used to determine baseline ambient air quality include the ECCC National Air Pollution Surveillance Program (NAPS) long-term air monitoring stations, literature values and onsite measurements. Regional stations from the ECCC NAPS network used for baseline air quality assessments include urban stations in Thunder Bay and Winnipeg (Ellen Station). These stations are considered conservative baselines for the Project site as they are influenced by urban sources and thus provide an overestimate air quality concentration. The onsite baseline air monitoring program was initiated in 2020 to measure suspended particulate matter, particle matter less than 10 µm and 2.5 µm in diameter, metals, nitrogen dioxide and sulphur dioxide at the Project site to establish existing conditions. The onsite measurements were used to refine background concentration estimates and compared to the regional concentrations used in the air quality assessment.

The results from NAPS Thunder Bay and Winnipeg datasets for the 90th percentile did not exceed the Ontario Ambient Air Quality Criteria (AAQC) and in most cases were well below the AAQC. The onsite data is comparable with the regional data.

20.1.1.3 Sound and Vibration

Sound and vibration baseline monitoring took place in April and June of 2021 at two locations; SP1, northwest of the Project; and at SP2, south of the Project. Sound monitoring was conducted in accordance with NPC-300 guidelines (MOECC 2013), and vibration monitoring was conducted in accordance with the NPC-119 guideline (MOE 1977b) for the assessment of potential damages to building structures due to vibration and NPC-207 (MOE 1977c) for the assessment of potential impacts related to human perception of vibrations. The sound monitoring data is characteristic of a rural (Class 3) area, in accordance with NPC-300. The background vibration at both locations are under 0.01 mm/s.

20.1.1.4 Geology and Physiography

The Project is underlain by glaciated terrain characteristic of a large part of the Canadian Shield. Land areas are generally of low relief with less than 30 m of local elevation, intersected by numerous lakes and watercourses. The local bedrock geology is primarily composed of three major groupings: trachyte porphyry intrusion that hosts the mineralization zones of the Springpole deposit, metavolcanic and siliciclastic rocks, which are host rocks that pre-date the mineralized intrusion and represent most of the bedrock in the vicinity of the proposed open pit, and metasediments located to the northeast of the proposed open pit.

Multiple geotechnical investigations have been conducted since 2020 to characterize the overburden geology and included geotechnical boreholes, test pits and hand auger holes as well as exploration boreholes (approximately 600 data points across the site) that contain information on overburden across the site. The overburden geology in the vicinity of the Project generally consists of surficial organics (i.e., peat) and glacial sediments. The overall thickness of overburden can be thin to absent in some locations of higher bedrock topography and tends to become thicker under lakes and ponds (i.e., in low-lying areas). On land, overburden thickness values are typically less than 5 m, while lake bottom sediment thickness is often greater than 5 m and, in some locations, greater than 40 m thick. Lake bed sediment thickness in the area of the proposed dike alignments is relatively thin and generally less than 2 m.

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20.1.1.5 Hydrogeology

Hydrogeological baseline characterization activities at site have been conducted to support design, and included multiple drilling campaigns with associated groundwater monitoring installations, hydrogeologic testing, groundwater monitoring, and corresponding desktop-based assessments.

The Project site is located in the Superior Province of the Canadian Shield physiographic region where the landscape is dominated by recent glacial activity. The undulating and subdued topography clearly shows the effects of glaciation, whereby cyclical ice advances and retreats eroded ancient mountain ranges and generally deposited reworked sediment in low-lying areas. Overburden is generally quite thin at the Project site, with frequent bedrock outcropping and an average overburden thickness of just 2.2 m on-land. The Project site area falls between two major watersheds, including Birch Lake and Springpole Lake.

A site specific hydrostratigraphic conceptualization has been developed through the synthesis of borehole data, test pits, outcrop mapping, and general observations from shoreline mapping. Hydrostratigraphy at site generally consists of overburden, a highly altered unconsolidated granular material (UGM) zone in the immediate vicinity of the proposed open pit, a zone of low rock quality designation rock (Low RQD zone) in the area of the proposed open pit, and the adjacent host rock surrounding the UGM and Low RQD zones. Lithologies of the host rock area generally include various metavolcanic and metasedimentary rocks and a trachyte porphyry intrusion (which is associated with the UGM zone).

Hydraulic conductivity has been assessed at the Project site based on packer testing, single well tests in monitoring wells, and pumping tests. Based on single well tests and packer tests, the geometric means of bedrock hydraulic conductivity range from 5.0 × 10-6 m/s in the area of the ore deposit (low RQD bedrock) to 6.4 × 10-8 in the area of the CDF (competent andesite rock).

Groundwater level monitoring infrastructure at site includes 86 monitoring wells and 10 multi-level vibrating wire piezometer (VWP) installations (total of 26 VWP units). Automated pressure transducers are installed in 27 of the monitoring wells and VWPs are equipped with dataloggers to facilitate temporal groundwater level monitoring at site. Long-term monitoring in monitoring wells and VWPs generally shows two archetypes of water level responses: 1) those strongly linked to Springpole Lake water level fluctuations and 2) those more indicative of responses to recharge/infiltration events (primarily the spring freshet).

Instantaneous water level monitoring was conducted in the fall of 2022. Measurements demonstrated that water levels generally coincide with local topography, with the highest water levels inland dissipating towards the lows and surface water features. This is consistent with the conceptual hydrogeological model; that groundwater recharge is driven by snowmelt/rainfall during the wet seasons and is focused at locations where the topography is favourable for infiltration. Groundwater flow is primarily focused through interconnected fractures in the bedrock (flows likely decreasing with depth), though will also be present at locations where thick overburden is present beneath the water table. Flows converge towards topography lows and natural surface water receivers, which act as the discharge point of groundwater.

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20.1.1.6 Hydrology

Baseline surface water hydrology programs have been conducted since 2011. Meteorological and hydrological data collection has generally taken place during three periods: 2011 to 2012, 2020 and 2021 to 2023. Several hydrometric monitoring campaigns have included flow monitoring stations and lake level monitoring stations.

In general, flow monitoring data from the smaller tributaries (less than 2 km^2^), show rapid responses to precipitation events, and produce zero flow during the late summer/fall periods. Slightly larger watersheds (2 to 10 km^2^) in the Project area have observed flow throughout the year. The much larger Springpole Lake watershed (1,372 km^2^), has consistently shown the highest annual flow in the spring. Flows measured at the outflow of Springpole Lake vary from 3.99 to 58.12 m^3^/s over the monitoring period. Rating curves and subsequent hydrographs were developed for Springpole Lake inflow and outflow stations. Based on an analysis of historical data from Water Survey of Canada (WSC) Sturgeon River at McDougall Mills station and the Springpole Lake outflow rating curve, the mean water level of Springpole Lake is approximately 391.28 masl. A low flow analysis was also conducted using multiple regional WSC stations, and the expected 7Q20 low flow at Springpole Lake outflow is 1.35 m^3^/s.

20.1.1.7 Surface Water

Baseline surface water quality monitoring completed from 2011 to 2023 included over 1,100 samples from 40 sampling locations.

Water quality results indicate that the surface waters in all monitored waterbodies are typical of oligotrophic lakes in northwestern Ontario, demonstrating limited nutrient availability, low turbidity and saturated to near-saturated dissolved oxygen concentrations. Levels of total suspended solids and total dissolved solids were generally very low. Water column profile results indicate that most lakes experience turnover in the spring and fall but remain stratified throughout the summer and winter months.

Concentrations of total and dissolved metals are very low, often at or below analytical detection limits, although there were a few occasions where measured baseline concentrations were higher than the applicable federal and provincial water quality guidelines for sampled waterbodies. These occurrences were irregular, generally associated with elevated total suspended solids levels and are considered representative of the natural heterogeneity of these lake systems. Parameters with concentrations outside of guidelines in the baseline condition include pH, total aluminum, phosphorus, total iron and total copper.

20.1.1.8 Aquatic Resources

Springpole Lake has a predominantly rocky shoreline and contains numerous islands and rocky shoals. The Birch River is its largest tributary, and it enters at the southwest end of Springpole Lake through a short section of rapids downstream from Cromarty Lake. There are also several small tributary streams flowing into Springpole Lake. The outflow of Springpole Lake is also through the Birch River, at the east end, into Gull Lake and then subsequently joins Cat Lake near Lake St. Joseph. Springpole Lake has a surface area of 2,861 ha; the Project footprint includes 6% of that area. Overall, Springpole Lake has a maximum depth of approximately 40 m and an average depth of 6.3 m. The portion of Springpole Lake that is within the Project footprint has an average depth of 13 m. The north basin of Springpole Lake, within which the Project is located, is 4.5 km wide and 6.5 km long. It is generally deeper and more open than the southeast arm of the lake and has three large basins exceeding 30 m in depth. The lake has three deep additional basins exceeding 20 m in depth. The southeast arm of Springpole Lake is 17.7 km long. Much of that length is a narrow channel bound by a steep bedrock wall along the north shore. Birch Lake has a surface area of 11,823 ha, with an irregular shape and a predominantly rocky shoreline. The maximum depth is 37 m, with an average depth of 7.4 m. The east end of the lake is deeper and more open than the west, which is characterized by narrow channels and comparatively shallow water.

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Beginning in 2011, an extensive and comprehensive assessment program was carried out to describe the fish community and habitat in surface water bodies that are within the Project area. Fish and fish habitat studies have been carried out on Springpole Lake, Birch Lake and numerous small unnamed waterbodies, and tributary streams in the Project area. The Springpole Lake fish community includes walleye, northern pike, yellow perch, rock bass, log perch, common white sucker, shorthead redhorse, common shiner, spottail shiner, lake whitefish, lake herring, lake trout, finescale dace, golden shiner and burbot. The fish community reported for Birch Lake consists of 19 different species, including lake trout, lake whitefish, northern pike, walleye, yellow perch, and other fish species. Among the small lakes surveyed; six host fish communities that include sport species including yellow perch and northern pike; four host only forage species; and two are considered devoid of fish. Of the small streams surveyed, the largest and best connected to lakes support yellow perch and other forage fish. The smaller, more ephemeral streams typically support only forage fish species.

Lake trout are utilizing multiple deep basins within Springpole lake as cold-water refuges during the summer months and move freely to both upstream and downstream lakes in the fall and return in the late spring. There are several lake trout spawning shoals in Springpole Lake. Aquatic assessments have demonstrated that the two deep basins adjacent to the open pit basin support equal or greater abundance of lake trout and lake whitefish, and the undisturbed (94%) portion of Springpole Lake will continue to support all resident species through the Project and into closure. No walleye has been observed spawning within the Project footprint in Springpole Lake, and no lake sturgeon or other endangered fish species are present in Springpole Lake.

20.1.1.9 Vegetation Communities and Wetlands

The Project is part of the Lac Seul Upland, which extends eastward from Lake Winnipeg in Manitoba to the Albany River in northwestern Ontario. Forest composition on the Project is typical of the Lac Seul Upland. Dominant tree species include trembling aspen, black spruce, white birch, balsam fir, and white spruce and jack pine. Understory ground cover species composition and abundance is typical of mesic mixed wood boreal sites and lacks microhabitats likely to harbor rare vascular plant species. A variety of common, early successional graminoids and herbaceous ground cover plants are prevalent on areas of the Project where mature timber has been removed or where the canopy is open and exposed to direct sunlight. Natural re-vegetation and succession has been observed to be rapid in areas of historical exploration.

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Vegetation assessments were conducted using the ecological land classification system. Baseline data was collected between 2011 to 2019 and resulted in 30 ecosites sampled for 90 vegetation plots. In 2021, 318 supplemental vegetation community surveys were completed to update existing information, and an additional 270 targeted vegetation surveys were conducted in 2022. Wetlands were assessed during the 2019 field season with 48 wetland ecosites surveyed. Further studies were completed in 2021 and throughout 2022.

As a result of the vegetation surveys, 397 species of plants were identified. Eleven species of plants are classified as rare in Ontario (Northern Marsh Violet, Floating Marsh Marigold, Alpine Woodsia, Nahanni Oak Fern, Small Yellow Pond-lily, Lakecress, Smooth-margin Nitrogen Moss, Red Dung Moss, Yellow Dung Moss and Cruet Dung Moss) and one species at risk was documented (Black Ash). The most commonly encountered species in 2012 were Speckled Alder, Labrador Tea, Bunch Berry, Sparse-flower Sedge, Stiff Clubmoss, Schreber’s Moss and Green Reindeer Lichen. In 2019, the most commonly encountered plant species were Schreber’s Moss, Peat Moss, Creeping Snowberry, Labrador Tea, Black Spruce, Knight’s Plume Moss, Grey Alder, Lingonberry and Mountain Fern Moss. Wetland ecosites comprised approximately 34% of the survey area and communities surveyed include swamps, marshes, bogs and fens.

20.1.1.10 Wildlife and Wildlife Habitat

Wildlife species and habitat types were evaluated by means of fieldwork and helicopter aerial surveys along selected routes, transects and survey/sampling points between 2011 and 2024.

Two Threatened bird species, the Eastern Whip-poor-will and Lesser Yellowlegs, have been observed in the vicinity of the Project. Additionally, six bird species of Special Concern, including the Bald Eagle, Barn Swallow, Canada Warbler, Common Nighthawk, Olive-sided Flycatcher and Rusty Blackbird, have also been observed in the vicinity of the Project; only the Bald Eagle was observed close to the Project site. Two species of bat considered Species at Risk were identified in the vicinity of the Project (Northern Myotis and Little Brown Myotis).

The Project is located within the northern portion of the Churchill range and adjacent to the Berens and Kinloch ranges for Boreal Caribou. Boreal Caribou are a valued species classified as Threatened under the provincial Endangered Species Act (ESA) and the federal Species at Risk Act (SARA). With the implementation of mitigation measures, such as the development and implementation of a habitat restoration program for Boreal Caribou, the potential effects on Boreal Caribou will be effectively managed.

Significant wildlife habitat was assessed for the baseline studies in accordance with provincial guidance (MNR 2000; MNRF 2010, 2014h) for the regional area; 17 habitats have been evaluated as candidate and 11 as confirmed. The following significant wildlife habitats were determined to be confirmed:

·	Waterfowl Stopover and Staging Areas (Aquatic)
·	Colonial Nesting Bird Breeding Habitat (Tree/Shrub)
·	Colonial Nesting Bird Breeding Habitat (Ground)
·	Bat Maternity Colonies
·	Bat Hibernaculum

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·	Regionally Rare Plant Species
·	Wild Rice Stand
·	Bald Eagle and Osprey Nesting Habitat
·	Aquatic Feeding Habitat
·	Marsh Bird Breeding Habitat
·	Habitat of Species of Conservation Concern

Not all of these categories were found in the Project development area and by following appropriate guidance documents and applying effective mitigation strategies, direct and indirect impacts will not result in a change in the form and function of significant wildlife habitat in the regional study area.

20.1.2 Environmental Monitoring

The follow-up and monitoring framework supports the overall environmental management for the Project. The follow-up monitoring will be implemented as part of the framework to verify predicted effects, evaluate the effectiveness of mitigation, and to measure compliance with permit conditions and statutory requirements. Monitoring is used to address uncertainties associated with effects predictions, identify any unanticipated effects, and provide input into corrective actions or adaptive management to limit those effects. Collectively, these actions improve the overall environmental performance of the Project.

The objectives of the follow-up and monitoring framework are to:

·	verify the accuracy of the effects assessment;
·	confirm the effectiveness of the measures implemented to mitigate adverse effects of the Project;
·	confirm compliance with commitments made during the EA process; and
·	confirm compliance with regulatory conditions of approval.

The follow-up and monitoring framework applies to the construction, operations, decommissioning and closure, and post-closure phases of the Project. In the event that monitoring results indicate that realized effects are appreciably different than predicted, further investigation will be undertaken and mitigation strategies may be modified as needed to reduce or eliminate unforeseen adverse effects.

Further monitoring details will be developed based on conditions of regulatory approvals issued by the federal and provincial regulatory agencies during permitting. The details of these programs will be developed in consultation with federal and provincial governments, and with Indigenous communities.

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20.1.3 Waste Management

Mine rock and processed tailings will be produced by the mine and co-disposed in the engineered CDF. The Project will produce approximately 146 Mt of PAG mine rock and 147 Mt of NAG non-metal leaching mine rock. NAG non-metal leaching mine rock will be used for CDF construction along with quarried construction rock. Approximately 81 Mm^3^ of construction material will be required to construct the CDF dams. Any surplus mine rock will require permanent storage.

Approximately 101 Mt of tailings will be produced by the process plant over the life of the mine. The Project will produce two tailings streams in order to best manage the potential for acid generation from the tailings in the long term:

·	Thickened non-acid generating (NAG) tailings (approximately 80.8 Mt): produced by passing a portion of the tailings through thickeners. Although thickened, the tailings are still able to be hydraulically conveyed through a high-density polyethylene (HDPE) tailings pipeline for final deposition in the north cell of the CDF.
·	Conventional slurry potentially acid generating (PAG) tailings (approximately 20.2 Mt): produced in a conventional slurry form at the process plant and transported to their final storage location in the dedicated south cell of the CDF by HDPE pipeline.

The CDF is proposed as a two-cell facility that provides the required storage capacity for mine rock and tailings. The CDF has been located on an area with low permeability andesite to provide a sound foundation with natural resistance to seepage. It has been designed to effectively use NAG non-metal leaching mine rock for construction purposes and to permanently store PAG mine rock, NAG thickened tailings and PAG conventional slurry tailings. The NAG tailings will be co-managed with the PAG mine rock in the north cell of the CDF, while the conventional PAG slurry tailings will be managed in the south cell of the CDF to mitigate acid rock drainage potential.

The CDF is designed to take advantage of the different properties of the mine wastes, (tailings and mine rock). In particular, the lower permeability of the tailings will be used to provide an oxygen barrier for the mine rock. The NAG tailings will function as an oxygen barrier thereby limiting the rate of oxidation and consequently metal leaching.

20.1.4 Water Management

Contact water arising from precipitation and groundwater is collected in ditches, sumps and ponds and transferred into the integrated site water management system for containment, treatment and discharge to the environment in accordance with applicable regulatory requirements, as needed. The water management strategy is to collect site runoff in local collection ponds. The largest ponds are the CDF internal pond, central water storage pond and ponds located within the open pit. The water collected in these ponds is considered as contact water and requires treatment through the effluent treatment plant prior to discharge to the environment. Designs and locations for perimeter ditching and ponds consider distances from nearby infrastructure and natural waterbodies and maintain setbacks from these features.

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The CDF internal pond collects water from both the north and south cells and from CDF perimeter seepage collection ponds. The primary drainage ponds and pumping systems will be installed at topographical low points surrounding the CDF footprint. Water collected in the CDF internal pond will be reclaimed to the plant/mill, reducing the need for freshwater demands from Birch Lake. Excess water will be pumped to the central water storage pond for monitoring, treatment, and discharge to the environment in accordance with applicable regulatory requirements to environment, as needed.

The central water storage pond is the ultimate collection point for contact water and will provide make-up water to the process plant as needed. Excess water will be pumped to the effluent treatment plant for treatment and subsequently discharged to the environment in accordance with applicable regulatory requirements.

The perimeter collection system will include ditches and pumping stations (drainage ponds) with pumps installed to facilitate collection of water and transfer to the central water storage pond.

20.2 Permitting Considerations

20.2.1 Environmental Permits

20.2.1.1 Federal Environmental Assessment

A Project description was submitted to the Impact Assessment Agency of Canada (IAAC) in February 2018, with environmental impact statement (EIS) Guidelines being issued by the Impact Assessment Agency of Canada in June 2018. Issuance of the EIS Guidelines confirmed that the Project is subject to environmental assessment under CEAA 2012, and associated regulations designating physical activities rather than under the new federal Impact Assessment Act (IAA).

An environmental impact statement/environmental assessment document (EIS/EA) was submitted to IAAC in November 2024, and the information presented in the EIS/EA is currently under review.

20.2.1.2 Federal Permitting Requirements

Federal approvals will be required to construct, operate and eventually close the Project. There are three primary federal departments that are anticipated to be involved with environmental approvals for the Project:

·	Fisheries and Oceans Canada (DFO) is responsible for administering the habitat protection provisions of the Fisheries Act (R.S.C., 1985, c. F-14) and the Species at Risk Act.
·	Environment and Climate Change Canada (ECCC) administers the Canadian Environmental Protection Act, 1999 (S.C. 1999, c. 33), the Migratory Birds Convention Act, 1994 (S.C. 1994, c.22), the Species at Risk Act (S.C. 2002, c. 29), and the pollution prevention provisions of the Fisheries Act, including amendments to Schedule 2 of the Metal and Diamond Mining Effluent Regulations.
·	Transport Canada (TC) is responsible for administering a wide range of legislation and regulations related to transportation and infrastructure, such as the Canada Transportation Act (S.C. 1996, c. 10), the Transportation of Dangerous Goods Act, 1992 (S.C. 1992, c.34), and the Canadian Navigable Waters Act (R.S.C., 1985, c. N-22).

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In addition to the requirement for assessment under CEAA, 2012, the following key federal permits may be required pending further regulatory advice:

·	Fisheries Act Authorization (DFO).
·	Canadian Navigable Waters Act (TC).
·	Schedule 2 of Metal and Diamond Mining Effluent Regulations (MDMER by ECCC).

Table 20‑1 summarizes the federal environmental approvals that are anticipated for the construction, operation and closure of the Project.

Table 20‑1: Anticipated Federal Environment-related Approvals

Approval/Instrument	Responsible<br> <br>Agency	Description

| Authorization Fisheries Act | DFO | Based on an assessment of impacts to fish and fish habitat, an authorization may be required for an undertaking or activity that may result in the death of fish or the harmful alteration, disruption or destruction of fish habitat. This may include: | | | | | · | establishment of the CDF and the central water storage pond; |

| | | · | in-water structures; and |

| | | · | mine dewatering groundwater effects and surface flow reductions that would cause impacts to watercourses supporting fisheries. |

| | | A financial guarantee is required to cover the cost of implementing the required offsetting plan. | |

| Schedule 2 Listing MDMER pursuant to the Fisheries Act | ECCC | Overprinting of water frequented by fish by mine waste (or other deleterious material) will require a listing under Schedule 2 of the MDMER. Potential areas of impact include the CDF and the central water storage pond. A financial guarantee is required to cover the cost of implementing the required compensation plan. | |

| Manufacturing, storage, and transportation of explosives Explosives Act | Natural Resources Canada | Explosives magazine, manufacturing facility and transportation require a federal permit, pursuant to Section 7 of the Explosives Act. If facility is owned by licensed explosives contractor, permit will be issued to them. | |

| Works in Navigable Waters Canadian Navigable Waters Act | TC | Alteration of navigable waters (e.g., dike in Springpole Lake) and crossing of navigable waters with infrastructure. | |

| Aeronautical Obstruction Clearance Canadian Aviation Regulations (SOR/96-433) of the Aeronautics Act | TC | Marking and lighting for structures that could interfere with aeronautical navigation, such as transmission line and airstrip (if developed). | |

| Transportation of Dangerous Good Act | TC | Focuses on the prevention of incidents when dangerous goods are imported, handled, offered for transport and transported. | |

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20.2.1.3 Provincial Environmental Assessment

First Mining has entered into a Voluntary Agreement with the Ministry of the Environment, Conservation and Parks (MECP) to undertake an Individual EA, under Section 3.0.1 of the provincial Environmental Assessment Act.

An EIS/EA was submitted to MECP concurrently with the EIS/EA submitted to IAAC in November 2024, and the information presented in the EIS/EA is currently under review.

20.2.1.4 Provincial Permitting Requirements

Provincial approvals will be required to construct, operate and eventually close the Project. Applications for these approvals will be submitted to the respective Ministries upon resolution of the EIS/EA process. There are three primary provincial ministries that are anticipated to be involved with environmental approvals for the Project:

·	Ministry of Energy and Mines (MEM) has responsibility to ensure the orderly development of mineral resources in Ontario, including primary responsibility for mine reclamation planning activities which includes a requirement to post financial assurance for implementation of closure measures by a third party.
·	Ministry of Natural Resources (MNR) is responsible for the protection and wise use of Crown resources including merchantable timber and aggregates.
·	MECP grants permits and approvals that address project aspects that are related to soil and water quality and quantity, air quality and noise, and SAR.

Ministry of Citizenship and Multiculturalism may also be involved with permitting Project components, although no permits are expected to be required. In addition, the Ministry of Transportation (MTO) may issue permit(s) if there is an infringement on provincial highway(s). The Ontario Energy Board (OEB) may also provide energy-related approval(s) for the Project, including approval to construct transmission lines.

Table 20‑2 provides a listing of the provincial approvals anticipated to be required or likely to be required for the Project.

Table 20‑2: Anticipated Provincial Environment-related Approvals

Approval	Responsible Ministry	Description of Activity/Facility Approved

| Environmental Compliance Approval – Air and Noise Environmental Protection Act | MECP | Release of air emissions and noise, such as from haul trucks (road dust), and facilities, such as the process plant. |

| Environmental Compliance Approval – Domestic Sewage Environmental Protection Act | MECP | Establishment and operation of a domestic sewage treatment plant. |

| Environmental Compliance Approval – Industrial Sewage Works Environmental Protection Act | MECP | Operation of the integrated water management system to provide for discharge to the environment, including but not limited to: central water storage pond, plant site pond and effluent treatment plant. |

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Approval	Responsible Ministry	Description of Activity/Facility Approved

| Overall Benefit Permit Endangered Species Act^(1)^ | MECP | Project-related effects to a protected species (in return for providing an overall benefit to the species in Ontario). |

| Permit to Take Water Ontario Water Resources Act | MECP | Taking of groundwater or surface water in excess of 50,000 litres per day, such as for a fresh water supply, or controlled dewatering the open pit and construction excavations. |

| Closure Plan Mining Act | MEM | Mine construction, operations and closure, including financial assurance for reclamation costs. The Closure Plan may also approve dams located outside watercourses and waterbodies. |

| Aggregate Resource License Aggregate Resource Act | MNR | Extraction of aggregate for construction purposes during the construction and operations phases. |

| Forestry Resource License Crown Forest Sustainability Act | MNR | Cutting of Crown merchantable timber, such as for site development and clearing for the access road and transmission line. |

| Land Use Permit/Sale of Crown Land/License of Occupation Public Lands Act | MNR | Tenure for long-term facilities on Crown land, such as for the transmission line or shoreline structures. |

| Scientific Collection Permit Fish and Wildlife Conservation Act | MNR | Conduct baseline fish/wildlife studies and relocate fish from watercourses and waterbodies. Also, for the destruction of beaver dams, if needed. |

| Work Permit Lakes and Rivers Improvement Act, Public Lands Act | MNR | Work/construction on Crown land. Construction of a dam or dike in any lake or river requires approval for the location of the dam, and its plans and specifications. |

| Clearance Letter Heritage Act | MCM | Confirms appropriate archeological and/or cultural heritage studies and mitigation, if required have been completed, for activity at the location to proceed. |

| Entrance Permit and Encroachment Permit Public Transportation and Highway Improvement Act | MTO | Allows access to construct a transmission line from a provincial highway and placement of a transmission line over a provincial highway, if required. |

| Leave to Construct Ontario Energy Board Act | OEB | Construction of a transmission line. |

Note:

1.	Endangered Species Act is being repealed and replaced with the Species Conservation Act; however, the new act is not yet in force.

20.2.1.5 Municipal Approvals

Since the Project is not located in an incorporated municipality, no municipal approvals are anticipated to be required.

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20.3 Social Considerations

20.3.1 Social and Community Setting

20.3.1.1 Socioeconomics

The region hosts a small number of remote tourism operations and seasonal camps. Remote designated tourism lakes identified in the Crown Land Use Policy Atlas Policy Report (MNRF, 2018) in the general vicinity of the Project include Birch Lake.

The three closest municipalities to the Project are Ear Falls, Red Lake and Sioux Lookout.

·	Ear Falls is located approximately 270 km north of Kenora and has a population of 924 (2021 Census). It consists mostly of rural residential areas. Economic growth is linked to resource development activity, dominated by the mining and manufacturing industries.
·	Red Lake is located north of the English River and has a population of 4,094 (2021 Census). Mining is the primary industry in Red Lake.
·	Sioux Lookout is located approximately 65 km north of the Trans-Canada Highway between Thunder Bay and Kenora and has a population of 5,839 (2021 Census). Key industries include health care and social assistance, retail trade, transportation and warehousing, and educational services.

20.3.1.2 Built Heritage Resources and Cultural Heritage Landscapes

Surveys of built heritage resources and cultural heritage landscapes were conducted in 2020 and 2021 on the Springpole property. One cultural heritage landscape has been determined to have potential and mitigation measures have been proposed.

20.3.1.3 Traditional Land Use Studies

Indigenous Traditional Land and Resource use (TLRU) studies describe Indigenous (First Nations and Métis) traditional land and resource uses in the area surrounding the Project site. First Mining has supported the development of TLRU studies throughout the EA process and will continue to support TLRU throughout all stages of the Project. TLRU information has been gathered and the following studies have been considered:

·	Cat Lake First Nation Indigenous Knowledge and Use Study: Kita-Ki-Nan Indigenous-led Assessment of the Springpole Project (CLFN, 2024a).
·	Cat Lake First Nation Socio-economic Baseline Study for the Proposed Springpole Gold Mine Project (CLFN, 2024b).
·	Lac Seul First Nation Indigenous Knowledge and Use Study: Kita-Ki-Nan Indigenous-led Assessment of the Springpole Project (LSFN, 2024a).
·	Lac Seul First Nation Socio-economic Baseline Study for the Proposed Springpole Gold Mine Project (LSFN, 2024b).

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·	Traditional Land Use and Occupancy and Traditional Ecological Knowledge Study Report for the Springpole Gold Project (MON, 2023).
·	Traditional Knowledge and Land Use Study for the First Mining Gold Corp. Springpole Mine Project. Completed by Know History Inc. Historical Services (MNO, 2021).
·	Springpole TKLUS Follow-up Report for NWOMC Completed by Know History Inc. (NWOMC, 2024).
·	Health, Socio-economic, Indigenous Knowledge and Land Use Baseline Study (SFN, 2024).
·	Wabauskang Traditional Knowledge and Use in the area of Springpole Gold Access Corridor Project (ArrowBlade, 2014).

TLRU practiced by Indigenous Peoples that could be affected by the Project include:

·	hunting and trapping locations are common around lakes and take advantage of trails and resource movement patterns within the surrounding forest;
·	preferred fishing locations were identified around waterbodies and watercourses where habitat, including Birch Lake and Springpole Lake, would support the various stages of aquatic resource, including spawning sites;
·	Indigenous communities place high value on a variety of plant species for their nutritional benefits, medicinal purposes and ceremonial uses; and
·	Indigenous communities noted that camps and cabins used for habitation, cultural practices and as spiritual sites are commonly found along lakeshores, including Birch Lake and Springpole Lake.

First Mining intends to continue incorporating the findings from TLRU studies completed by the local Indigenous communities in the Project area to refine the Project moving forward.

20.3.2 Engagement and Consultation

First Mining has aimed to establish positive and constructive relationships with local Indigenous communities, public stakeholders, and government agencies throughout the EA processes and will continue to provide opportunities over the life of the Project. The provincial and federal EA processes each have requirements for consultation under the Ontario Environmental Assessment Act and CEAA 2012. First Mining has undertaken coordinated consultation activities beyond what is required by the provincial and federal EA processes to maximize opportunities for engagement with Indigenous communities, public stakeholders and government agencies. First Mining has provided significant capacity funding throughout the process to support meaningful opportunities to understand the Project and share feedback towards First Mining’s ability to receive questions and understand concerns such that they can be addressed through information sharing, Project design optimizations, mitigation measures, monitoring and adaptive management.

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20.3.2.1 Indigenous Communities

The federal government identified Cat Lake First Nation (CLFN), Slate Falls Nation (SFN), Lac Seul First Nation (LSFN), Wabauskang First Nation (WFN), Mishkeegogoamang Ojibway Nation (MON), Ojibway Nation of Saugeen (ONS) and the Métis Nation of Ontario in 2018 (updated in 2020), while in 2018 the provincial government identified CLFN, SFN, LSFN, WFN, MON, ONS, Pikangikum First Nation, and Métis Nation of Ontario, as potentially impacted by the Project or having an interest in the Project. The Métis Nation of Ontario, Region 1, as represented by the Northwest Ontario Métis Community (NWOMC), office is located in the City of Dryden, 185 km from the Project site.

The communities located in closest proximity to the Project include CLFN, SFN, LSFN and MON. In 2017, CLFN, LSFN, and SFN entered into a Shared Territory Protocol agreement. Under this agreement, CLFN, LSFN and SFN are referred to as the “Shared Territory Protocol Nations” (STPN) and have agreed to certain geographic areas referred to as the “Shared Territory”; to share ownership in development and management of natural lands and resources within their territory, and to maximize benefits for each community. As of December 2023, CLFN and LSFN have continued to work together on many aspects of Project consultation, while SFN has moved forward on its own and is currently working independently on Project consultation. First Mining and MON signed a Long-Term Relationship Agreement in July 2025. Cat Lake First Nation and Lac Seul First Nation signed a Process Agreement with First Mining in October 2024, to support an Anishinaabe-Led Impact Assessment for the Project.

Consultation with the communities has primarily focused on the key EA milestones, including sharing baseline studies, the assessment of alternatives, the comprehensive EIS/EA, traditional knowledge and traditional land use studies/information as well as other topics of importance to communities (environmental protection, community health and wellness, economic and employment opportunities, and education and training). Enhanced opportunities to participate in the planning process were provided based on the level of interest communities expressed in the process. To date, CLFN, LSFN, MON, NWOMC and SFN have expressed a higher level of interest compared to other communities. ONS has expressed an interest in staying informed about the Project and participation in training, employment and other potential opportunities.

Information has been distributed by First Mining to Indigenous communities throughout the EA process to keep them informed of ongoing Project-related activities. First Mining has engaged with Indigenous communities throughout the technical review process of the EIS/EA through a variety of flexible methods, such as distributing Ojibway and English summary documents and videos, presenting findings through virtual webinars, hosting Community Information Sessions and meetings/workshops, where accepted, using social media, and continuing to relay information to community contacts through email, meetings and phone calls.

20.3.2.2 Public Stakeholders

First Mining has engaged with Public Stakeholders to understand their knowledge of the Project area and to identify how the Project may affect their interests. During preparation of the draft EIS/EA, First Mining shared Project information related to the proposed Project components, EIS/EA and ongoing updates with members of the Public to receive feedback on their interests. First Mining and the public have engaged through email, telephone, and meetings to discuss the Project and topics such as land use, access, and potential employment and business opportunities.

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20.3.2.3 Government Agencies

First Mining has engaged government agencies throughout the EA process and focused on opportunities for a broad list of government agency reviewers at the EA milestones. Throughout the EA process, First Mining met with representatives from federal government agencies and provincial government ministries to share information about the Project and receive feedback at each phase.

The following Federal agencies have been engaged:

·	Impact Assessment Agency of Canada
·	Environment and Climate Change Canada
·	Fisheries and Oceans Canada
·	Indigenous Services Canada
·	Health Canada
·	Natural Resources Canada
·	Transport Canada.

The following Provincial agencies have been engaged:

·	Ministry of the Environment, Conservation and Parks;
·	Ministry of Natural Resources;
·	Ministry of Energy and Mines;
·	Ministry of Citizenship and Multiculturalism; and,
·	Ministry of Indigenous Affairs and First Nations Economic Reconciliation.

Federal and provincial agencies which are outside the list of core agencies may become more involved in the Project as it proceeds through other phases.

20.4 Closure and Reclamation Planning

A general rehabilitation and closure approach to meet the objectives of Ontario Mining Act and Regulation 35/24 has been developed and described below. The overall objective for closure is to return the Project site to a productive condition after mining is complete that is capable of supporting plant, wildlife and fish communities, and other applicable land uses.

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20.4.1 Closure and Reclamation Plans

Closure concepts will be refined as engineering advances during permitting resulting in a more robust set of site closure objectives and a detailed closure plan, as required by the Mining Act. The primary goal of decommissioning and closure of the mine site is to establish a site that is physically, chemically and biologically stable. The Project footprint will be rehabilitated to a productive and natural state as practicable. The reclamation and decommissioning/closure objectives for the Project include:

·	re-establishing natural drainage;
·	rehabilitating disturbed lands;
·	confirming site runoff meets regulatory criteria;
·	establishing a self-sustaining vegetative cover; and
·	creating of functional wetland habitat.

Progressive rehabilitation activities that can be performed prior to final closure and that do not pose a barrier to daily operations will be considered for progressive reclamation. Progressive rehabilitation of affected areas will be fully considered during operations. Some potential opportunities include:

·	regular backhaul of waste material off site
·	decommissioning and salvage of infrastructure used only for exploration and construction
·	initiation of revegetation studies during operations to evaluate soil amendments and seed mixes to maximize the success of the final revegetation program
·	recontouring and revegetation of disturbed areas from exploration and construction phases that are not needed during operations;
·	progressive reclamation of the CDF perimeter dams; and
·	advancement of the designated fish habitat development area to final contours in preparation for completion during closure.

Closure decommissioning and rehabilitation takes place once mining has permanently stopped. Activities will include decommissioning of infrastructure and buildings, site preparation and planting of disturbed sites, and surface preparation to facilitate natural revegetation. Closure is expected to be completed five years after operation ceases. The general closure activities that will be completed for the site will include:

·	Reclamation of the open pit, including regrading overburden slopes to a stable side slope, and installation of enhancements to support fish habitat and increase biodiversity.
·	Controlled refilling of the open pit basin.

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·	Removal of hazardous substances.
·	Cleaning and removal of equipment.
·	Demolishing and salvage of site buildings and infrastructure.
·	Infilling or breaching of water storage ponds.
·	Water treatment.
·	Ground cover application, re-sloping, and revegetation of disturbed sites.

The operational design and the decommissioning and closure concept for the CDF have been developed to promote long-term chemical and physical stability, minimize erosion, provide long-term environmental protection and minimize long-term maintenance requirements. The CDF closure concept involves the following measures:

·	Construct an overflow spillway at the south cell perimeter dam to safely pass the inflow design flood to the environment.
·	Deposit NAG tailings over the entire north cell surface to fully cover the PAG mine rock and limit oxygen ingress;
·	Vegetate the tailings or, if necessary, place and grade an erosion protection cover over the entire north cell surface and direct all runoff to the south cell.
·	Deposit NAG tailings or other suitable soil cover in the south cell of the CDF to remove excess pond capacity and provide cover over PAG tailings.
·	Breach perimeter collection ponds and allow runoff and seepage water to report to environment once water quality requirements are met.

Post closure activities typically include site security, water treatment, and environmental/reclamation monitoring to ensure progression towards closure objectives. Post closure monitoring and assessments will continue until the site has met the rehabilitation objectives at which time the property would be released to the Crown.

20.4.2 Closure Cost Estimates

Closure cost estimates are detailed in Section 21.

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21 CAPITAL AND OPERATING COSTS

21.1 Introduction

The facilities at Springpole will consist of an open pit mine, primary crushing, SAG mill and ball mill grinding, flotation, flotation tailings leaching and CIP, flotation concentrate regrind/leaching/CCD, carbon elution, Merrill Crowe, and refinery, tailings thickening/cyanide destruction, reagent mixing, and associated infrastructure.

21.2 Capital Costs

21.2.1 Overview

The estimate is derived from data as derived from drawings, models, material take offs and lists generated for this study. The capital costs include all associated infrastructure as defined within the scope of work.

The capital cost estimate has been summarized in Table 21‑1 by work breakdown structure (WBS) and Table 21‑2 by major discipline stated in United States Dollars (US$) with a base date of Q4 2025 and with no provision for escalation.

Table 21‑1: Estimate Summary Level 1 Major Facility

WBS Level 1	Description	Initial CAPEX<br> <br>(US$M)	Sustaining CAPEX<br> <br>(US$M)	Closure Costs<br> <br>(US$M)	Total CAPEX<br> <br>(US$M)

| 1000 | Mining | 302.5 | 304.1 | 40.5 | 647.1 |

| 2000 | Site Development | 39.6 | - | - | 39.6 |

| 3000 | Process Plant | 348.6 | - | - | 348.6 |

| 4000 | On-Site Infrastructure | 75.5 | - | - | 75.5 |

| 5000 | Off-Site Infrastructure | 47.0 | - | - | 47.0 |

| Total Direct Costs | | 813.2 | 304.1 | 40.5 | 1157.9 |

| 6000 | Indirects | 68.6 | - | - | 68.6 |

| 7000 | EPCM Services | 70.2 | 18.5 | - | 88.8 |

| 8000 | Owners Costs | 24.4 | - | - | 24.4 |

| Total Indirect Costs | | 163.3 | 18.5 | - | 181.8 |

| 9000 | Provisions (Contingency) | 127.5 | - | - | 127.5 |

| Project Total | | 1,104.1 | 322.6 | 40.5 | 1,467.2 |

Note:

1.	Closure costs noted are associated with the WMF closure costs. Pit closure costs are in Sustaining Capital costs

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Table 21‑2: Initial Estimate by Major Discipline

Disc.	Description	Initial CAPEX (US$M)

| A | Architectural | 42.9 |

| B | Earthworks | 101.5 |

| C | Concrete | 39.3 |

| S | Structural Steel | 31.7 |

| F | Platework | 30.6 |

| M | Mechanical Equipment | 167.1 |

| P | Piping | 44.5 |

| E | Electrical Equipment | 43.4 |

| L | Electrical Bulks | 20.9 |

| I | Instrumentation | 14.8 |

| N | Mobile Equipment | 1.5 |

| R | Third Partys | 275.1 |

| Total Direct Costs | | 813.3 |

| O | Owners Costs | 24.4 |

| T | Project Delivery | 70.3 |

| U | Field Indirects | 52.9 |

| V | Spares/First Fills/Vendor Reps | 15.7 |

| Total Indirect Costs | | 163.3 |

| Y | Provisions/Contingency | 127.5 |

| Project Total | | 1,104.1 |

Table 21‑3: CAPEX Contributor Definition

WBS	Description	Capex Contributor

| 1000 | Mining | |

| 1100 | Mining Development Surface | AGP |

| 1200 | Mine/Lake Dewatering | AGP |

| 1300 | Haul Roads | AGP |

| 1400 | Waste Rock Storage | AGP & WSP |

| 1500 | Ore Stockpile | Ausenco |

| 1600 | Mining Equipment | AGP |

| 1700 | Mine Ancillary Services | Ausenco |

| 1800 | Mine Explosives Magazine | AGP |

| 2000 | Site Development | |

| 2100 | Dike No.1 | WSP |

| 2200 | Dike No.2 | WSP |

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WBS	Description	Capex Contributor

| 2400 | Bulk Earthworks | Ausenco |

| 3000 | Process Plant | |

| 3100 | Crushing, Stockpile/Reclaim | Ausenco |

| 3200 | Grinding | Ausenco |

| 3300 | Flotation | Ausenco |

| 3400 | Flot Tailings Leaching | Ausenco |

| 3500 | Concentrate Leaching | Ausenco |

| 3600 | ADR Plant | Ausenco |

| 3700 | Refinery | Ausenco |

| 3800 | Reagents and Services | Ausenco |

| 3900 | Tailings Filtration | Ausenco |

| 4000 | On-Site Infrastructure | |

| 4100 | Power Supply and Distribution | Ausenco |

| 4200 | Ancillary Buildings | Ausenco |

| 4300 | Site Services | Ausenco |

| 4400 | Mobile Equipment | Ausenco |

| 5000 | Off Site Infrastructure | |

| 5100 | Off-site Roads | Existing – no work required |

| 5200 | HV Power Line (OHPL) | First Mining |

| 5300 | Water Supply | Ausenco |

| 6000 | Project Indirects (Field, Camp, Vendors, Spares) | Ausenco and AGP |

| 7000 | EPCM | Ausenco and AGP |

| 8000 | Owner’s costs | First Mining |

| 9000 | Provisions (Contingency) | Ausenco and AGP |

21.2.2 Basis of Estimate

The capital cost estimate is an update of the previous phases of work developed in CAD and presented in USD in Q4 2025 with a new mine plan and using recent equipment pricing. The original quantities were based on layouts from data from projects and research in Ausenco's internal database and knowledge gained from similar operations. These were scaled with updated layouts where mechanical equipment and tanks were revised to meet the new mine plan data.

The capital cost estimate conforms to Class 4 guidelines for a PFS level estimate with a -20%/+25% accuracy, according to the Association for the Advancement of Cost Engineering International (AACE International).

The CAPEX is a quantitative based cost estimate, with engineering developed material take-offs with factored quantities, semi-detailed unit costs and budgetary quotations for major equipment.

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The structure of the estimate is a build-up of the direct & indirect cost of the current quantities; this includes the installation/construction hours, unit labour rates and contractor distributable costs, bulk and miscellaneous material and equipment costs, any subcontractor costs, freight and growth.

The methodology applied to develop the estimate is as follows:

·	Updated a priced Mechanical Equipment List and Electrical Equipment List with new process equipment engineering.
·	Updated bulk material pricing based on recent, suitable reference projects.
·	Updated the installation cost for equipment and bulks.
·	Updated requirements for freight.
·	Determined and agreed on foreign exchange rates.
·	Updated growth allowances for each estimate line item.
·	Updated Indirect costs.
·	Determined the estimate contingency value.
·	Currency exchange rate 0.74 US$ - 1 C$.

21.2.3 Mine Capital Costs

The mining capital cost estimate is grouped into various categories of the WBS. These include:

·	1100 - Mine Development - Surface
·	1200 - Mine – Bay Dewatering
·	1400 - Waste Rock Storage
·	1500 – Ore Stockpile
·	1600 - Mining Equipment
·	1800 – Mine Explosive Magazines
·	1900 – Mining General
·	2300 – Dike Management

Mining initial and sustaining capital estimates were provided by AGP and are all-inclusive costs for their scope of work, as per the following Table 21‑4. This does not include CAPEX costs for earthworks, mechanical and electrical equipment, off-plot pipe runs and buildings estimated by Ausenco with earthworks quantities for the waste rock facility by WSP.

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Table 21‑4: Mining Capital Costs Estimate Summary

WBS	Level 1 Description	Initial CAPEX<br> <br>(US$M)	Sustaining CAPEX<br> <br>(US$M)	Closure Costs<br> <br>(US$M)	Total CAPEX<br> <br>(US$M)

| 1100 | Mine Development Surface | 77.8 | 191.5 | 34.7 | 304.0 |

| 1200 | Mine Dewatering surface (excl. lake dewatering) | 8.0 | 2.1 | - | 10.1 |

| 1300 | Haul Roads | 4.1 | - | - | 4.1 |

| 1400 | Waste Rock Storage (clearing and basin shaping) | 88.6 | - | 5.5 | 94.1 |

| 1500 | Ore Stockpile | 5.2 | - | - | 5.2 |

| 1600 | Mining Equipment | 40.1 | 23.3 | - | 63.4 |

| 1800 | Mine Explosives Magazine | 0.05 | - | - | 0.05 |

| 1900 | Mining General | 0.6 | 0.6 | - | 0.6 |

| 2300 | Dikes (road access, clearing, general fill using mine waste) | 3.0 | - | - | 3.0 |

| Total Direct Costs | | 227.6 | 217.5 | 40.2 | 484.6 |

| 7000 | EPCM services | Incl with Direct costs | - | - | Incl with direct costs |

| 9000 | Provisions (Contingency) | 11.4 | - | - | 11.4 |

| Project Total | | 238.3 | 217.5 | 40.2 | 496.0 |

21.2.3.1 Mine Development, Surface (WBS 1100)

Mining activity commences in advance of the process plant achieving commercial production. This includes the movement of 17.7 Mt of waste and placement of 2.3 Mt of mill feed material in a stockpile adjacent to the primary crusher. The cost covers all associated management, drilling, blasting, loading, hauling, support, engineering and geology departments labour, grade control costs and financing costs.

Pit electrification is included in this portion of the cost with some in Year -1 and extension of the electrical system in Year 1 as part of sustaining capital.

21.2.3.2 Dewatering (WBS 1200)

The initial capital dewatering costs are primarily the controlled dewatering of the open pit basin behind the dikes. The total initial capital is for the controlled dewatering of the open pit basin. Sustaining capital is for the mine in-pit dewatering.

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21.2.3.3 Haul Roads (WBS 1300)

This category includes the construction of an estimated 13.2 km of haul roads and access roads.

21.2.3.4 Waste Rock Storage (WBS 1400)

The mining during the pre-production period includes the development of a quarry, CDF Quarry, for use in haul road construction and overall site requirements. A smaller sized equipment fleet (91-t class trucks) will be used in road, dike, and waste storage facility preparation. A small fleet of the larger mining equipment would also be planned for this period to excavate the quantity of material required for the CDF Embankments.

The larger mining fleet will become operational in Year -1, when pre-production mining commences from the open pit area as the water level drops from the controlled dewatering of the open pit basin. This work will widen/improve haul roads that had been pioneered, advancement of the waste storage facility and continued stockpiling of mill feed material in preparation of the process plant commissioning.

The development of the waste rock storage facility by the smaller equipment fleet is included as well as the cost associated with clearing and grubbing. This category also includes estimate for the drainage and seepage collection around the perimeter of the waste rock storage facility area. This category also includes estimate for the drainage and seepage collection around the perimeter of the waste rock storage facility area for approximately 9,000 m of ditching and 12 collection basins.

21.2.3.4.1 Waste Rock Storage Facility/Dry Tailings Storage Facility (WBS 1400)

The Waste Rock Storage Facility design and quantities have been completed by WSP. Earthworks and piping costing was completed by Ausenco based on recently completed Ausenco Feasibility Studies in Ontario where pricing was sourced from the market. AGP provided costing for supply and haulage of rock from the site quarry and mine with larger rock (600 minus and above) directly placed and smaller aggregate requiring sizing below 600mm minus delivered to a crushing facility.

21.2.3.5 Ore Stockpile (WBS 1500)

Preparation of the ore stockpile locations which includes the foundation and drainage is covered in this cost category. This includes the high, mid, and low-grade stockpiles adjacent to the plant and ROM hopper.

21.2.3.6 Mining Equipment (WBS 1600)

The mining equipment capital costs reflect the use of financing of the major equipment and most support equipment. Equipment prices for major equipment is based on current quotations from local vendors. A 20% down payment is included in the capital cost for those units financed. The remaining cost is included in operating costs discussed in Section 21.3. An allowance for the cost for the base fleet management system to properly track mill feed, and waste types is included in this category.

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The base costs provided by the vendors are included in a calculation for each unit cost calculation and options added to that including the equipment fleet management modules. In the case of the electric/hydraulic shovels it also includes switchgear and initial trailing cable requirements. The capital cost if it was to be purchased outright is shown for comparison. The cost of financing and down payment of some of the major equipment is shown in Table 21‑5 as used in the PFS.

Table 21‑5: Major Mine Equipment – Capital Cost, Full Finance Cost and Down Payment

Equipment	Unit	Capacity	Full Finance Cost	Down Payment	Capital Cost

| | | | US$M | US$ | US$ |

| Production Drill | mm | 140 | 1.7 | 0.3 | 1.4 |

| Production Drill (Electric) | mm | 251 | 5.4 | 0.9 | 4.5 |

| Production Loader | m^3^ | 23 | 9.4 | 1.6 | 7.9 |

| Hydraulic Shovel (Electric) | m^3^ | 36 | 17.7 | 3.0 | 14.8 |

| Haulage Truck | t | 240 | 6.6 | 1.1 | 5.5 |

| Haulage Truck | t | 91 | 2.1 | 0.4 | 2.0 |

| Crusher Loader | m^3^ | 13 | 2.8 | 0.5 | 2.3 |

| Production/Support Excavator | m^3^ | 7 | 1.4 | 0.3 | 1.4 |

| Track Dozer | kW | 474 | 2.2 | 0.4 | 1.8 |

| Grader | kW | 163 | 2.0 | 0.3 | 1.8 |

The cost of spare truck boxes, loader buckets, and dozer blades are included in the capital cost for the major equipment cost estimate.

The distribution of capital costs is completed using the number of units required within a period. If new or replacement units are needed, that number of units and using the unit cost applied against them (20% of that for major equipment) is used to determine the capital cost in that period. There is no allowance for escalation in any of these costs.

The balancing of equipment units based on operating hours is completed for each major piece of mine equipment. The smaller equipment was based on number of units required, based on operational experience. This includes such things as pickup trucks (dependent on the field crews), lighting plants, mechanics trucks, etc. Additional support equipment for snow removal and site water control was included to accommodate the expected climatic conditions.

The most significant piece of major mine equipment is the haulage trucks. At the peak of mining, twenty-five 240 t units are necessary to maintain mine production. This happens in Year 6. The maximum hours per truck/per year are set at 6,000. There are periods where the maximum hours per unit are below what the maximum possible can be. In those situations, increasing the maximum on the number of trucks still leaves residual hours required to complete the material movement, therefore, the number of total trucks is unchanged. In these cases, the hours required are distributed evenly across the number of trucks on site and available. The other major mine equipment is determined in the same manner.

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With a mine life of eight years the major equipment does not require a replacement cycle. Support equipment is replaced within the mine life. The support equipment is usually replaced on several year’s basis. For example, pickup trucks are replaced every three years, with the older units possibly being passed down to other departments on the mine site, but for capital cost estimating new units are considered for mine operations, engineering, and geology.

The number of pieces of major equipment required by year are shown in Table 21‑6.

Table 21‑6: Mine Equipment on Site

Equipment	YR -2	YR -1	YR 1	YR 2	YR 3	YR 4	YR 5	YR 6	YR 7	YR 8	YR 9	YR 10

| Production Drill (140mm) | 0 | 1 | 1 | 1 | 1 | 2 | 2 | 2 | 2 | 2 | 0 | 0 |

| Production Drill (251mm) | 0 | 1 | 2 | 3 | 3 | 3 | 4 | 3 | 3 | 2 | 0 | 0 |

| Production Loader | 1 | 1 | 2 | 2 | 2 | 2 | 2 | 3 | 1 | 1 | 1 | 1 |

| Hydraulic Shovel (Electric) | 0 | 1 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 1 | 0 | 0 |

| Haulage Truck (91 t) | 5 | 5 | 0 | 0 | 0 | 0 | 0 | 0 | 0 | 0 | 0 | 0 |

| Haulage Truck (240 t) | 4 | 5 | 15 | 16 | 18 | 20 | 22 | 25 | 22 | 12 | 2 | 1 |

| Crusher Loader | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 0 | 0 |

| Production/Support Excavator | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 0 | 0 |

| Track Dozer | 0 | 4 | 4 | 4 | 4 | 4 | 4 | 4 | 4 | 4 | 1 | 1 |

| Grader | 0 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 1 | 1 |

The smaller production drills, production excavator and haulage trucks (91 t) are used in the construction of the dikes, roads, and preparation of the CDF. After the main mine equipment arrives on site this equipment falls to a support role as required. The smaller drills are used for pre-shear drilling to keep the larger electric drills working on the more productive patterns. The production excavator will be used to clean the hanging wall and footwall of the ore zones to minimize dilution.

The expected equipment life is:

·	production drill (140 mm)	= 25,000 h
·	production drill (251 mm)	= 35,000 h
·	hydraulic shovel (36 m3)	= 80,000 h
·	production loader (23 m3)	= 36,000 h
·	haulage truck (240 t)	= 72,000 h
·	haulage truck (91 t)	= 50,000 h
·	track dozer	= 25,000 h

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·	grader	= 25,000 h
·	support excavator	= 35,000 h

Other support equipment is normally determined in number of years and varies by its duty in the mine. Light plants for example are replaced every four years. The integrated tool carrier for site support is purchased once at the Project start and is not replaced over the mine life.

21.2.3.7 Mine Explosive Magazine (WBS 1800)

This category accounts for pad preparation for the mine explosive magazines. The cost of the magazines is provided as part of the explosives cost (magazine monthly rental from vendor).

21.2.3.8 Mine General (WBS 1900)

This category accounts for capital cost items related to the mine engineering area.

21.2.3.9 Dike Roads (WBS 2320)

This category accounts for 4.1 km of haul road construction.

21.2.3.10 Overland Pipelines

The estimate allows for the supply and installation of all major overland (off-plot) piping, which includes pipework, fittings, valves special pipe items and supports, for the following services:

·	Fresh water – Birch Lake to Process Plant
·	Potable Water – Process Plant to camp
·	Reclaim Water – Collection Pond to Process Plant
·	Sewage – from camp to sewage treatment plant
·	Sewage – from Process Plant to sewage treatment plant
·	Water Discharge – sewage treatment plant to water treatment plant
·	Excess Contact Water – contact water pond to water treatment plant
·	Treated Water - discharge to Springpole Lake
·	Mine Pit Dewatering – pit to contact water pond
·	Tailings Seepage – tailings seepage pond to contact water pond
·	Runoff Water from ROM pile – ROM pad to collection water pond

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The overland pipelines were individually preliminary designed and compiled in an Overland pipeline List. Flows, specifications, sizing and lengths of each line have been developed by Engineering as the basis for the cost estimate as part of the 2021 PFS.

The supply pricing for off plot piping was escalated to current Q4 2025 pricing to be in line with recent projects and supply pricing includes supply and delivery of pipe, fittings, valves, and supply of heat tracing and insulation.

Installation hours and costs are based on in-house typical installation hours for each size and specification.

21.2.4 Process Plant Capital Costs

The process plant is designed for conventional process operations. The plant as estimated will treat 30,000 t/d or 1,250 t/h. The process equipment requirements are based on the process flowsheet and process design criteria as defined in Section 17 of this report. Costs have been built up using the equipment lists for the Project, together with the layout and MTOs. Each discipline is detailed below. Total costs by area are shown in Table 21‑7.

Table 21‑7: Process Plant Initial Direct Costs

WBS	Level 1 Description	TOTAL CAPEX (US$M)

| 3100 | Crushing. Stockpile/Reclaim | 33.0 |

| 3200 | Grinding | 94.3 |

| 3300 | Flotation | 28.9 |

| 3400 | Flotation Tailings Leaching, CIP and Detox | 59.2 |

| 3500 | Concentrate Leaching, CCD and Detox | 62.7 |

| 3600 | ADR Plant | 6.8 |

| 3700 | Merrill Crowe, Refinery | 25.4 |

| 3800 | Reagents & Services | 24.4 |

| 3900 | NAG Slurry Tailings | 14.0 |

| Process Building Total | | 348.6 |

21.2.4.1 Crushing, Stockpile, Reclaim (3100)

The materials handling and crushing circuit cost includes the following key equipment:

·	primary gyratory crusher
·	mill feed apron feeders
·	materials handling equipment

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21.2.4.2 Grinding (3200)

The grinding circuit cost includes the following key equipment:

·	SAG mill
·	Ball mill
·	cyclone feed pump box
·	classification cyclone cluster

21.2.4.3 Flotation (3300)

The flotation circuit cost includes the following key equipment:

·	rougher flotation cells
·	rougher flotation concentrate pump box
·	flotation tailings pump box
·	trash screens

21.2.4.4 Flotation Tailings Leaching, CIP and Detox (3400)

The flotation tailings leach, carbon adsorption and detoxification circuit cost includes the following key equipment:

·	tailings thickener
·	leach/CIP tanks and agitators
·	inter-tank carbon screens
·	carbon sizing screen
·	detoxofication tanks and agitators
·	carbon safety screen
·	detoxification product pump box

21.2.4.5 Concentrate Leaching, CCD and Detox (3500)

The flotation concentrate leach circuit cost includes the following key equipment:

·	classification cyclone cluster
·	regrind mills

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·	regrind feed and discharge pump boxes
·	concentrate thickener
·	leach tanks and agitators
·	CCD circuit
·	barren solution tank
·	detoxofication tanks and agitators
·	detoxification product pump box and tailings pumps

21.2.4.6 Carbon Elution and Regeneration Plant (3600)

The main equipment in this cost includes:

·	loaded carbon screen
·	acid wash carbon column
·	stripping circuit cost includes the following key equipment:

o	elution column
o	strip solution heater with heat exchanger
o	strip and pregnant solution tanks
o	carbon regeneration kiln

21.2.4.7 Merrill Crowe & Refinery (3700)

The Merrill Crowe circuits, electrowinning circuit and gold room costs include the following key equipment:

·	clarifier
·	pregnant solution tank
·	clarification filters
·	deaeration tower
·	precipitate filters
·	mercury retort
·	flux mixer
·	induction furnace with bullion moulds and slag handling system

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·	bullion vault and safe
·	dust and fume collection and gas scrubbing system
·	gold room security system

21.2.4.8 Reagents and Services (3800)

Cost for storage and preparation of key plant reagents include the following:

·	pebble lime
·	sodium cyanide
·	sodium hydroxide
·	hydrochloric acid
·	copper sulphate (pentahydrate)
·	sulphur dioxide
·	activated carbon
·	flocculant
·	coagulant
·	flotation collector (PAX)
·	flotation frother (MIBC)
·	zinc dust
·	lead nitrate
·	oxygen generation

21.2.4.9 Tailings Disposal (3900)

The main equipment in this cost includes:

·	tailings surge tanks and agitators
·	tailings pumps

21.2.4.10 Concrete Supply & Installation

The scope of the civil concrete works allows for all new concrete work in the process plant and relevant on-site facilities as detailed in the WBS.

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Material take-offs were prepared by engineering as part of the 2021 PFS and are based on calculations derived from either factored historical quantities from previously executed projects or estimated from general arrangement drawings and sketches updated with new process equipment.

MTOs for major structures including foundations, footings, walls, pedestals, slab on grade and elevated concrete have been developed based on these calculations. Updated layouts for changes in process design and mechanical equipment have been used to scale MTO quantities in this study in line with new equipment and tank sizing.

Concrete rates are based on a recently completed Ausenco Feasibility Study in Ontario. Budget pricing was sourced from the market for supply and delivery of batched concrete including supply/operation of the batch plant and installation of concrete.

The concrete supply rates are inclusive of supply of batched concrete with separate mobilization and demobilization identified in the field indirects.

The basis for the total cost of installed concrete (excluding growth) is the following:

·	Material costs - includes formworks, concrete from an on-site batch plant, required embedments and reinforcement steel.
·	Labour includes categorized installation hours multiplied by the “All in” labour rate.
·	Total Concrete for the processing plant is US$33.5million, 23% of total installed Mechanical Equipment.

21.2.4.11 Structural Steel

The scope of the structural steel works allows for all new steel work in the process plant and relevant on-site facilities as detailed in the WBS.

All structural steel quantities were estimated from historical executed projects and factored accordingly to align with the Springpole process design. The steel quantities include light, medium, heavy and extra heavy steel designations and miscellaneous steel including grating, handrail and stair treads.

Structural steel pricing is based on recently completed Ausenco Feasibility Studies in Ontario and British Columbia where budget pricing was sourced from the market for supply and delivery to site of fabricated structural steel and other elements such as floor grating, handrailing, stair treads etc.

Total structural steel for the processing plant is US$31.3 million, 22% of total installed mechanical equipment.

21.2.4.12 Architectural

The scope of the architectural buildings work is noted below. Building sizes have been scaled-off general arrangement drawings and from similar historical projects. The Architectural works covers the superstructure portion of the buildings. Overhead cranes, concrete works and internal support steel have been accounted for separately in their respective engineering discipline MTOs. A building list has been developed, for which the cost estimate is based on for the following:

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Pre-Engineered metal-clad buildings:

·	Primary Crusher Building
·	Grinding Building
·	Process Building
·	Gold-room
·	Reagents building

Pre-Engineered fabric-clad buildings:

·	Stockpile Dome
·	Warehouse and Workshop building
·	Truck Shop building

Modular:

·	Administration and Dry building
·	Mine Office
·	Gatehouse
·	Construction camp (will become the Permanent Operations Camp after construction) included in Indirect costs.

Supply and Install rates for the above superstructures are based on a recently completed Ausenco Feasibility Study in Ontario. Budget pricing was sourced from the market for supply and installation packages for the architectural buildings.

Total Architectural for the processing plant is US$29.6 million, 20% of total installed Mechanical Equipment.

21.2.4.13 Mechanical Equipment

The estimate includes the supply and installation of all new mechanical equipment identified in the PFS study for the process plants and related on-site facilities as detailed in the WBS within Ausenco scope.

The mechanical equipment list has been updated to reflect all the required new mechanical equipment. Equipment sizing has been by process engineering and mechanical engineering, with a principle of selecting proven designs used to ensure equipment selection was robust. The mechanical equipment was specified utilising project specific equipment datasheets highlighting agreed upon process performance criteria and were accompanied by typical engineering specifications.

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The items and quantities for the mechanical equipment list are based on the following:

·	PFS Process flow drawings
·	Equipment datasheets
·	Layouts and general arrangement drawings

Pricing for the major mechanical equipment items have been based on budgetary quotations, using competitive pricing submissions from equipment suppliers. Where budgetary quotes were not obtained, existing Ausenco database pricing has been used. For minor equipment, in-house historical pricing and estimates were used.

By cost, 74% of equipment has been priced using budgetary supplier quotations, and 26% from using Ausenco historical data.

Total Process Plant Mechanical Equipment supply costs are US$144.7 million.

21.2.4.14 Platework

The estimate allows for the supply and installation of all platework and liners identified in the PFS study for the process plant and relative onsite facilities as detailed in the WBS within Ausenco scope.

MTOs were prepared for chute work, launders, hoppers, bins, tanks and silos. Platework liners were quantified in metric tonnes or square metres for rubber lining by engineering. This list of platework MTOs is part of the mechanical equipment list.

Platework quantities were prepared by engineering based on design calculations, previous similar designs, drawings and sketches.

Platework pricing is based on a recently completed Ausenco Feasibility Study in Ontario where budget pricing was sourced from the market for supply, shop detailing, fabrication and painting of bulk steel plate and rubber lining rates including delivery to site.

Total Platework for the processing plant is US$25.2 million, which represents 17% of total installed Mechanical Equipment.

21.2.4.15 Process Plant Piping

The estimate allows for the supply and installation of all pipework, fittings, valves, special pipe items and supports in the PFS study for the process plant and relative onsite facilities as detailed in the WBS within Ausenco scope.

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Process plant piping has not been quantified. Process piping costs have been factored off the total installed process plant mechanical costs.

The process plant piping costs have been factored off the total installed mechanical costs by WBS level 3 process areas. The factors allow for pipe, fittings, supports, valves, paint, special pipe items and flanges. The overall factored piping costs is US$30.4 million, and equate to a blended 21% of total installed mechanical costs and aligns with Ausenco’s historical benchmarks.

21.2.4.16 Electrical Equipment

The estimate allows for the supply and installation of all the electrical equipment for the process plant and relative onsite facilities as identified for the PFS study and as detailed in the WBS within Ausenco scope.

An Electrical Equipment List (EEL) has been prepared and aligns with the current mechanical equipment list and load list.

Electrical equipment pricing carried in the estimate has been based on budgetary vendor supply quotations for major equipment, and Ausenco’s historical data from recent projects for minor equipment. Total Electrical Equipment supply costs is US$21.8 million or 15% of total installed mechanical.

21.2.4.17 Electrical Bulks

The estimate allows for the supply and installation of all electrical bulks in the PFS study for the process plant and relative onsite facilities as detailed in the WBS within Ausenco scope.

MTOs for electrical bulks have not been developed. Costs have been factored from the installed mechanical equipment costs for each WBS level 3 process area. The factors include for cables, tray, terminations, wiring, testing, grounding, installation, etc.

The overall electrical bulk costs total US$18.5million and equates to a blended 13% of total installed mechanical costs and aligns with Ausenco’s historical benchmarks.

21.2.4.18 Instrumentation

The estimate allows for the supply and installation of all instruments and bulks in the PFS study for the process plant and relative onsite facilities as detailed in the WBS within Ausenco scope. An allowance for a plant control system has been included as well.

MTOs for process instrumentation have not been developed. Costs have been factored from the installed mechanical equipment costs for each WBS level 3 process area.

The overall instrumentation bulk factored costs total US$12.5 million and equates to a blended 9% of total installed mechanical costs and aligns with Ausenco’s historical benchmarks.

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21.2.5 Infrastructure Capital Costs

The scope covers the following:

·	On-site infrastructure earthworks
·	Internal roads
·	Off-site roads
·	Power supply and distribution
·	Water treatment plants and utilities
·	Plant Truckshop
·	Plant services (fuel, air, water, power)

Material take offs (MTO’s) have been developed for the agreed scope of work using two main methods. Quantities for bulk excavations and backfilling for plant and roads have been generated from AutoCAD Civil 3D models, subsequent quantities for items such as trenching, drainage, retaining, structural cut, detailed backfill have been taken from AutoCAD 2D drawings and design sketches. Fill material includes site won, imported and mine waste materials, the assumption that structural fill material is available from a quarry 2 km from the process site has been made.

Sub-contract rates have been based on recently completed Ausenco Feasibility Studies in Ontario and British Columbia where budget pricing was sourced from the market.

These costs are summarized below in Table 21‑8.

Table 21‑8: Onsite and Offsite infrastructure Initial Direct Costs

WBS	Level 1 Description	TOTAL CAPEX<br> <br>(US$M)

| 4100 | Power Supply & Distribution | 21.5 |

| 4200 | Anciliary Buildings | 17.0 |

| 4300 | Site Services | 35.5 |

| 4400 | Mobile Equipment | 1.5 |

| Total On-Site Infrastructure Direct Costs | | 75.5 |

| 5100 | Off-site Roads | 0.5 |

| 5200 | Power Supply | 46.2 |

| 5300 | Water Supply | 0.3 |

| Total Off-Site Infrastructure Direct Costs | | 47.0 |

| Total | | 122.6 |

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21.2.5.1 Overhead Powerline to Site and Switch Station

The costs for the main powerline to site (230 kV, 89 km long) are included in Table 21‑9. This pricing was originally provided by First Mining based on an estimate from a third party in the 2021 PFS. This estimate has been escalated to 2025 and benchmarked against recent, suitable reference projects to ensure accurate pricing. It is an all-inclusive cost (design, supply and install) for their scope of work, as per the following:

·	Grid tie-in
·	230 kV Overhead Power line to site – 89 km long

Substation pricing was provided by First Mining with a 2025 contractor quotation.

Table 21‑9: Third-Party (OHPL) Estimate Summary

WBS	LVL 1 Description	Total CAPEX<br> <br>(US$M)

| 4110 | Main Substation | 6.0 |

| 5200 | Overhead Powerline – 89 km | 46.2 |

| Total Direct Costs | | 52.2 |

| 6000 | Indirects | Incl with Direct costs |

| 7000 | EPCM Services | Incl with Direct costs |

| Total Indirect Costs | | Incl with Direct costs |

| 9000 | Provisions (Contingency) | 5.3 |

| Project Total | | 57.5 |

21.2.6 Indirect Capital Costs (WBS 6000)

Project Indirect Costs include all costs that are necessary for project completion but are not directly attributable to the construction of specific physical facilities of the plant or associated infrastructure but are required to be provided as support during the construction period. These items are as follows:

·	Field Indirects
·	Camp Accommodations and Messing
·	Vendor Representatives
·	Spares and First Fills
·	Startup and Commissioning
·	Heavy-lift Cranage (>100 t capacity)

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·	EPCM and Expenses
·	Owner’s Costs.

21.2.6.1 Field Indirects (WBS 6100)

Field Indirects are items or services provided by either the Owner or EPCM contractor as common facilities and services which are not covered/managed by the general contractor’s indirect costs (distributable costs).

These costs typically include but are not limited to:

·	Temporary construction facilities – site office for EPCM contractor, site services, temporary fencing, temporary roads and parking.
·	Construction support – site clean-up and waste disposal, material handling, maintenance of buildings and roads, testing and training, service labour, site transport, site surveys, QA-QC and security.
·	Common pool of construction equipment, tools and supplies purchased by the owner or EPCM contractor – mobile equipment and tools, consumables, and purchased utilities.
·	Material transportation & storage on-site incurred by the owner or EPCM contractor – agents, staging and marshalling etc.
·	Heavy lift equipment - an allowance for heavy lift cranage (>100 t capacity cranage) has been included for extra-heavy lifts which are not included in the contractor’s distributables (within the all-in SMP labour rates).

Field Indirects have been factored from the Total Direct Costs (less Mining) at a combined 3.6% and total US$19.6 million.

21.2.6.2 Construction Camp and Catering (WBS 6200)

A Construction camp has been allowed for in the estimate and will become the permanent operations accommodations upon completion of the construction. It is sized for a new 450-bed modular style construction camp including a commercial kitchen, dining/recreation room, laundry facility. The camp cost (based on Ausenco’s historical data) is based on US$51,800/bed for construction.

In addition, a cost for operations and maintenance of the camp during the construction period of the Project has been included and is based on Ausenco’s historical data at a cost of US$59.2/person per day.

Costs are summarized in Table 21‑10 below.

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Table 21‑10: Construction Camp Costs

WBS	Description	Cost<br> <br>US$M	Comments

| 6220 | Construction camp and messing | 33.4 | 450 beds at $70,000 CAD per bed $80/person/day |

| Total | | 33.4 | - |

21.2.6.3 Vendor Representatives (WBS 6400)

Costs for vendor representatives for construction and commissioning have been allowed for in the cost estimate and based on Ausenco’s historical percentages of 0.5% of the total supply cost of equipment, totaling US$1.3 million.

21.2.6.4 Spares (WBS 6500)

Major mechanical spares for construction and commissioning have been included in the estimate, based on Ausenco’s historical percentages of total equipment supply. See below Table 21‑11.

Table 21‑11: Spares Costs

WBS	Description	Cost<br> <br>US$M	Comments

| 6510 | Commissioning Spares | 1.3 | 1% of total equipment supply |

| 6520 | Capital (Critical) Spares | 5.3 | 4% of total equipment supply |

| 6530 | Operating Spares | 0.6 | 0.5% of total equipment supply |

| Total | | 7.2 | - |

21.2.6.5 First Fills (WBS 6500)

First fills include the costs for the initial construction first fills for installed equipment and process commissioning first fills which consist of chemicals, fuels and lubricants etc. and have been calculated by first principles.

21.2.6.6 EPCM Costs (WBS 7000)

EPCM services costs cover such items as engineering and procurement services (home office based), construction management services (site based), project office facilities, IT, staff transfer expenses, secondary consultants, field inspection and expediting, corporate overhead and fees. An amount of US$70.3 million has been included in the cost estimate for Ausenco’s EPCM services. AGP’s EPCM services is included in their direct mining costs in Section 21.2.3.

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21.2.7 Owner (Corporate) Capital Costs (WBS 8000)

Ausenco, on behalf of the client, has included the PFS Owner’s costs into the cost estimate as a placeholder. An amount equivalent to 3% of total directs costs has been included.

Owner’s costs typically include, but are not limited to, the following:

·	Corporate Overheads & Office
·	Environmental Monitoring
·	Site Office
·	Setup & Running Costs
·	Staff & Labour
·	Bonding

21.2.8 Estimate Contingency (WBS 9000)

Contingency is a provision of funds for unforeseen or inestimable costs within the defined project scope relating to the level of engineering effort undertaken and estimate/engineering accuracy and applied to provide an overall level of confidence in costs and schedule outcomes (in this case, targeting an 50% confidence level). The contingency is meant to cover events or incidents that occur during the course of the Project which cannot be quantified during the estimate preparation and does not include any allowance for Project risk.

It is important to note that contingency does not cover scope changes, force majeure, adverse weather conditions, and changes in government policies, currency fluctuations, escalation and other project risks.

A summary of the contingency contributors is noted below Table 21‑12.

Table 21‑12: Contingency Summary

Scope Owner	Cost<br> <br>US$M

| Contingency – Ausenco | 99.3 |

| Contingency – AGP Mining | 11.3 |

| Contingency – WSP | 16.9 |

| Contingency – Owners | - |

| Total | 127.5 |

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21.2.9 Sustaining Capital

Sustaining capital is only required for Mining (WBS1000) as there are no plant expansions with the updated mine plan and associated ore.

Table 21‑13: Estimate Summary Level 1 Sustaining Capital

WBS Level 1	Description	TOTAL SUSTAINING<br> <br>(US$M)

| 1000 | Mining | 304.1 |

| 2000 | Site development | 0 |

| 3000 | Process plant | 0 |

| 4000 | On-site infrastructure | 0 |

| 5000 | Off-site infrastructure | 0 |

| 8000 | Owner cost | 18.5 |

| Project Total | | 322.6 |

21.2.10 Closure Costs

Closure costs include covering of the waste rock storage area at the end of the mine life, noting that the current mine plan incorporates progressive reclamation during mining. Additional costs have been allocated for remediation of the depleted stockpiles, and the removal of the processing plant and site facilities in addition to the normal site closure costs.

Additional to the closure cost is the restoration of the lake area with the removal breaching of the dikes. Additional lake area is created within the mine quarry footprint upon closure and the lake level reaching its pre-mining level.

Mine closure costs have been estimated by AGP at US$40.5 million. A salvage credit equal to 10% of mining equipment direct costs has been applied to the overall closure cost, resulting in a net closure estimate of US$36.5 million. Plant closure costs have been excluded.

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21.3 Operating Costs

21.3.1 Overview

The costs considered on-site operating costs are those related to mining, processing, maintenance, power, water treatment and general and administrative activities.

Total Mining costs are US$12.44/t. Mining costs (mined) are US$2.57/t. The average process operating cost is US$10.72/t processed, and the annual G&A costs are US$27 million.

A summary of the average operating costs is shown below in Table 21‑14.

Table 21‑14: Operating Cost Summary

Cost Area	Total (US$M/a)	US$/t

| Mining | 106.8 | 12.44 |

| Process | 113.1 | 10.72 |

| G&A | 27.0 | 2.56 |

| Total | 246.9 | 23.15 |

21.3.2 Mine Operating Costs

The mine operating costs have been estimated from base principles with vendor quotations for repair and maintenance costs and other suppliers for consumables. Key inputs to the mine cost are fuel and labour. The price provided for the Project was US$1.014/litre delivered to the site. The mine truck and support equipment fleets are diesel powered. The large production drills, hydraulic shovels and dewatering pumps are electric powered and used a price of US$0.0488/kW hr.

21.3.2.1 Labour

Labour costs for the various job classifications were obtained from salary surveys in Ontario and other operations. A burden rate of 40% was applied to the various rates. Labour was estimated for both staff and hourly on a 12-hour shift basis utilizing a rotation of either two weeks on/two weeks off or four days on, three days off. Mine positions and salaries are shown in Table 21‑15.

Table 21‑15: Mine Staffing Requirements and Annual Employee Salaries

Position	Estimated Number of Personnel	Annual Salary<br> <br>(US$/a)

| Mine Maintenance | | |

| Maintenance Superintendent | 1 | 170,940 |

| Maintenance General Supervisor | 1 | 158,508 |

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Position	Estimated Number of Personnel	Annual Salary<br> <br>(US$/a)

| Maintenance Shift Supervisor | 4 | 119,140 |

| Maintenance Planner/Contract Administration | 2 | 129,293 |

| Clerk | 1 | 82,155 |

| Subtotal | 9 | |

| Mine Operations | | |

| Mine Operations/Technical Superintendent | 1 | 170,940 |

| Mine General Supervisor | 1 | 158,508 |

| Senior Shift Supervisor | 4 | 134,369 |

| Junior Shift Supervisor | 4 | 108,262 |

| Road Crew/Services Supervisor | 1 | 119,658 |

| Clerk | 1 | 82,155 |

| Subtotal | 12 | |

| Mine Engineering | | |

| Chief Engineer | 1 | 190,624 |

| Senior Engineer | 1 | 136,752 |

| Open Pit Planning Engineer | 2 | 122,870 |

| Geotechnical Engineer | 1 | 119,658 |

| Blasting Engineer | 1 | 119,658 |

| Blasting/Geotechnical Technician | 2 | 79,772 |

| Dispatch Technician | 1 | 79,772 |

| Surveyor/Mining Technician | 2 | 101,735 |

| Surveyor/Mining Technician Helper | 2 | 74,074 |

| Clerk | 1 | 82,155 |

| Subtotal | 14 | |

| Geology | | |

| Chief Geologist | 1 | 170,318 |

| Senior Geologist | 1 | 139,860 |

| Grade Control Geologist/Modeler | 2 | 109,298 |

| Sampling/Geology Technician | 4 | 85,010 |

| Clerk | 1 | 82,155 |

| Subtotal | 9 | |

| Total | 44 | |

The mine staff labour remains constant from Year 2 until Year 8 when positions are removed as the mine winds down. During the pre-production period and Year 1 there are trainer positions in mine operations.

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Hourly employee labour force levels in mine operations and maintenance fluctuate with production requirements. The Year 4 hourly labour requirements are shown in Table 21‑16.

Table 21‑16: Hourly Manpower Requirements and Annual Salaries (Year 4)

Position	Estimated Number of Personnel	Annual Salary<br> <br>(US$/a)

| Mine General | | |

| General Equipment Operator | 8 | 83,186 |

| Road/Pump Crew | 8 | 83,186 |

| General Mine Labourer | 8 | 82,699 |

| Trainee | 4 | 73,700 |

| Light Duty Mechanic | 3 | 126,481 |

| Tire Technician | 4 | 111,400 |

| Lube Truck Driver | 4 | 105,562 |

| Subtotal | 39 | |

| Mine Operations | | |

| Driller | 20 | 109,454 |

| Blaster | 2 | 109,454 |

| Blast Helper | 4 | 81,726 |

| Loader Operator | 8 | 114,319 |

| Hydraulic Shovel Operator | 8 | 114,319 |

| Haul Truck Driver | 80 | 107,022 |

| Dozer Operator | 12 | 108,477 |

| Grader Operator | 6 | 108,477 |

| Crusher Loader Operator | 3 | 108,477 |

| Snowplow/Water Truck | 7 | 83,186 |

| Subtotal | 150 | |

| Mine Maintenance | | |

| Heavy Duty Mechanic | 36 | 130,737 |

| Welder | 21 | 130,737 |

| Electrician | 1 | 130,737 |

| Apprentice | 7 | 116,386 |

| Subtotal | 65 | |

| Total | 254 | |

Labour costs are based on an Owner-operated scenario with First Mining responsible for the maintenance of the equipment with its own employees.

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Overseeing all the mine operations, maintenance, engineering, and geology functions will be a Technical Superintendent. This person would have the Mine General Supervisor and Maintenance Superintendent reporting to them, as well as the Chief Engineer and Chief Geologist.

The Mine General Supervisor would have the Shift Supervisors report directly to them.

The mine will have four mine operations crews, each with a Senior Shift Supervisor who will have one Junior Shift Supervisor reporting to them. For the mine life, there will also be a Road Crew/Services Supervisor responsible for roads, drainage, and pumping around the mine. This person would also be a backup Senior Mine Shift Supervisor. The Training Supervisor will only be required on site until the end of Year 1, at which time the position will be eliminated. The Mine Operations department will have its own clerk.

The Chief Engineer will have one Senior Engineer and two open pit engineers reporting to them. The Blasting Engineer will be included in the short-range planning group and will double as Drill and Blast Supervisor as required. The Geotechnical Engineer will cover all aspects of the wall slopes and waste dumps together with shared technicians in blasting.

The short-range planning group in engineering will also have two surveyor/mine technicians and two surveyors/mine helpers. These people will assist in the field with staking, surveying, and sample collection with the geology group; the group will have a clerk to assist the team.

In the Geology department, there will be one Senior Geologist reporting to the Chief Geologist. There will also be two grade control geologists/modellers; one will be in short range and grade control drilling, and the other will be in long range/reserves. There will also be four grade control/sampling technicians and one clerk.

Four Mine Maintenance Shift Supervisors will report to the Maintenance General Supervisor who in turn will report to the Maintenance Superintendent. As well, there will be two maintenance planners/contract administrators and a clerk.

The hourly labour force will include positions for the light duty mechanic, tire technicians, and lube truck drivers. These positions will all report to Maintenance. There will generally be one of each position per crew. Other general labour will include General Mine Labourers (two per crew) and Trainees (one per crew) plus two road/pump crew personnel per crew for water management/snow removal.

The drilling labour force is based on one operator per drill, per crew while operating. This peaks at 24 drillers in Year 5 and then drops down over time as the drilling hours are diminished.

Shovel and loader operators will peak at 20 in Year 6. Haulage truck drivers will peak at 100 in Year 6 and then drop off to the end of the mine life.

Maintenance factors are used to determine the number of heavy duty mechanics, welders and electricians that are required and are based on the number of equipment operators. Heavy duty mechanics work out to 0.25 mechanics required for each drill operator for example. Welders are 0.25 per operator and electricians are 0.05 per operator.

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The number of loader, truck and support equipment operators is estimated using the projected equipment operating hours. The maximum number of employees is four per unit to match the mine crews.

21.3.2.2 Equipment Operating Costs

The vendors provided repair and maintenance (R&M) costs for each piece of equipment selected for the PFS. Fuel consumption rates were estimated from the supplied information and knowledge of the working conditions. The costs for the R&M are expressed in US$/h form.

Tire costs were also collected from various vendors for the sizes expected to be used. Estimates of tire life are based on AGP’s experience and discussions with the vendors. The operating cost of the tires is also expressed in a US$/h form. The life of the haulage truck tires is estimated at 4,500 h/tire with proper rotation from front to back. Each truck tire is estimated to cost US$51,467 so the cost per hour for tires is US$68.62/h for the truck using six tires in the calculation.

Ground engaging tools (GET) costing is estimated from other projects and is an area that would be refined once the Project is operational.

Drill consumables are estimated as a complete drill string using the parts list and component lives provided by the vendor. Drill productivity is estimated at 25.7 m/h for mill feed and waste rock material. The equipment costs used in the estimate are shown in Table 21‑17.

Table 21‑17: Major Equipment Operating Costs – No Labour (US$/h)

Equipment	Fuel/Power	Lube/Oil	Tires/ Undercarriage	Repair & Maintenance	GET/ Consumables	Total

| | (US$/h) | (US$/h) | (US$/h) | (US$/h) | (US$/h) | (US$/h) |

| Production Drill (140mm) | 91.24 | 9.12 | - | 94.31 | 92.30 | 286.97 |

| Production Drill (251mm) | 12.21 | - | 4.88 | 72.88 | 162.33 | 252.30 |

| Production Loader (23m3) | 111.52 | 11.15 | 104.19 | 145.34 | 12.21 | 384.42 |

| Hydraulic Shovel (37m3) | 110.87 | - | - | 278.28 | 13.02 | 402.18 |

| Haulage Truck (240 t) | 148.01 | 14.80 | 68.62 | 139.55 | 4.88 | 375.88 |

| Haulage Truck (91 t) | 76.04 | 7.61 | 17.03 | 84.55 | 2.44 | 187.66 |

| Track Dozer | 76.04 | 7.61 | 8.14 | 58.31 | 4.07 | 154.16 |

| Grader | 22.30 | 2.23 | 9.57 | 41.29 | 4.07 | 79.47 |

| Production Excavator | 60.83 | 12.17 | - | 125.06 | 6.51 | 204.57 |

21.3.2.3 Drilling

Drilling in the open pit will use down the hole hammer drill rigs with 140 mm bits with the small diesel drill and rotary bits with the 251 mm electric drill. The material is designed to be blasted smaller and finer to improve productivity and reduce maintenance costs as well as improve plant performance. The drilling pattern parameters are shown in Table 21‑18.

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Table 21‑18: Drill Pattern Specifications

Specification	Unit	Production Drill - Small			Production Drill – Large

| Bench Height | m | 12 | 12 | 12 | 12 | 12 |

| Sub-drill | m | 0.8 | 0.8 | 0.0 | 1.30 | 1.30 |

| Blasthole Diameter | mm | 140 | 140 | 140 | 251 | 251 |

| Pattern Spacing - Staggered | m | 4.8 | 4.6 | 1.65 | 7.7 | 7.7 |

| Pattern Burden – Staggered | m | 4.2 | 4.0 | 2.0 | 6.7 | 6.7 |

The sub-drill is included to allow for caving of the holes in weaker zones, reducing re-drill requirements or short holes that would affect bench floor conditions.

The parameters used to estimate drill productivity are shown in Table 21‑19

Table 21‑19: Drill Productivity Criteria

Drill Activity	Unit	Small Drill	Production Drill

| Pure Penetration Rate | m/min | 0.55 | 0.50 |

| Hole Depth | m | 12.8 | 13.3 |

| Drill Time | min | 23.3 | 26.6 |

| Move, Spot and Collar Hole | min | 3.00 | 3.00 |

| Level Drill | min | 0.50 | 0.50 |

| Add Steel | min | 0.50 | 0.0 |

| Pull Drill Rods | min | 1.50 | 1.0 |

| Total Setup/Breakdown Time | min | 5.50 | 4.50 |

| Total Drill Time per Hole | min | 28.8 | 31.1 |

| Drill Productivity | m/h | 26.7 | 25.7 |

21.3.2.4 Blasting

An emulsion product will be used for blasting to provide water protection. With the high rainfall, large snowmelt and working below lake level it is expected that a water-resistant explosive will be required. The powder factors used in the explosive calculation are shown in Table 21‑20.

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Table 21‑20: Design Powder Factors

	Unit	Small Drill		Production Drill

| Powder Factor | kg/m^3^ | 0.75 | 0.84 | 0.83 | 0.83 |

| Powder Factor | kg/t | 0.28 | 0.31 | 0.29 | 0.29 |

The blasting cost is estimated using quotations from a local explosives vendor. The emulsion price is US$77.49/100 kg. The mine will be responsible for guiding the loading process, including placement of boosters/Nonels, and stemming and firing the shot.

The explosives vendor will also lease the explosives and accessories for a monthly cost. A service charge for the vendors pickup trucks, pumps, labour, and cost of the explosives plant are included. The total monthly cost is estimated at US$258,704/month.

21.3.2.4.1 Loading

Loading costs for both mill feed and waste are based on the use of electric hydraulic shovels and front-end loaders. The shovels will be the primary diggers with the front-end loaders as backup/support units. The average percentage of each material type that the various loading units are responsible for is shown in Table 21‑21. This highlights the focus of the shovels over the loaders.

Table 21‑21: Loading Equipment Parameters

	Unit	Electric Hydraulic Shovel	Front End Loader

| Bucket Capacity | m^3^ | 37 | 23 |

| Truck Capacity Loaded | t | 240 | 240 |

| Waste Tonnage Loaded | % | 70 | 30 |

| Mill Feed Tonnage Loaded | % | 75 | 25 |

| Bucket Fill Factor | % | 91 | 95 |

| Cycle Time | sec | 38 | 42 |

| Trucks present at loading unit | % | 80 | 80 |

| Loading Time | min | 2.60 | 4.20 |

The trucks present at the loading unit refers to the percentage of time a truck is available to be loaded. To maximize truck productivity and reduce operating costs, it is more efficient to slightly under-truck the loading unit. One of the largest operating cost items is haulage and minimizing this cost by maximizing the truck productivity is crucial to lower operating costs. The value of 80% comes from the standby time shovels typically encounter due to a lack of trucks.

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21.3.2.5 Hauling

Haulage profiles were determined for each pit phase for the primary crusher or the waste rock facility destinations. Cycle times were generated for the appropriate period tonnage by destination and phase to estimate the haulage costs. Maximum speed on the trucks is limited to 50 km/h for tire life and safety reasons although few locations in the mine plan appeared to offer the truck the opportunity to accelerate to that velocity.

21.3.2.6 Support Equipment

Support equipment hours and costs are determined on factors applied to various major pieces of equipment. For the PFS some of the factors used are shown in Table 21‑22.

Table 21‑22: Support Equipment Operating Factors

Mine Equipment	Factor	Factor Units

| Track Dozer | 25% | Of haulage hours to maximum of 4 dozers |

| Grader | 15% | Of haulage hours to maximum of 3 graders |

| Crusher Loader | 20% | Of loading hours to maximum of 1 loader |

| Snowplow/Water Truck | 9% | Of haulage hours to maximum of 2 trucks |

| Pit Support Backhoe | 8% | Of loading hours to maximum of 1 backhoe |

| Road Crew Backhoe | 3 | hours/day/unit |

| Road Crew Dump Truck | 3 | hours/day/unit |

| Road Crew Loader | 3 | hours/day/unit |

| Lube/Fuel Truck | 6 | hours/day/unit |

| Mechanics Truck | 12 | hours/day/unit |

| Blasting Loader | 8 | hours/day/unit |

| Blaster’s Truck | 8 | hours/day/unit |

| Integrated Tool Carrier | 4 | hours/day/unit |

| Light Plants | 12 | hours/day/unit |

| Pickup Trucks | 10 | hours/day/unit |

These factors resulted in the need for four track dozers, three graders, and one support backhoe. Their tasks include clean-up of the loader faces, roads, waste storage facility, and blast patterns. The graders will maintain the crusher and waste haul routes. In addition, snowplows/water trucks have the responsibility for patrolling the haul roads for snow removal and controlling fugitive dust for safety and environmental reasons. The small backhoe and road crew dump trucks will be responsible for cleaning out sedimentation ponds and water ditch repairs.

The hours generated in this manner are applied to the individual operating costs for each piece of equipment. Many of these units are support equipment so no direct labour is allocated to them due to their variable function. The operators come from the General Equipment operator pool.

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21.3.2.7 Grade Control

Grade control will be completed with a separate fleet of reverse circulation (RC) drill rigs. They will drill the deposit off on a 10 x 5 m pattern in areas of known mineralization taking samples each metre. The holes will be inclined at 60°.

In areas of low-grade mineralization or waste the pattern spacing will be 20 x 10 m with sampling over 6 m. These holes will be used to find undiscovered veinlets or pockets of mineralization. Over the life of the mine, 769,750 m of drilling are expected to be completed for grade control work. A total of 862,000 samples will be assayed from that drilling.

These grade control holes serve to define the mill feed grade and contacts.

Samples collected will be sent to the assay laboratory and assayed for use in the short range mining model.

No additional costing for blasthole sampling has been included. This may be an opportunity when more knowledge has been gained operationally on the gold deportment and representativeness of blast hole samples.

Costs associated with this separate drill program are tracked as a distinct line item for the mining cost. The drill crew is one driller and two helpers with oversight by the Mine Geology department. The cost of this drilling is expected to average over US$4.8 million/a.

21.3.2.8 Dewatering

Pit dewatering will be an important part of mining. Significant volumes need to be pumped initially to allow the open pit to advance as the open pit basin is being dewatered, in addition to the normal rain/snow amounts.

Reviewing past data collected and comparing this to the proposed mining area allowed AGP to make an estimate at a PFS level for the water volume required to be pumped. Initial pumping in Year -1 is expected to be 1.8 Mm^3^. That climbs to 6.4 Mm^3^/a in Year 8 as the pit drives deeper.

The dewatering will be completed with a set of six pumps in the pit and two pumps on the surface to push the water to the settling ponds. These pumps will be electric to reduce the cost of this operation.

Additional dewatering in the form of horizontal drill holes are part of the dewatering costs. These holes will be campaigned and will be part of the sustaining mine capital.

Dewatering is expected to cost US$4.4 million over the mine life.

21.3.2.9 Leasing

Leasing of the mine fleet is considered a viable option to reduce initial capital. Various vendors offer this as an option to help select their equipment.

Indicative terms for leasing provided by the vendors are:

·	down payment	= 20% of equipment cost

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·	term length	= 3-5 years (depending on equipment)
·	interest rate	= CORRA (2.5%) plus 6.5%
·	residual	= $0

The initial capital, down payments, and annual leasing costs were shown previously in the capital cost area of this section.

The support equipment fleet is calculated in the same manner as the major mining equipment.

All of the major mine equipment, and the majority of the support equipment where it was considered reasonable, is leased. If the equipment has a life greater than the lease term length, then the following years onward of the lease do not have a lease payment applied. In the case of the mine trucks, with an approximate 10-year working life, the lease would be complete, and the trucks would simply incur operating costs after that time. For this reason, the operating cost would vary annually depending on the equipment replacement schedule and timing of the leases.

Utilizing the leasing option adds US$0.59/t moved to the mine operating cost over the life of the mine. On a cost per tonne of mill feed basis, it is US$2.58/t mill feed.

21.3.2.10 Total Mine Costs

The total LOM operating costs per tonne of material mined, moved (including rehandle) and per tonne of mill feed processed are shown in Table 21‑23.

In the “General” category is the cost associated with a Contractor-operated crushing plant to make stemming material and road crush. That cost averages approximately US$3.8 million/a.

Portions of the estimated total LOM operating costs have been allocated to different categories for economic analysis purposes. Table 21‑24 shows the quantities of the operating cost estimate that have been allocated to the different categories, notably Pre-Production stripping, Capitalized stripping, and the CDF Waste Storage Facility.

Table 21‑23: Total Open Pit Mine Operating Cost Estimate - With Leasing

Cost Area	Total (US$M)	US$/t<br> <br>mined	US$/t moved	US$/t milled	% of Total

| General | 126.91 | 0.31 | 0.28 | 1.24 | 10% |

| Drill | 77.55 | 0.19 | 0.17 | 0.76 | 6% |

| Blast | 157.03 | 0.38 | 0.36 | 1.54 | 12% |

| Load | 89.54 | 0.21 | 0.20 | 0.88 | 7% |

| Haul | 442.45 | 1.07 | 0.99 | 4.34 | 34% |

| Support | 106.63 | 0.26 | 0.24 | 1.04 | 8% |

| Grade Control | 39.74 | 0.10 | 0.09 | 0.39 | 3% |

| Dewatering | 4.37 | 0.01 | 0.01 | 0.04 | 0% |

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Cost Area	Total (US$M)	US$/t<br> <br>mined	US$/t moved	US$/t milled	% of Total

| Subtotal | 1044.21 | 2.54 | 2.34 | 10.23 | - |

| Leasing Costs | 263.07 | 0.64 | 0.59 | 2.58 | 20% |

| Total | 1307.28 | 3.17 | 2.93 | 12.82 | - |

Table 21‑24: Adjusted Open Pit Mine Operating Cost Estimate - With Leasing

Cost Area	Units	Total<br> <br>Estimated<br> <br>Cost	Allocated to<br> <br>Pre-Production Stripping	Allocated to Capitalized Stripping	Allocated to<br> <br>CDF	Total<br> <br>Production<br> <br>OPEX

| General | US$M | 126.9 | 11.7 | - | - | 115.2 |

| Drill | US$M | 77.6 | 3.8 | - | - | 73.7 |

| Blast | US$M | 157.0 | 8.4 | - | - | 148.7 |

| Load | US$M | 89.5 | 4.2 | - | - | 85.3 |

| Haul | US$M | 442.4 | 14.5 | 155.8 | - | 272.2 |

| Support | US$M | 106.6 | 5.8 | 34.6 | - | 66.2 |

| Grade Control | US$M | 39.7 | 1.5 | - | - | 38.3 |

| Dewatering | US$M | 4.4 | 0.1 | - | - | 4.2 |

| Subtotal | US$M | 1044.2 | 50.1 | 190.4 | 0.0 | 803.8 |

| Leasing Costs | US$M | 263.1 | 27.4 | - | 34.0 | 201.7 |

| Total | US$M | 1307.3 | 77.5 | 190.4 | 34.0 | 1005.4 |

| Mined Tonnage | Mt | 411.5 | 20.0 | 391.5 | - | 391.5 |

| Unit Cost | US$/t mined | 3.17 | 3.87 | 0.49 | - | 2.57 |

21.3.3 Process Operating Costs

The process operating cost estimate is based on a throughput of 30,000 t/d. The process consists of crushing, grinding, flotation, concentrate regrind, concentrate and tailings dewatering and leaching, concentrate leach CCD, carbon recovery, concentrate and tailings Merrill Crowe, cyanide detoxification, and water treatment.

The process operating costs average to US$10.72/t over the LOM. Table 21‑25 provides a detailed summary of process plant operating costs over the LOM.

Table 21‑25: Process Plant Operating Cost Summary

Cost Area	Total (US$M/a)	US$/t	% of Total

| Reagents | 60.44 | 5.72 | 53.4% |

| Consumables | 18.71 | 1.77 | 16.5% |

| Plant Maintenance | 3.93 | 0.37 | 3.5% |

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Cost Area	Total (US$M/a)	US$/t	% of Total

| Power | 15.41 | 1.46 | 13.6% |

| Laboratory | 0.42 | 0.04 | 0.4% |

| Labour (O&M) | 12.84 | 1.22 | 11.4% |

| Mobile Equipment | 1.38 | 0.13 | 1.2% |

| Total | 113.1 | 10.72 | 100% |

21.3.4 Basis of Estimate

The following items provide the basis of estimate for the processing operating cost estimate:

·	Power cost of US$0.0488/kWh was based on in-house benchmarks.
·	A diesel cost of US$1.014/litre was provided via quotation from AGP Mining.
·	Process plant throughput of 30,000 t/d or 10.95 Mt/a was used.
·	Crushing circuit availability is assumed to be 75%, grinding, flotation, and leaching availability is assumed to be 92%.
·	Process plant material and equipment are purchased as new.
·	Process plant staff count is estimated from first principles to operate and maintain the respective process unit operations, salaries are based on in-house benchmarks.
·	Process plant mobile equipment is leased and then owned.
·	Reagent consumption rates are based on metallurgical testwork results and unit costs from vendor quotations.

21.3.4.1 Reagents and Consumables

Reagents, grinding media, and various processing consumables are required in the process. The consumption rates of each of the consumable items are based on the metallurgical testwork and design criteria outlined in Section 13 and Section 17 respectively at the design process plant throughput of 30,000 t/d. The total costs of the reagents averages to US$60.44M/a and the total costs of the processing consumables averages to US$18.71M/a.

21.3.4.2 Maintenance Consumables

Annual maintenance consumable costs were calculated based on a total installed mechanical capital cost by area using a weighted average factor of 4%. The total maintenance consumable operating cost is US$3.93M/a or US$0.37/t of feed.

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21.3.4.3 Power

Power operating costs are calculated from an estimate of annual power consumption using a unit cost of US$0.0488/kWh. The annual power consumption for the processing plant is based on the average utilization of each motor on the mechanical equipment list for the process plant.

The process plant nominal energy consumption is estimated to be 327.3 GWh/a (29.9 kWh/t). The total power operating cost is US$15.41M/a or US$1.46/t of feed.

21.3.4.4 Laboratory Services

Operating costs for the assay lab were estimated based on a review of benchmark projects with similar flowsheets and sampling requirements. The lab is expected to handle grade control samples, mill solid and aqueous samples, water testing, doré quality testing, and other miscellaneous tests as required. Annual operating costs (excluding labour and power) are estimated at US$0.42M/a.

21.3.4.5 Labour

Labour includes all processing, assay laboratory and maintenance labour costs.

Processing production labour was developed using benchmarks from similar projects and includes operation departments such as metallurgy, mill operations, maintenance and the assay lab.

Each position was defined and classified as salary and wages. Costs included taxes and benefits. For process labour the annual cost is US$12.84M/a or US$1.22/t of feed.

The estimated total labour force was estimated at 134. The estimate was based on providing a labour force to support continuous operations at 24 h/day and 365 days/a.

21.3.4.6 Mobile Equipment

Vehicle costs are based on a scheduled number of light vehicles and mobile equipment for the process plant. The costs include fuel, annual maintenance, spares and tires, annual insurance, and equipment leasing costs. Mobile equipment costs for the process plant are estimated at US$1.38 M/a.

21.3.5 Infrastructure Operating Costs

Infrastructure operating costs have been considered under general and administrative operating costs. These are detailed below in Section 21.3.6.

21.3.6 General and Administrative Operating Costs

General and administrative (G&A) operating costs cover the expenses of the operating departments, and a summary is presented in Table 21‑26.

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G&A costs were developed using Ausenco’s in-house data on existing Canadian operations. The costs were estimated based on the following items:

·	Site maintenance (including G&A mobile equipment, on-site road maintenance),
·	Labour (management, procurement, IT, safety & security, etc.),
·	Camp and rotational travel cost,
·	Human Resources and Health & Safety (including personnel, recruiting, training, community relations, personal protective equipment and first-aid),
·	Environmental (including sampling and TSF operation),
·	Water treatment,
·	Administration Costs (including office supplies, IT hardware & software, and support services), and
·	Contract services (including insurance, sanitation, licence fees and legal fees).

The total annual G&A cost was estimated at US$27M/a during production or US$2.56/t plant feed.

Table 21‑26: G&A Cost Summary

G&A Expenses	Average Annual Costs (US$M)	US$/t Processed

| Labour | 5.14 | 0.49 |

| Camp & Travel | 10.26 | 0.97 |

| G&A - Expenses | 7.49 | 0.71 |

| Water treatment | 4.06 | 0.38 |

| Total | 27.0 | 2.56 |

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22 ECONOMIC ANALYSIS

22.1 Forward-Looking Information

The results of the economic analyses discussed in this section represent forward-looking information as defined under Canadian securities law. The results depend on inputs that are subject to known and unknown risks, uncertainties, and other factors that may cause actual results to differ materially from those presented here. Information that is forward looking includes the following:

·	Mineral resource estimates.
·	Forecasted commodity prices and exchange rates.
·	Planned mine production schedule.
·	Estimated mining and process recovery rates.
·	Expectations as to mining dilution.
·	Capability to mine in areas earlier exploited.
·	Timing and amount of projected future production.
·	Sustaining costs and proposed operating costs.
·	Assumptions as to closure costs and closure requirements.
·	Assumptions as to environmental, permitting and social risks.

Additional risks to the forward-looking information include the following:

·	Variations in the costs of production from what was assumed.
·	Unrecognized environmental risks.
·	Unexpected reclamation expenses.
·	Unanticipated variations in the quantity of mineralized material, grade or recovery rates.
·	Accidents, labour disputes and other risks of the mining industry.
·	Geotechnical or hydrogeological considerations during mining differing from what was assumed.
·	Failure of mining methods to operate as anticipated.
·	Failure of plant, equipment or processes to operate as anticipated.

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·	Changes to assumptions on the availability of electrical power or the power rates used in the operating cost estimates and financial analysis.
·	Ability to maintain the social license to operate.
·	Modifications to interest rates.
·	Changes to tax rates.

22.2 Methodologies Used

The Project has been evaluated using a discounted cashflow (DCF) analysis based on a 5% discount rate. Cash inflows consist of annual revenue projections. Cash outflows consist of capital expenditures, including pre-production costs, operating costs, taxes, royalties, and streaming costs. These are subtracted from the inflows to arrive at the annual cash flow projections. Cash flows are taken to occur at the midpoint of each period. It must be noted that tax calculations involve complex variables that can only be accurately determined during operations and, as such, the actual post-tax results may differ from those estimated. A sensitivity analysis was performed to assess the impact of variations in metals price, discount rate, head grade, recovery, total operating cost, and total capital costs.

The capital and operating cost estimates developed specifically for this project are presented in Section 21 in Q3 2025 USD. The exchange rate used in the economic model is 1.35 CAD/USD. The economic analysis has been performed on a constant dollar basis with no inflation.

22.3 Financial Model Parameters

22.3.1 Assumptions

The base case metal price estimates used in the economic study were US$3,100/oz for gold and US$35.50 /oz for silver. These metal prices were determined using consensus expert projections and recent economic studies as described in Section 19. The forecasts used are meant to reflect the average metals price expectation over the life of the Project. No price inflation or escalation factors were considered. As a result, there is a possibility that the forecast may differ and that the price of the commodity may change.

The economic analysis also used the following assumptions:

·	Construction and commissioning period will be two years.
·	The total mine life is 9.4 years.
·	Cost estimates are in constant Q3 2025 USD with no inflation or escalation factors considered.
·	Results are based on 100% ownership with revenue from gold doré production, with a combination of 2.0% and 1.0% net smelter return (NSR) royalties applied to portions of the mineral reserve following royalty buybacks totalling US$5.2 million.

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·	A silver stream agreement on 25% of payable silver produced following a stream buyback totalling US$22.5 million at a contract price of US$7.50/oz Ag factored for inflation starting in the third year of production. The revenue lost to this stream is represented as a cost in financial modelling.
·	Capital costs are funded with 100% equity (no financing assumed).
·	All cash flows are discounted to the start of the construction period using a mid-period discounting convention.
·	All metal products will be sold in the same year they are produced.
·	Project revenue will be derived from the sale of gold doré.
·	Currently, there are no contractual refining arrangements.
·	Currency exchange rate 0.74 US$ - 1 C$.

22.3.2 Taxes

The Project has been evaluated on a post-tax basis to provide an approximate value of the potential economics. The tax model was compiled by Ausenco and reviewed by V.M. Denike Inc. as retained by First Mining. All tax calculations are based on the tax regime as of the date of this technical report. At the effective date of this report, the Project is assumed to be subject to:

·	the Canadian federal corporate income tax of 15%;
·	the Ontario provincial corporate income tax of 10%; and
·	the Ontario Capital Minimum Tax of 2.7%.

The tax calculations are based on the following assumptions:

·	The tax attribute opening balances are based on information provided by First Mining and include project-related sunk costs and pre-permit expenses.
·	Costs relating to NSR royalties and Silver Stream are included as a deduction for federal and provincial income tax calculations.
·	Actual taxes payable will be impacted by corporate activities and current tax benefits which are excluded from consideration.

The culmination of these taxes results in an estimated total tax of US$1,452 million over the LOM.

22.3.3 Working Capital

An estimate of working capital has been incorporated into the economic analysis based on the following assumptions.

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Table 22‑1: Estimate of Working Capital

Description	Time

| Inventories | 30 days |

| Accounts Payable | 30 days |

22.3.4 Closure Costs and Salvage Value

Closure costs and salvage value are applied after the end of the life-of-mine. Closure costs were estimated to be US$40.5 million, excluding closure of the process plant, and salvage value was estimated to be US$4.1 million, which was factored as 10% of the mining equipment initial direct capital costs.

22.3.5 Royalties

Based on the agreements in place as of the date of this technical report, NSR royalties of 2.0% and 1.0% on metals recovered from distinct portions of the deposit are used for the economic analysis following the application of royalty buybacks totaling US$5.2 million.

22.3.6 Off-Site Costs

The following off-site costs and sale terms used for the economic analysis were determined using industry benchmarks and standard industry practices with input from First Mining.

Table 22‑2: Smelter Term Summary

Description	Units	Value

| Au Payability in Doré | % | 99.9 |

| Ag Payability in Doré | % | 95.0 |

| Au Refining & Transportation | US$/oz Au | 3.70 |

| Ag Refining & Transportation | US$/oz Ag | 0.37 |

22.4 Economic Analysis

The economic analysis was performed using a 5% discount rate. The pre-tax net present value discounted at 5% (NPV5%) is US$3,228 million, the internal rate of return (IRR) is 53.8%, and the payback period is 1.4 years from the start of commercial production. On a post-tax basis, the NPV5% is US$2,150 million, the IRR is 40.9%, and the payback period is 1.8 years from the start of commercial production. A summary of project economics is tabulated in Table 22‑3. The economic analysis was performed on an annual cashflow basis; the cashflow output is shown in Table 22‑4 and cashflow is represented graphically in Figure 22‑1 on a post-tax basis.

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Figure 22‑1: Undiscounted, Unlevered, Free Cash Flow – Post Tax Basis

Source: Ausenco, 2025

Table 22‑3: Economic Analysis Summary

Item	Units	LOM	Years 1 to 5

| Gold Price | US$/oz | 3,100 | 3,100 |

| Silver Price | US$/oz | 35.50 | 35.50 |

| Foreign Exchange Rate | CAD:USD | 0.74 | 0.74 |

| Total Tonnes Processed | Mt | 102.0 | 53.6 |

| Total Tonnes Waste Mined | Mt | 309.5 | 217.0 |

| Mill Feed Grade – Au | g/t | 0.94 | 1.09 |

| Mill Feed Grade – Ag | g/t | 4.9 | 5.7 |

| Mine Life | Years | 9.4 | 5.0 |

| Mill Throughput | t/d | 30,000 | 30,000 |

| Average Strip Ratio | waste:ore | 3:1 | 3.2:1 |

| Average Recovery Rate – Au | % | 86.0 | 86.7 |

| Average Recovery Rate – Ag | % | 86.2 | 87.1 |

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Item	Units	LOM	Years 1 to 5

| Total Production – Au | koz | 2,648 | 1,626 |

| Total Production – Ag | koz | 13,842 | 8,518 |

| Average Annual Production – Au | koz/a | 281 | 325 |

| Average Annual Production – Ag | koz/a | 1,468 | 1,704 |

| Total Payable Metal – Au | koz | 2,646 | 1,624 |

| Total Payable Metal – Ag | koz | 13,150 | 8,092 |

| Total Revenue – LOM | U$M | 8,668 | 5,323 |

| Average Annual Revenue | US$M/a | 920 | 1,065 |

| Total EBITDA – LOM | US$M | 6,079 | 3,830 |

| Average Annual EBITDA | US$M/a | 645 | 766 |

| Total Operating Costs – LOM | US$M | 2,361 | 1,352 |

| Average Annual Operating Cost | US$M/a | 251 | 270 |

| Mining Cost | US$/t mined | 2.57 | 2.41 |

| Mining Cost | US$/t milled | 9.87 | 11.93 |

| Processing Cost | US$/t milled | 10.72 | 10.74 |

| G&A and Site Services Cost | US$/t milled | 2.56 | 2.58 |

| Total Operating Cost | US$/t milled | 23.15 | 25.26 |

| Total Cash Cost^1^ | US$/oz Au | 802.4 | 742.4 |

| All-in Sustaining Cost^2^ | US$/oz Au | 937.9 | 876.2 |

| Initial Capital Cost | US$M | 1,104 | - |

| Sustaining Capital Cost | US$M | 322 | 217 |

| Closure Cost | US$M | 40 | - |

| Salvage Value | US$M | 4 | - |

| NPV5% | US$M | 3,228 | 2,150 |

| IRR | % | 53.8 | 40.9 |

| Payback Period | Years | 1.4 | 1.8 |

| Undiscounted Cash Flow | US$M | 4,588 | 3,137 |

| NPV5%: Initial Capital Cost | NPV:Capex | 2.9 | 1.9 |

Note:

1.	Cash costs consist of mining costs, processing costs, mine-level G&A and refining charges, royalties, and streaming costs

| 2. | AISC includes cash costs plus sustaining capital, closure cost and salvage value |

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Table 22‑4: Cashflow Statement on an Annualized Basis

Macro Assumptions	Units	Total/ Avg	-2	-1	1	2	3	4	5	6	7	8	9	10	11

| Gold Price | US$/oz | 3,100 | -- | -- | 3,100 | 3,100 | 3,100 | 3,100 | 3,100 | 3,100 | 3,100 | 3,100 | 3,100 | 3,100 | 3,100 |

| Silver Price | US$/oz | 36 | -- | -- | 36 | 36 | 36 | 36 | 36 | 36 | 36 | 36 | 36 | 36 | 36 |

| Revenue | US$M | 8,668 | -- | -- | 1,103 | 1,222 | 1,027 | 997 | 974 | 859 | 988 | 994 | 354 | 151 | -- |

| Off-Site Costs | US$M | (15) | -- | -- | (2) | (2) | (2) | (2) | (1) | (1) | (2) | (2) | (1) | (0) | -- |

| Royalties | US$M | (122) | -- | -- | (21) | (18) | (14) | (11) | (12) | (10) | (15) | (16) | (5) | (2) | -- |

| Silver Stream Cost | US$M | (91) | -- | -- | (8) | (14) | (18) | (11) | (6) | (7) | (10) | (12) | (4) | (2) | -- |

| Operating Cost | US$M | (2,361) | -- | -- | (261) | (278) | (291) | (269) | (254) | (271) | (266) | (233) | (168) | (70) | -- |

| EBITDA | US$M | 6,079 | -- | -- | 812 | 910 | 702 | 705 | 701 | 569 | 695 | 730 | 177 | 77 | -- |

| Initial Capex | US$M | (1,104) | (397) | (706) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| Sustaining Capex | US$M | (322) | -- | -- | (38) | (36) | (28) | (51) | (64) | (54) | (34) | (17) | -- | -- | -- |

| Royalty Buyback | US$M | (5) | -- | (5) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| Stream Buyback | US$M | (23) | -- | (23) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| Closure Capex | US$M | (40) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | (40) |

| Salvage Credit | US$M | 4 | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | 4 | -- |

| Pre-Tax Unlevered Free Cash Flow | US$M | 4,589 | (397) | (734) | 774 | 874 | 674 | 654 | 636 | 515 | 662 | 714 | 177 | 81 | (40) |

| Income Tax, Government Royalties and Profit Sharing | US$M | (1,452) | -- | -- | (162) | (203) | (152) | (169) | (177) | (142) | (190) | (206) | (39) | (11) | -- |

| Post-Tax Unlevered Free Cash Flow | US$M | 3,137 | (397) | (734) | 612 | 670 | 522 | 484 | 459 | 373 | 472 | 508 | 138 | 70 | (40) |

| Waste Mined Total | kt | 309,482 | -- | 17,709 | 41,859 | 32,557 | 25,175 | 47,435 | 52,218 | 51,803 | 31,676 | 9,049 | -- | -- | -- |

| Mineralized Material Mined Total | kt | 102,015 | -- | 2,295 | 12,160 | 18,672 | 24,825 | 7,565 | 2,782 | 8,197 | 13,631 | 11,889 | -- | -- | -- |

| Total Mill Feed | kt | 102,015 | -- | -- | 9,752 | 10,950 | 10,950 | 10,950 | 10,950 | 10,950 | 10,950 | 10,950 | 10,950 | 4,664 | -- |

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Macro Assumptions	Units	Total/ Avg	-2	-1	1	2	3	4	5	6	7	8	9	10	11

| Mill Feed - Au Grade | g/t | 0.94 | -- | -- | 1.24 | 1.22 | 1.00 | 1.00 | 1.00 | 0.87 | 1.00 | 0.99 | 0.39 | 0.39 | -- |

| Mill Feed - Ag Grade | g/t | 4.90 | -- | -- | 4.39 | 6.81 | 8.37 | 5.50 | 3.19 | 3.85 | 4.91 | 6.25 | 2.00 | 2.00 | -- |

| Total Metal Content - Au | koz | 3,079 | -- | -- | 390 | 429 | 352 | 352 | 352 | 308 | 352 | 349 | 137 | 58 | -- |

| Total Metal Content - Ag | koz | 16,065 | -- | -- | 1,375 | 2,399 | 2,947 | 1,935 | 1,124 | 1,354 | 1,730 | 2,199 | 703 | 299 | -- |

| Average Recovery - Au | % | 86.0% | -- | -- | 88.1% | 86.7% | 86.0% | 86.3% | 86.4% | 86.2% | 86.0% | 85.9% | 79.0% | 79.0% | -- |

| Average Recovery - Ag | % | 86.2% | -- | -- | 84.7% | 88.1% | 89.7% | 86.4% | 82.3% | 83.7% | 85.6% | 87.4% | 78.6% | 78.6% | -- |

| Total Metal Produced - Au | koz | 2,648 | -- | -- | 343.5 | 371.6 | 302.7 | 303.8 | 304.3 | 265.0 | 302.8 | 299.9 | 108.4 | 46.2 | -- |

| Total Metal Produced - Ag | koz | 13,842 | -- | -- | 1,164.8 | 2,113.0 | 2,642.5 | 1,672.6 | 924.8 | 1,132.9 | 1,480.3 | 1,922.8 | 552.6 | 235.3 | -- |

| Total Payable Gold | koz | 2,646 | -- | -- | 343.2 | 371.3 | 302.4 | 303.5 | 304.0 | 264.8 | 302.5 | 299.6 | 108.3 | 46.1 | -- |

| Total Payable Silver | koz | 13,150 | -- | -- | 1,106.6 | 2,007.4 | 2,510.3 | 1,588.9 | 878.5 | 1,076.3 | 1,406.3 | 1,826.6 | 524.9 | 223.6 | -- |

| Total Operating Costs | US$M | (2,361) | -- | -- | (261) | (278) | (291) | (269) | (254) | (271) | (266) | (233) | (168) | (70) | -- |

| Mine Operating Costs | US$M | (1,006) | -- | -- | (127) | (133) | (146) | (124) | (109) | (126) | (121) | (88) | (24) | (8) | -- |

| Mill Processing Costs | US$M | (833) | -- | -- | (79) | (90) | (90) | (90) | (90) | (90) | (90) | (90) | (90) | (38) | -- |

| G&A Costs Total | US$M | (261) | -- | -- | (28) | (28) | (28) | (28) | (28) | (28) | (28) | (28) | (28) | (12) | -- |

| Total Unit Operating Costs | US$/t Milled | (23.1) | -- | -- | (26.7) | (25.4) | (26.6) | (24.5) | (23.2) | (24.8) | (24.3) | (21.3) | (15.4) | (15.0) | -- |

| Total Offsite Charges | US$M | (15) | -- | -- | (2) | (2) | (2) | (2) | (1) | (1) | (2) | (2) | (1) | (0) | -- |

| Au Transport & Refining Charges | US$M | (9.8) | -- | -- | (1.3) | (1.4) | (1.1) | (1.1) | (1.1) | (1.0) | (1.1) | (1.1) | (0.4) | (0.2) | -- |

| Ag Transport & Refining Charges | US$M | (4.9) | -- | -- | (0.4) | (0.7) | (0.9) | (0.6) | (0.3) | (0.4) | (0.5) | (0.7) | (0.2) | (0.1) | -- |

| NSR Royalties | US$M | (122) | -- | -- | (21) | (18) | (14) | (11) | (12) | (10) | (15) | (16) | (5) | (2) | -- |

| Silver Stream Cost | US$M | (91) | -- | -- | (8) | (14) | (18) | (11) | (6) | (7) | (10) | (12) | (4) | (2) | -- |

| C1 Cash Cost^1^ | US$/oz Au | 802 | -- | -- | 733 | 650 | 777 | 778 | 796 | 951 | 801 | 662 | 1,466 | 1,427 | -- |

| C3 Cash Cost^2^ | US$/oz Au | 938 | -- | -- | 844 | 747 | 870 | 946 | 1,006 | 1,156 | 912 | 717 | 1,466 | 1,339 | -- |

| Total Initial Capital | US$M | (1,104) | (397) | (706) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 1000 - Mining | US$M | (302) | (115) | (187) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

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Macro Assumptions	Units	Total/ Avg	-2	-1	1	2	3	4	5	6	7	8	9	10	11

| 2000 - Process Plant | US$M | (40) | (16) | (24) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 3000 - Additional Process Facilities | US$M | (349) | (122) | (227) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 4000 - On Site Infrastructure | US$M | (75) | (26) | (49) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 5000 - Off Site Infrastructure | US$M | (47) | (16) | (31) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 6000 - Indirects | US$M | (69) | (24) | (45) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 7000 - Project Delivery | US$M | (70) | (25) | (46) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 8000 - Owner's Costs | US$M | (24) | (9) | (16) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 9000 - Provisions | US$M | (128) | (45) | (83) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| Total Sustaining Capital | US$M | (322) | -- | -- | (38) | (36) | (28) | (51) | (64) | (54) | (34) | (17) | -- | -- | -- |

| 1000 - Mining | US$M | (304) | -- | -- | (34) | (32) | (24) | (47) | (60) | (54) | (34) | (17) | -- | -- | -- |

| 2000 - Process Plant | US$M | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 3000 - Additional Process Facilities | US$M | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 4000 - On Site Infrastructure | US$M | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 5000 - Off Site Infrastructure | US$M | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 6000 - Indirects | US$M | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 7000 - Project Delivery | US$M | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| 8000 - Owner's Costs | US$M | (19) | -- | -- | (4) | (4) | (4) | (4) | (4) | -- | -- | -- | -- | -- | -- |

| 9000 - Provisions | US$M | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- |

| Closure Cost and Salvage Credit | US$M | (36) | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | -- | 4 | (40) |

| Total Capital Expenditures Including Salvage Value | US$M | (1,462) | (397) | (706) | (38) | (36) | (28) | (51) | (64) | (54) | (34) | (17) | -- | 4 | (40) |

Note:

1.	Cash Costs consist of mining costs, processing costs, mine-level G&A, offsite charges, royalties, and streaming costs, less by-product credits.

| 2. | All-In Sustaining Costs includes cash costs plus sustaining capital, closure costs, and salvage credits |

| 3. | Dollar figures in Real 2025 US$M unless otherwise noted |

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22.5 Sensitivity Analysis

A sensitivity analysis was conducted on the base case pre-tax and post-tax NPV and IRR of the Project, using the following variables: metal prices, discount rate, total operating cost, initial capital cost, recovery, head grade, and exchange rate. Table 22‑5 shows the post-tax sensitivity analysis results; pre-tax sensitivity results are shown in Table 22‑6. Sensitivity to gold price is shown in Table 22‑7.

As shown in Figure 22‑2, the sensitivity analysis revealed that the Project is most sensitive to changes in commodity price, and exchange rate, and less sensitive to total operating cost and initial capital cost.

Figure 22‑2: Sensitivity Analysis

Source: Ausenco, 2025

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Table 22‑5: Post-tax Sensitivity

Post-Tax Sensitivity to Metal Price

| Post-Tax NPV5% (USM) Sensitivity to Discount Rate | | | | | | | | | | | |

| Discount Rate | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $1,821 | $2,364 | $2,907 | $3,450 | 3,994 | 1.0% | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| | $1,534 | $2,016 | $2,499 | $2,982 | 3,465 | 3.0% | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| | $1,289 | $1,720 | $2,150 | $2,582 | 3,013 | 5.0% | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| | $986 | $1,352 | $1,719 | $2,085 | 2,452 | 8.0% | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| | $819 | $1,149 | $1,480 | $1,810 | 2,141 | 10.0% | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| Total OPEX | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $1,528 | $1,958 | $2,389 | $2,820 | 3,252 | (20.0%) | 32.1% | 38.1% | 43.8% | 49.2% | 54.5% |

| | $1,409 | $1,839 | $2,270 | $2,701 | 3,132 | (10.0%) | 30.4% | 36.6% | 42.4% | 47.9% | 53.1% |

| | $1,289 | $1,720 | $2,150 | $2,582 | 3,013 | -- | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| | $1,169 | $1,601 | $2,031 | $2,462 | 2,893 | 10.0% | 26.9% | 33.4% | 39.4% | 45.0% | 50.4% |

| | $1,049 | $1,481 | $1,912 | $2,343 | 2,774 | 20.0% | 25.1% | 31.7% | 37.8% | 43.6% | 49.1% |

| Post-Tax NPV5% (USM) Sensitivity to Initial CAPEX | | | | | | | | | | | |

| Initial CAPEX | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $1,444 | $1,875 | $2,307 | $2,738 | 3,170 | (20.0%) | 36.4% | 43.6% | 50.4% | 56.9% | 63.0% |

| | $1,366 | $1,797 | $2,229 | $2,660 | 3,092 | (10.0%) | 32.2% | 38.9% | 45.2% | 51.2% | 56.9% |

| | $1,288 | $1,719 | $2,150 | $2,582 | 3,014 | -- | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| | $1,210 | $1,641 | $2,072 | $2,504 | 2,936 | 10.0% | 25.7% | 31.6% | 37.2% | 42.4% | 47.5% |

| | $1,130 | $1,564 | $1,994 | $2,426 | 2,858 | 20.0% | 23.1% | 28.8% | 34.0% | 39.0% | 43.7% |

| Post-Tax NPV5% (USM) Sensitivity to Mill Recovery | | | | | | | | | | | |

| Mill Recovery | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $595 | $941 | $1,288 | $1,633 | 1,977 | (20.0%) | 17.3% | 23.2% | 28.7% | 33.7% | 38.5% |

| | $942 | $1,332 | $1,719 | $2,107 | 2,495 | (10.0%) | 23.2% | 29.3% | 35.0% | 40.3% | 45.4% |

| | $1,289 | $1,720 | $2,150 | $2,582 | 3,013 | -- | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| | $1,634 | $2,108 | $2,582 | $3,057 | 3,531 | 10.0% | 33.8% | 40.3% | 46.5% | 52.3% | 57.9% |

| | $1,833 | $2,332 | $2,831 | $3,330 | 3,829 | 20.0% | 36.3% | 42.9% | 49.2% | 55.1% | 60.8% |

| Post-Tax NPV5% (USM) Sensitivity to Head Grade | | | | | | | | | | | |

| Head Grade | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $595 | $941 | $1,288 | $1,633 | 1,977 | (20.0%) | 17.3% | 23.2% | 28.7% | 33.7% | 38.5% |

| | $942 | $1,332 | $1,719 | $2,107 | 2,495 | (10.0%) | 23.2% | 29.3% | 35.0% | 40.3% | 45.4% |

| | $1,289 | $1,720 | $2,150 | $2,582 | 3,013 | -- | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| | $1,634 | $2,108 | $2,582 | $3,057 | 3,531 | 10.0% | 33.8% | 40.3% | 46.5% | 52.3% | 57.9% |

| | $1,979 | $2,497 | $3,014 | $3,531 | 4,049 | 20.0% | 38.6% | 45.4% | 51.8% | 57.9% | 63.8% |

| Post-Tax NPV5% (USM) Sensitivity to Exchange Rate | | | | | | | | | | | |

| Exchange Rate | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $2,154 | $2,693 | $3,232 | $3,771 | 4,310 | (20.0%) | 40.9% | 47.8% | 54.4% | 60.7% | 66.7% |

| | $1,673 | $2,152 | $2,631 | $3,110 | 3,589 | (10.0%) | 34.3% | 40.9% | 47.1% | 53.0% | 58.6% |

| | $1,289 | $1,720 | $2,150 | $2,582 | 3,013 | -- | 28.7% | 35.0% | 40.8% | 46.5% | 51.8% |

| | $973 | $1,367 | $1,758 | $2,149 | 2,541 | 10.0% | 23.7% | 29.9% | 35.5% | 40.9% | 45.9% |

| | $709 | $1,070 | $1,431 | $1,789 | 2,148 | 20.0% | 19.3% | 25.3% | 30.8% | 36.0% | 40.8% |

All values are in US Dollars.

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Table 22‑6: Pre-tax Sensitivity

Pre-Tax Sensitivity to Metal Price

| Pre-Tax NPV5% (USM) Sensitivity to Discount Rate | | | | | | | | | | | |

| | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $2,685 | $3,478 | $4,272 | $5,065 | 5,858 | 1.0% | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | $2,299 | $3,004 | $3,709 | $4,414 | 5,119 | 3.0% | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | $1,970 | $2,600 | $3,228 | $3,858 | 4,487 | 5.0% | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | $1,562 | $2,097 | $2,633 | $3,168 | 3,703 | 8.0% | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | $1,337 | $1,820 | $2,303 | $2,785 | 3,268 | 10.0% | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $2,318 | $2,948 | $3,577 | $4,206 | 4,835 | (20.0%) | 42.3% | 50.1% | 57.6% | 64.6% | 71.4% |

| | $2,144 | $2,774 | $3,403 | $4,032 | 4,661 | (10.0%) | 40.1% | 48.2% | 55.7% | 62.9% | 69.7% |

| | $1,970 | $2,600 | $3,228 | $3,858 | 4,487 | -- | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | $1,796 | $2,425 | $3,055 | $3,684 | 4,313 | 10.0% | 35.7% | 44.1% | 51.9% | 59.2% | 66.3% |

| | $1,622 | $2,251 | $2,881 | $3,510 | 4,139 | 20.0% | 33.4% | 42.0% | 49.9% | 57.4% | 64.5% |

| Pre-Tax NPV5% (USM) Sensitivity to Initial CAPEX | | | | | | | | | | | |

| | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $2,179 | $2,808 | $3,438 | $4,067 | 4,696 | (20.0%) | 47.9% | 57.3% | 66.2% | 74.5% | 82.5% |

| | $2,075 | $2,704 | $3,333 | $3,963 | 4,592 | (10.0%) | 42.5% | 51.2% | 59.4% | 67.2% | 74.6% |

| | $1,970 | $2,600 | $3,228 | $3,858 | 4,487 | -- | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | $1,866 | $2,495 | $3,124 | $3,754 | 4,383 | 10.0% | 34.1% | 41.8% | 49.0% | 55.9% | 62.4% |

| | $1,761 | $2,391 | $3,020 | $3,649 | 4,279 | 20.0% | 30.8% | 38.1% | 44.9% | 51.4% | 57.6% |

| Pre-Tax NPV5% (USM) Sensitivity to Mill Recovery | | | | | | | | | | | |

| | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $962 | $1,465 | $1,969 | $2,472 | 2,975 | (20.0%) | 23.2% | 30.9% | 37.9% | 44.5% | 50.8% |

| | $1,466 | $2,032 | $2,599 | $3,165 | 3,731 | (10.0%) | 30.9% | 38.8% | 46.1% | 53.0% | 59.6% |

| | $1,970 | $2,600 | $3,228 | $3,858 | 4,487 | -- | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | $2,474 | $3,167 | $3,859 | $4,551 | 5,243 | 10.0% | 44.6% | 53.1% | 61.1% | 68.7% | 76.0% |

| | $2,765 | $3,493 | $4,222 | $4,950 | 5,678 | 20.0% | 47.7% | 56.3% | 64.5% | 72.2% | 79.6% |

| Pre-Tax NPV5% (USM) Sensitivity to Head Grade | | | | | | | | | | | |

| | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $962 | $1,465 | $1,969 | $2,472 | 2,975 | (20.0%) | 23.2% | 30.9% | 37.9% | 44.5% | 50.8% |

| | $1,466 | $2,032 | $2,599 | $3,165 | 3,731 | (10.0%) | 30.9% | 38.8% | 46.1% | 53.0% | 59.6% |

| | $1,970 | $2,600 | $3,228 | $3,858 | 4,487 | -- | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | $2,474 | $3,167 | $3,859 | $4,551 | 5,243 | 10.0% | 44.6% | 53.1% | 61.1% | 68.7% | 76.0% |

| | $2,979 | $3,734 | $4,489 | $5,244 | 5,999 | 20.0% | 50.8% | 59.7% | 68.0% | 76.0% | 83.6% |

| Pre-Tax NPV5% (USM) Sensitivity to Exchange Rate | | | | | | | | | | | |

| | (20.0%) | (10.0%) | -- | 10.0% | 20.0% | | (20.0%) | (10.0%) | -- | 10.0% | 20.0% |

| | $3,234 | $4,020 | $4,807 | $5,593 | 6,380 | (20.0%) | 53.9% | 62.9% | 71.4% | 79.5% | 87.3% |

| | $2,532 | $3,231 | $3,930 | $4,629 | 5,329 | (10.0%) | 45.3% | 53.8% | 61.9% | 69.5% | 76.8% |

| | $1,970 | $2,600 | $3,228 | $3,858 | 4,487 | -- | 37.9% | 46.1% | 53.8% | 61.1% | 68.0% |

| | $1,511 | $2,083 | $2,655 | $3,227 | 3,799 | 10.0% | 31.5% | 39.5% | 46.8% | 53.8% | 60.4% |

| | $1,128 | $1,652 | $2,177 | $2,701 | 3,226 | 20.0% | 25.8% | 33.6% | 40.7% | 47.4% | 53.8% |

All values are in US Dollars.

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Table 22‑7: Sensitivity to Gold Price

Parameter	Units	-20%	-10%	Base Case	+13%	Spot

| Gold Price | (US$/oz) | $2,450 | $2,800 | $3,100 | $3,500 | $4,200 |

| Silver Price | (US$/oz) | $35.50 | $35.50 | $35.50 | $35.50 | 51.00 |

| FX | C$:US$ | 0.74 | 0.74 | 0.74 | 0.74 | 0.71 |

| Pre-Tax NPV5% | US$M | $1,963 | $2,645 | $3,229 | $4,008 | $5,610 |

| Pre-Tax IRR | % | 37.8% | 46.7% | 53.8% | 62.8% | 82.1% |

| Post-Tax NPV5% | US$M | $1,284 | $1,750 | $2,150 | $2,684 | $3,784 |

| Post-Tax IRR | % | 28.6% | 35.4% | 40.9% | 47.8% | 62.6% |

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23 ADJACENT PROPERTIES

There are several mineral exploration properties adjacent to the Springpole property covering portions of the Birch-Uchi Greenstone Belt. The area has seen sporadic exploration over the past 90 years and hosts several past-producing mines (Figure 23‑1).

The Qualified Person has been unable to verify the information and the information is not necessarily indicative of the mineralization on the property that is the subject of the technical report.

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Figure 23‑1: Adjacent Properties Map

Source: First Mining, 2025. Note: Tenure information sourced from the Ontario Mining Lands Administration System (MLAS), November 2025

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24 OTHER RELEVANT DATA AND INFORMATION

This section is not relevant to this technical report.

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25 INTERPRETATION AND CONCLUSIONS

25.1 Introduction

The QPs note the following interpretations and conclusions in their respective areas of expertize, based on the review of data available for this Report.

25.2 Exploration

Regional exploration within the First Mining tenure has confirmed the presence of favourable lithologies, structural features, alteration patterns, and localized mineralization typically associated with gold in the region. Historical data, combined with recent geophysical and geochemical surveys, highlight multiple prospective targets for future exploration.

Based on current knowledge, the Birch-Uchi Greenstone Belt remains prospective for additional gold mineralization. Continued exploration, including detailed mapping, geochemical and geophysical surveys, and targeted drilling is recommended to further evaluate the regional mineral potential.

25.3 Metallurgical Testwork

Head assays confirm the presence of deleterious elements mercury and arsenic. Mercury is present at concentrations that warrant mercury control. Some areas of elevated copper and zinc concentrations are seen which will result in higher than average cyanide consumption.

·	Average abrasion index is 0.16 g which is considered low.
·	BWi ranges from 7.3 to 19.3 kWh/t which is considered very soft to hard, ore hardness is highly variable.
·	Calculated Axb ranges from 28-224 which is considered competent to soft, ore competence is highly variable.

Ore hardness (BWi) does not vary significantly per rock group, however ore competence (calculated Axb) varies widely by rock group, with average values of Axb for Basement, Intrusive and Volcanic of 35, 87 and 128 respectively. The Basement rock group is considered competent while Intrusive and Volcanic rock groups are soft. For optimal mill design and operation, the proportion of Basement rock group in the mill feed will need to be managed.

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Small improvements in residue grades and overall extractions were seen at a finer primary grind P80 size in the range of 75, 100, 125, and 150 µm. These improvements were not deemed substantial enough to warrant finer grinding. Target primary grind size P80 selected is 150 µm.

Variability test results show no apparent relationship between the zone or lithology of the variability composite samples and the flotation mass pull and gold and silver distribution to concentrate. The mass recovered to concentrate increases with an increase in sulphur feed grade. Mass and metal recovery to concentrate was highly variable.

A 1–2% increase in leach extraction of gold was seen when reducing the target regrind P80 size from 17 µm to 10 µm. However, the specific energy increases from 38.8 kWh/t to 86.6 kWh/t for the finer grind size, which is unlikely to be economic. A regrind target size of 17 microns was selected.

The leach variability tests in Program BL758 show that:

·	leach extraction of gold and silver from flotation concentrate and flotation tailings was highly variable;
·	concentrate leach is essentially complete after eight hours and tailings leach after 10 h. No relationship is apparent between the zone or lithology of the composite and the overall gold recovery expected;
·	the overall gold extraction varied with head grade, and the low grade samples performed slightly worse than their higher grade counterparts; and
·	there was no significant reduction in overall gold extraction in the samples with high base metal content or arsenic.

Average sodium cyanide consumption from the Production and Master Composite samples was 0.12 kg/t of tailings for the tailings leach circuit and 2.87 kg/t of concentrate for the concentrate leach circuit. Average lime consumption was 1.33 kg/t of tailings and 3.98 kg/t of concentrate.

The overall flotation and leach extraction for the production, master and variability composites ranged between 79.2 and 94.5% for gold over a range of calculated head grades from 0.47 to 6.02 g/t. The overall flotation and leach extraction for the production, master and variability composites ranged between 72.9 and 95.8% for silver over a range of calculated head grades from 0.92 to 29.86 g/t.

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Oxygen uptake testwork showed that the concentrate is a high consumer of oxygen at 4 kg/t of leach feed, while the tailings is a low consumer of oxygen at 0.07 k/t of leach feed.

Optimal rougher concentrate leach conditions were determined to be a density of 45% solids with a pre-aeration period of 8 hours. The optimal rougher tailings leach conditions were set at 53% solids. Leaching at a higher solids content for both the concentrate and tailings streams showed no adverse impact on leach extraction.

A Merrill Crowe circuit was selected for precious metal recovery from the concentrate leach solution because of the mineralized material’s high silver to gold ratio and variable flotation and leach extraction. Analysis of the pregnant leach solution produced from rougher concentrates shows no species of concern that would negatively impact precipitation efficiency.

Cyanide detoxification conditions for the master composites 3 and 4 at 45% solids were optimized to be a 5:1 ratio of SO2 to CNWAD, 25 mg/L copper and 30 minutes retention for rougher concentrates. For rougher tailings at 53% solids, conditions included a 5:1 ratio of SO2 to CNWAD and 60 minutes retention. Under these conditions, the process met the targets for cyanide detoxification.

Rougher concentrates generated from the Merrill-Crowe master composites were evaluated to determine thickening and rheological characteristics and pressure filtration parameters. MC3 achieved an underflow density of 52% utilizing 60 g/t of flocculant at a 0.5 t/m^2^/h loading rate in dynamic thickening testing. MC4 achieved an underflow density of 55% with with 60 g/t of flocculant at a 0.5 t/m^2^/h loading rate. Pressure filtration tests achieved a filter cake moisture content of 10.4% for MC3 and 15.8% for MC4.

The sheared viscosities of MC3 and MC4 rougher concentrates were measured using a Brookfield DV2T viscometer with a vane spindle. Slurry viscosity was 0.051 to 0.052 Pa.s at the shear rate of 130 s^-1^ which is acceptable for pumping applications. Viscosity was 0.53 to 0.64 Pa.s at the shear rate of 4.5 s^-1^ which is acceptable for mixing and screening applications.

Jenike and Johanson (J&J) performed flow property testing for crushed ore and tailings samples to support the design and calculations for the major bulk solids handling components of the front-end crushing and conveying system and the tailings handling system. Solids flow property test data collected identified there are minimal design risks for the major bulk solids handling systems for the project.

A flowsheet maintaining separate high sulphide and low sulphide leach tailings streams reduces the acid generation potential in the tailings disposal facilities.

25.4 Quality Assurance and Quality Control

The quality assurance/quality control (QA/QC) programs instituted by Gold Canyon and First Mining and conducted by the laboratories utilized for the assay work are of a standard generally consistent with current industry practice.

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During the 2022 QA/QC program, First Mining noticed that some of the standard reference materials (SRM) for silver were consistently reporting higher than the expected value, although still largely within the accepted 3 standard deviation threshold. This issue affected a total of 5,292 core samples over 45 workorders assayed between May 26 and November 16, 2022. To quantify the potential error that can be attributed to the possible overestimation of silver assays, the QP carried out an investigation of the possible affected assay results. Because most of the samples with possible high Ag bias are mostly low-grade, and only 204 assay values (0.51% of all assays used for the silver estimate) are greater than 5 g/t, the QP is of the opinion that the inclusion of the 2022 silver assay data will not materially affect the results of the mineral resource estimate.

The analysis for gold and silver confirms an acceptable degree of reproducibility of samples for gold and a very good degree of reproducibility of samples for silver.

There is no evidence of bias in either gold or silver as a function of grade and First Mining has its own QA/QC procedures for deciding which assay batches are acceptable or not and which samples need to be re-assayed because of failed QA/QC.

The drill hole database from 2003 through 2025 is of a standard acceptable for public reporting of Mineral Resources.

25.5 Mineral Resource Estimate

In the opinion of the QP, the resource evaluation reported herein is a reasonable representation of the global gold and silver resources found in the Project at the current level of sampling. The mineral resources in this Report were estimated in accordance with current Canadian Institute of Mining, Metallurgy and Petroleum (CIM) standards definitions, and guidelines, and reported using the 2014 CIM Definition Standards and, at a 0.20 g/t Au cut-off, include 191 Mt grading 0.78 g/t Au and 4.6 g/t Ag classified as Indicated mineral resources and 64 Mt grading 0.38 g/t Au and 3.1 g/t Ag classified as Inferred mineral resources.

The revised mineral resource estimate was based on a gold price of USD$2,450/oz and a silver price of USD$27.50/oz, both considered reasonable economic assumptions by the QP. To establish a reasonable prospect of economic extraction in an open pit context, the resources were defined within an optimized pit shell with pit walls set at 35 to 45° based on domain. Assumed metallurgical recoveries of 87.2% for gold and 85.5% for silver were used. Mining costs were estimated at CDN$2.30/t of total material, processing costs estimated at CDN$14.50/t, and general and administrative (G&A) costs estimated at CDN$0.90/t. A cut-off grade (COG) of 0.2 g/t Au was calculated and is considered to be an economically reasonable value corresponding to breakeven mining costs. Approximately 90% of the revenue for the proposed Project is derived from gold, with 10% derived from silver.

Industry standard mining, process design, construction methods and economic evaluation practices were used to assess the Project. In the QP’s opinion, there is adequate geological and other pertinent data available to generate a PFS.

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Based on current knowledge and assumptions, the results of this study show that the Project has positive economics and should be advanced to the next level of study, a Feasibility Study.

As with almost all mining ventures, there are a large number of risks and opportunities that can influence the outcome of the Project. Most of these risks and opportunities are based on a lack of scientific information (test results, drill results, etc.) or the lack of control over external drivers (metal prices, exchange rates, etc.).

Subsequent higher-level engineering studies are needed to further refine these risks and opportunities, identify new ones, and define mitigation or opportunity implementation plans.

While a significant amount of information is still required to do a complete assessment, at this point there do not appear to be any fatal flaws for the Project.

25.6 Mining Methods

The PFS mine plan is based on open pit mining. This pit will provide the open pit feed material necessary to maintain the process plant feed rate at 30,000 t/d while operational.

The Springpole pit will be a three-phase design which provides 102.0 Mt of mill feed grading 0.94 g/t Au, and 4.90 g/t Ag. Waste from this pit will total 309.5 Mt for a strip ratio of 3.0 (waste:mill feed).

The mill feed cut-off used is 0.25 g/t AuEQ, accounting for the contribution of both gold and silver. This is equivalent to a gold only cut-off of 0.27 g/t. During the mine operation material would be stockpiled to optimize the plant feed grade and defer lower-grade material until later in the mine schedule. Three grade bins are used for the stockpiles including: low grade (0.25 - 0.50 g/t AuEQ), medium grade (0.50 – 0.75 g/t AuEQ) and high grade (+0.75 g/t AuEQ).

The pits will be built on 12 m benches with safety berm placement each 24 m. Inter-ramp angles will vary from 22 to 54° in rock depending upon the wall orientation. Overburden uses a 30° inter-ramp angle with 12 m between berms. Minimum mining widths of 35 - 40 m were maintained in the design with preferred bench widths of 60 m or more. Ramps are at maximum 10% gradient and vary in width from 27.1 m (single lane width) to 35.5 m (double lane width). They have been designed for a 240-tonne haulage truck.

The main fleet will consist of up to four 251 mm rotary drills and two 140 mm drills, two 37 m^3^ electric hydraulic shovels and three 23 m^3^ front end loaders. The truck fleet will total 25 – 240-tonne trucks at the peak of mining. This is due to the long hauls from the pit to the CDF. The usual assortment of dozers, graders, small backhoes, and other support equipment is considered in the equipment cost estimate. A smaller front-end loader (13 m^3^) will be stationed at the primary crusher.

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Waste material from the pit will be stored in the CDF. NAG/non-metal leaching material will be used for the outer Embankments while PAG material will be co-disposed with process plant tailings. The majority of the NAG waste rock material from the open pit will be contained within the CDF (134.6 Mt), but a small portion of NAG material will be backfilled into Phase 2 near the end of the mine life. This reduces the overall haul length and helps in reclamation of the pit. An estimated total of 15.4 Mt will be backfilled into the pit.

The mine equipment fleet is anticipated to be financed to lower initial capital requirements. The LOM operating cost, during the production period, is estimated at C$3.47/t mined (US$2.57/t mined). This includes equipment financing of C$0.70/t mined (US$0.52/t mined). Some operating costs associated with the NAG/non-metal leaching waste rock material for the CDF have been categorized or allocated to sustaining capital as Capitalized Stripping. This item is estimated at C$0.66/t mined (US$0.49/t mined).

Pre-production stripping costs are estimated at C$104.7 M (US$77.5 M), including capitalized operating costs and equipment leasing costs during pre-production period. Other initial mine capital including equipment is C$92.2 M (US$ 68.3 M) and C$200.8 M (US$148.7 M) for additional earthworks, road construction, water management in and around the CDF Area, plus controlled dewatering of the open pit basin. Sustaining capital is estimated at C$293.7 M (US$ 217.5 M), including the aforementioned Capitalized Stripping, and equipment capital during the production period.

In addition to the open pit, two quarries will be established during the pre-production period and at closure. These quarries will be used to provide rock material for various mine infrastructure including haul roads, dikes, CDF Embankments, and to meet site fill requirements for other infrastructure. Approximately 46.2 Mt of material is planned to be excavated from the CDF quarry, while 26.2 Mt of material is planned from the fish habitat development area.

25.7 Recovery Methods

The selected flowsheet aligns with conventional practices in the industry. Comminution, flotation, precious metal extraction and recovery of precious metals, destruction of free cyanide and handling of tailings are achieved through conventional processes that are commonly used in the industry for similar projects with no significant elements of technological innovation. Previous studies, coupled with historical and new testwork results and financial evaluations, were used to develop the resulting flowsheet suitable for the blend of rock groups and the feed grades expected over the LOM.

The plant is designed for a throughput of 30 kt/d, with overall availability and utilization of 92%. The primary crusher circuit design is set at 75%, and the milling, flotation and leach circuit is set at 92% availability and utilization. The project has an estimated life of 10 years.

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The flowsheet includes a mercury retort and gas capture and scrubbing system for mercury recovery.

25.8 Infrastructure

25.8.1 Summary

Key Project infrastructure as envisaged in the PFS includes: open pit mine area including mine haul roads and ramps; dikes for hydraulic isolation of the mine pit following bay dewatering; site main access roads, administrative access roads and maintenance roads, site main gate and guard house; mine administration and dry building, permanent camp accommodation; process plant e-room; crushing area e-room; reagent storage building; gold room; plant workshop and warehouse; truck shop and warehouse, tire changing facility, truck wash building; fuel facility, fuel office complex; fresh water intake; 230 kV overland and 25 kV power distribution lines; fresh water intake pumping supply and treatment; CDF; waste water treatment plant and explosives magazine. The main access road will be a private extension of the existing Wenesaga Road for forestry services and has been constructed up to 15 km from the Project site.

25.8.2 Dikes

Two dikes will be constructed across Springpole Lake to isolate the open pit during operations (dewatering and mining). The dikes will be constructed by placing rockfill sourced from pre-production phase. Rockfill dumping will advance from the shoreline along the proposed centreline to construct a 28 m wide fill surface. The combined length of the two dikes will be about 1060 m. The maximum constructed dam height will be approximately 15 m.

The dikes will consist of a zoned embankment shell and a hydraulic barrier comprised of a slurry wall and a grout curtain. The central zone of the embankment will be constructed with a 25 mm minus rockfill to support installation of the slurry wall, while the outer zone of the embankment will be constructed with 100 mm minus rockfill. The dike embankment and slurry wall will isolate the open pit area from Springpole Lake and limit seepage. The grout curtain may be required to further reduce seepage through the bedrock and limit pore pressure at the downstream toe of the embankment.

The closure and reclamation objectives for the Project will include flooding the dewatered region to Springpole Lake near final water level, and once the water quality meets all regulatory requirements, lowering of the dikes in a controlled manner and reconnection of the reclaimed basin to Springpole Lake.

25.8.3 Co-Disposal Facility

The CDF will store approximately 101 Mt of tailings and 146 Mt of PAG/metal leaching waste rock generated over the mine life. The CDF perimeter embankment is primarily constructed using NAG/NML waste rock generated from open pit mining. A total of 81 Mm^3^ of fill material is required for construction of the CDF perimeter dams. The south cell dam will be constructed with 3H:1V upstream slopes and 2H:1V downstream slopes. The north cell starter dam will also be constructed with same slopes and subsequent raises will be completed as centreline raises.

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The CDF perimeter embankments will be constructed year around due to the significant annual raise of 6 m. The average perimeter dam height is approximately 75 m. Contact water captured at the perimeter seepage collection system will be pumped back into the CDF. Water collected in the south cell pond will be reclaimed to the plant/mill. Excess water will be pumped to the central water storage pond for monitoring, treatment, and discharge to environment as needed.

At closure, north cell will be graded and provided with erosion protection cover to continue directing runoff to south cell. NAG tailings and/or suitable soil and granular cover will be provided over the south cell PAG tailings to remove excess water and safely pass IDF through overflow spillway.

25.9 Capital Cost Estimate

The capital cost estimate was developed in Q4 2025 using existing project designs which were updated with recent 2025 mine plans supplied by AGP and geotechnical designs from WSP. The costing have been built up by:

·	Budgetary quotes and data from projects from internal databases for mechanical equipment
·	Preliminary layouts for architectural, civil and structural disciplines, priced from vendor quotes and historical data from relevant reference projects
·	Piping and electrical factored from mechanical equipment installation
·	Contributions from WSP and AGP for geotechnical and mining respectively have been incorporated into this estimate

25.10 Operating Cost Estimate

The operating cost estimate was developed in Q4 2025 using data from vendor quotations, projects, studies and previous operations from internal databases. The operating cost estimate is approximately ±25 to ±15% accurate. The estimate covers the mining, processing, maintenance, power and general and administrative activities. Section 21 includes a summary of the operating expenses.

The average mining cost is US$12.44/t. The average process plant operating cost is US$10.72/t processed, and the annual G&A cost is US$27.0 million.

25.11 Economic Analysis

The economic analysis was performed by assuming a 5% discount rate. Cash flows have been discounted to the start of construction, assuming that the project execution will be made and major project financing will be carried out at this time.

The pre-tax NPV discounted at 5% is US$3,228 million; the IRR is 53.8%, and payback period is 1.4 years. On a post-tax basis, the NPV discounted at 5% is US$2,150 million, the IRR is 40.9%, and the payback period is 1.8 years. Cumulative post-tax unlevered free cash flow totals US$3,137 million.

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The sensitivity analysis revealed that the Project’s NPV and IRR are most sensitive to changes in metal price and exchange rate.

25.12 Risks and Opportunities

25.12.1 Risks

25.12.1.1 Metallurgical Testing

The extent of high copper throughout the deposit and the effect on cyanide consumption and the associated impacts in cyanide detox needs to be quantified.

The Basement rock group is considered competent while Intrusive and Volcanic rock groups are soft. For optimal mill design and operation, the proportion of Basement rock group in the mill feed will need to be managed.

25.12.1.2 Mining Methods

·	Southwest Pit Wall Slope: Flatter slopes may be required due to poor rock quality. Additional geotechnical drilling to accurately estimate expected slopes could mitigate the risk.
·	Water inflows: Additional pumping requirements and higher than anticipated site discharge to environment. Additional hydrogeological drilling/modelling to better understand water inflows could mitigate the risk.

25.12.1.3 Recovery Methods

The proposed recovery methods for the Project are based on historical work and two major metallurgical test programs carried out recently on composite and variability samples. Comminution, flotation and cyanide leaching are selected methods which are well-developed with a low technology risk.

There is a risk that throughput and reagent consumptions may differ from predictions in this report if actual plant feed differs substantially from samples tested.

A potential risk facing the comminution circuit design is the fact that the hardest mill feed occurs in year 1, combined with significant variation in ore competency across the ore types. Without ore blending, the SAG mill will likely be constrained when processing Basement ore and underutilized when treating softer ore types.

Using the current JK Dropweight Parameter (Axb) value of 65.2 in the process design criteria will optimize SAG mill capacity over the LOM but will result in reduced throughput when processing hard ore, namely in year 1.

There is a risk that target grind sizes will not be met and that leach recovery will differ from predictions due to ore variability and the associated difficulty in effectively controlling the unit processes.

There is a throughput risk for the concentrate regrind, leach and metal recovery circuit if sulphur grades in the mill feed exceed the design value. Blending will be required to ensure that high sulphur grades which drive mass pull to concentrate are avoided.

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25.12.1.4 Infrastructure

25.12.1.4.1 Co-Disposal Facility

·	Timely availability of suitable quality NAG/NML mine rock to support perimeter dam raising needs to be confirmed and integrated into the mine plan during future design stages. Additional NAG non-metal leaching rock could likely be quarried from elsewhere (Section 18.5.9) at increased cost
·	Additional closure costs could be incurred for vegetation or other requirements not defined at the time of this report
·	If the planned oxygen barriers are insufficient to prevent oxidation, there would be water quality concerns. Long term collection and treatment of seepage could be required. This risk will need to be addressed with additional geochemistry and hydrogeological studies and modeling to confirm performance requirements and designs of the planned CDF oxygen barriers, advance designs and identify additional contingencies
·	Process water quality was a key factor influencing CDF water quality during operations and initial studies indicated that water treatment would be needed at the tailings discharge to manage water quality risks. Process water quality data was limited at the time the initial work was undertaken and will need to be further evaluated as part of metallurgical and geochemical programs. Water treatment needs may change, or costs may be higher than estimated based on currently available data. Water treatment engineering studies are also required
·	The bedrock quality (e.g., hydraulic conductivity) poorer than indicated by the current investigations would be addressed by additional grouting beyond the allowances in the current cost estimate or other mitigation measures that could increase costs. Additional site investigations along the CDF perimeter dams are planned to further characterize shallow bedrock hydraulic conductivity and thick overburden areas to better define the grouting requirements and the costs
·	Quantities of stripping and fill are based on currently available information. If these quantities are higher than currently estimated additional costs would be incurred. Additional site investigation and testing will refine volumes and extraction scheduling of available NAG rock from potential CDF quarry and fish compensation quarry improving the cost estimate
·	The CDF design concept is dependent, in part, on the ratio of NAG and PAG tailings. If the PAG tailings are a greater percentage than currently estimated, the CDF design concept may not be viable. Prediction of this ratio is primarily dependent on the estimates of ore quantities and characteristics, and the effectiveness of the mill process to remove sulphides from the NAG tailings stream
·	If quantities of PAG or metal leaching rock exceed the quantity that can be encapsulated in NAG tailings, the CDF concept may not be viable. Estimates of the quantity of PAG mine rock (and characteristics) from the mine model and quantities NAG tailings quantity are the primary uncertainties
·	A greater proportion of the PAG tailings or PAG/ML rock than currently estimated and can be accommodated within the current design would require adjusting the design concept to accommodate resulting in potential additional costs and/or scheduling delays. Additional ore and mine rock quantity estimates and characterization are required to address this issue

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·	Impact to seepage quality due to the presence of organics under the tailings and PAG mine rock in the CDF. Additional clearing, grubbing and stripping of organics from the entire footprint of the CDF would increase cost. Additional geochemical studies are required to determine if this is necessary or confirm that it is not necessary
·	The stability of the east side of the CDF is dependent, in part, on the long-term stability (through closure) of the adjacent pit wall. Long term stability of the pit wall needs to be demonstrated or adjustments made to the CDF or pit shell
·	The CDF rate of rise is about 6 m/yr which is high compared to other sites. This requires coordination of water levels, PAG and NAG rock placement, tailings, etc. and possibly winter construction. This could impact costs and/or production

25.12.1.4.2 Dikes

·	Timely availability of NAG non-metal leaching mine rock to support dike construction needs to be confirmed and integrated into the mine plan during future design stages. Additional NAG non-metal leaching rock could likely be quarried from elsewhere (Section 18.5.9) at increased cost
·	Quantities of stripping and dredging for dike construction are based on currently available information. These quantities may be higher than currently estimated. Additional site investigation and testing will refine volumes
·	The bedrock quality below the west dike may be poorer than indicated by the current investigations due to a potential low RQD zone in bedrock. Additional site investigations (boreholes and geophysical surveys) are planned to further evaluate the extent and properties of this sand zone between the pit area and the west dike. If confirmed, this could require additional grouting beyond the allowances in the current cost estimate or other mitigation measures that could increase costs
·	The tie-in between the slurry wall and bedrock may be challenging where bedrock profile changes significantly. Additional site investigations (geophysical surveys) along the dike alignment are planned to further investigate the bedrock
·	A potential local soft clay deposit could pose a challenge to dike slope stability potentially increasing cost. Additional site investigations (boreholes and geophysical surveys) are proposed along dike alignment and select cross sections to better characterize the overburden. If a weak clay deposit is confirmed, mitigation measures such as a toe berm or other alternatives may be considered to enhance slope stability

25.12.1.5 Environmental Risks

Environmental risks associated with the Project at the current stage may include the following:

·	Regulatory risk associated with the EIS/EA process, including delays, onerous conditions, or unfavourable resolution.
·	Regulatory risk associated with the permitting process for construction and operation of the project, including delays, onerous conditions.

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·	Technical risk, including data gaps which may be identified during construction and operations permitting activities, or modelling predictions which differ from observed outcomes.

Regulatory and external risks are being managed through ongoing discussions and engagement with regulatory agencies and Indigenous communities for the project to ensure that appropriate information is provided to advance Project activities. Technical risks are being managed through the development of collection of applicable monitoring and other data to support project permitting and anticipated construction and operations monitoring programs.

25.12.2 Opportunities

25.12.2.1 Metallurgical Testing

Review flotation concentrate mass recovery against sulphur grades from the latest mine plan. There may be an opportunity to optimize concentrate regrind power requirements. The review also needs to confirm required flotation tailings sulphur grades required for disposal.

25.12.2.2 Mining Methods

·	Southwest Pit Wall Slope: Steeper slopes which will help reduce waste movement. Additional geotechnical drilling to accurately estimate expected slopes would be required to assess potential opportunity
·	PAG/NAG characterization: Reduced PAG material could result in a smaller CDF height. Additional geochemical testwork to improve material characterization would be required to assess potential opportunity
·	PAG material stored in Pit backfill: Reduced size of CDF and potential improved mine closure situation. Geochemical testwork to mimic subaqueous disposal for permitting would be required to assess potential opportunity

25.12.2.3 Recovery Methods

The Basement rock group is considered competent while Intrusive and Volcanic rock groups are soft. There is an opportunity to reduce the size of the SAG mill by limiting the amount of Basement rock in the mill feed through blending.

25.12.2.4 Co-Disposal Facility

Sourcing construction rock from the CDF quarry creates capacity below current ground surface (i.e., the void left by the quarried rock) potentially lowering the height of the dams. This represents a potential cost saving. The construction rock sources can be further optimized as the design progresses

Low RQD mine rock could potentially be processed (crushing and/or screening) for use as transition and/or filter materials if the quality is sufficient and the design requirements met. This could save the cost of processing higher strength rock. The use of the low RQD NAG rock can be further considered during future design, however it is also possible that this potential beneficial use may not be fully evaluated until bulk samples are available (i.e., during operations)

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·	Migration of tailings into the PAG mine rock in the north cell would reduce the required height of the CDF. There are opportunities during the future design stages and CDF operation to assess potential volume gain and optimize the CDF dam heights

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26 RECOMMENDATIONS

26.1 Recommended Work Program

The recommended work program is intended to advance the Project by improving confidence in the Mineral Resource estimate and supporting future technical studies. The recommended scope of work and the associated estimated costs for the proposed program are summarized below.

26.2 Geology

26.2.1 Inferred Resource Upgrade

In various pit areas it was noted that with a short drill program, some of the material currently classed as Inferred mineral resources, could potentially be converted to Indicated mineral resources. This could result in material currently classified as waste being processed at the mill. The bulk of the Inferred mineral resources at the Project are in the southern part of the pit, and approximately 10,000 m of drilling would be required to potentially support the upgrade of the material in this area to the Indicated mineral resource category.

26.2.2 Density Measurements

To date there are 626 core samples that have undergone bulk density testing at the Project. While sufficient for a PFS study, additional bulk density testwork is recommended to further improve the density model. This could be completed on existing core kept in storage on site at the Project and additional forward drilling should it be completed.

26.2.3 Geology Estimated Budget

The following is the estimated budget for the proposed drilling program and other study work for the continued development of the Mineral Resource estimates at the Springpole Gold project. The estimated budget for these proposed programs is shown in Table 26‑1.

Table 26‑1: Estimated Budget of Proposed Work

Proposed Work	Unit Rate (US$)	Approximate Cost (US$)

| Inferred Material Upgrade | | |

| Diamond Drilling (~10,000 m) | $296/m | 2,960,000 |

| Bulk Density Measurement; 500 samples | $25.9/sample | 12,950 |

| Subtotal | | 2,972,950 |

| Contingency | | 297,295 |

| Total | | 3,270,245 |

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26.3 Exploration

Further regional exploration is warranted across the Project tenure to support the development of an exploration target pipeline. Recommended work programs include an induced polarization (IP) geophysical survey to delineate prospective structures, as well as field programs comprising detailed geological mapping, geochemical soil surveys, and channel sampling to refine targets and assess surface mineralization. Follow-up drilling is recommended for both regional exploration targets and resource extensional targets. All geological, geochemical, geophysical, and drilling data will be integrated to update exploration models and guide subsequent work.

Table 26‑2: Exploration Program

Proposed Work	Unit Rate (US$)	Approximate Cost (US$)

| Field Work Program | | $74,000 |

| Regional Target IP Geophysical Survey | | $222,000 |

| Exploration Drilling | 12,000m @ $296/metre | $3,552,000 |

| 3D Geological Modelling | | $37,000 |

| Sub Total | | $3,885,000 |

| Contingency | | $388,500 |

| Total | | $4,273,500 |

26.4 Metallurgical Testing

Additional work is recommended ahead of the Feasibility Study phase as follows:

·	Supplementary variability testing to address gaps in spatial coverage of the deposit, particularly in Zones B and C and areas representing the early years of operation. This is needed to provide increased confidence in the comminution circuit design through an increased number of samples as the deposit is highly variable.
·	Variability tests should be carried out at the conditions selected for design and considering the full flowsheet.
·	Determine flotation recovery targets for sulphide sulphur required to produce non-acid generating tailings.
·	Vendor testing of leached concentrate pressure and vacuum filtration as an alternative to CCD.
·	Improved ore characterization to enable prediction of power requirement by year (by rock group or RQD) and a more precise mass recovery model. (Recommended testing includes whole rock analysis and mineralogy on feed and concentrate characterization composites and variability samples.)
·	Track Tellurium assays and recovery in testwork as a critical mineral of interest.

Metallurgical testwork and management thereof as described above is expected to cost between US$500,000 to US$600,000.

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26.5 Mining Methods

26.5.1 Mine Geotechnical

AGP recommends implementing a drilling and logging program which includes 5 drill holes in the south, and a single land-based drill hole in the north.

The quantity of sampling material, in terms of number of locations and volume or weight, must be representative of the extent of the Sand-Zone. The minimum requirement will be determined and included in the site-specific Logging Manual prior to the drilling campaign being started, with the number and location of samples probably needing to be adjusted in-field as drilling progresses. Care should be taken during the design and preparation phase to improve the chances of acquiring sufficient material to update the characterization through laboratory testing.

If in-situ bulk density measurements indicate that the condition of the sand material is loose, the following context should be considered when planning the laboratory testing program:

·	During the pit excavation process, if the sands remain saturated and are subject to shear deformation, pore water pressure can be generated.
·	If the sand’s hydraulic conductivity is relatively low, then the pore water pressure cannot be dissipated, and the sand could show an undrained behaviour.
·	The strength parameter to be used in stability analyses in this context is the undrained shear strength (), which depends on the sand relative bulk dry density.

26.5.2 In-Situ Characterization

The borehole geophysical probing of the PVC-lined Sand-zone drill holes should include the following:

·	Acoustic Televiewer (ATV) with structural interpretation.
·	Full waveform sonic (FWS) processed for physical properties.
·	Gamma-gamma Density (GGD) with near and far field data.
·	Borehole Magnetic Resonance (BMR) processed for porosity and hydraulic conductivity.

Pre-existing ATV data close to (<100 m) the proposed drill hole SRK23-005 could be used for the interpretation of structures. If suitable ATV data are not available, however, then the hole should be drilled, logged, and probed using ATV, mechanical caliper and GGD tools.

Surface geophysics lines will be run (as part of a broader characterization program being done by First Mining) between the drill hole collars, on ice, to assist with inter-hole lateral interpretation of material boundaries detected down-hole.

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26.5.3 Laboratory Testing for Rock

The following testing is recommended for rock samples recovered at 30 m stations within the rock domains:

·	Unconfined Compressive Strength (UCS), Triaxial Strength, Brazilian Tensile Strength, calipered density, moisture content, and Shear Strength testing of selected joint pairs.
·	Mineralogy and geochemical characterization of selected hand-samples, which will include main mineral types and bulk geochemistry.
·	In-field and other strength testing requirements are detailed in the site-specific Rock Logging Manual.

26.5.4 Laboratory Testing for the Sand-zone Material

The following tests are recommended for the bulk samples obtained from the Sand-zone to improve the physical description and determination of the critical friction angle for slope design and stability:

·	Particle Size Distribution (PSD) test, and visual examination of particle shape (e.g. using a stereo microscope).
·	Mineralogy and geochemical characterization, including main mineral types and geochemical analysis using X-ray fluorescence (XRF) methods.
·	Triaxial consolidated undrained (CU) shear strength tests, including pore water-pressure measurements.

If the material’s undrained behaviour is needed for stability analysis, then the following tests are required:

·	Hydraulic conductivity test using a flexible wall permeameter. Physical characterization (PSD) of all logs in sand can be used for initial screening of hydraulic conductivity values.
·	Determination of minimum and maximum void ratio values from the BMR downhole data to support the determination of the sand’s relative bulk dry density.
·	Triaxial consolidated undrained (CU) shear strength tests, including pore water-pressure measurements in samples reconstituted at the in-situ relative bulk dry density, to support the determination of the undrained shear strength ().

The budget for the geotechnical program is expected to cost US$570,000

26.5.5 Mining Methods

Building on the knowledge and the work completed in the PFS, the following is recommended for advancing the open pit design work to a Feasibility level of study:

·	Blasting Study

o	Further evaluation of pattern sizes and powder factors is required to enhance production and productivity

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·	Equipment Costs and Fleet Selection

o	Update the equipment operating and capital costs from vendors
o	Examine alternative vendors for specific equipment
o	Optimize the size of haulage trucks

■	current fleet is 25 trucks of 240 tonne class
■	further study needed to determine most cost-effective size

o	Examine truck box configurations for tailings haulage to avoid carry back while maximizing carrying capacity of lower density tailings

·	Review if autonomous trucks are an opportunity

o	Further study recommended
o	Technical and social benefits concerns need to be examined

·	Ore Sampling Protocols need to be established

o	Definition of ore/waste contacts
o	Sample size selection
o	Determine if blasthole samples can supplement the proposed Reverse Circulation (RC) samples

·	Haul Road Design

o	Detailed design of north end haul road
o	Detailed layout of haul roads for pre-production period

·	Pit Electrification Optimization

o	Examine the placement of infrastructure to bring power into the pits for shovels, drills, and dewatering

These recommendations are typically included in the normal cost of open pit design and engineering; therefore, no additional budget is listed beyond that which is allocated for the Feasibility Study. For the Feasibility Study the mine engineering aspect is expected to cost $400,000 with the additional studies.

26.6 Infrastructure

26.6.1 Dikes and CDF

The following tasks are recommended to advance the dikes and CDF design to feasibility level:

·	Site investigation to support feasibility design of CDF and Dikes. This task is currently in progress and expected to be completed by Spring 2026.

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·	Additional metallurgical studies to generate representative tailings and process water for geochemical test programs and modelling for the Feasibility Study.
·	Geotechnical testwork on tailings: This will include range of geotechnical tests to characterize engineering properties of tailings to support the CDF feasibility design.
·	Geochemical programs including additional assessment of the metal leaching and acid rock drainage potential of tailings, mine rock and quarry to support the feasibility design.
·	Complete geotechnical and other field investigations – additional geotechnical data is required to inform foundation design for infrastructure including the Process Plant and ancillary facilities, including targeted boreholes, in-situ testing, and laboratory characterization of soils to define bearing capacity, settlement, bedrock depth, frost susceptibility, and other key ground conditions.
·	Additional hydrogeological and geochemical modelling to refine the oxygen barrier design.
·	Depositional modelling with a more detailed mine plan to understand the requirements to raise the structures at the required rates and coordinate tailings deposition, mine rock placement and water management interactions during operations.
·	Water treatment assessment to confirm the viability and performance of proposed treatment approaches for CDF water quality management.
·	Feasibility Study.

o	Advance the CDF design to feasibility level design.
o	Advance the East and West Dikes design to feasibility level design.

Table 26‑3: Recommended Work Program

Program Component	Description	Estimated Total Cost (US$M)

| Site Investigations (currently on-going) | Site investigation assessing foundation conditions at process plant and ancillary facilities footprint, CDF and Dikes. | 5 |

| Geotechnical testing on tailings, rock and soils | Testing to assess engineering properties of tailings | 0.2 |

| Geochemical testing on tailings, mine rock and quarry | Testing to assess geochemical characteristics of tailings, mine rock and quarry | 1.3 |

| Environmental Studies | As required to support design development and construction planning | 0.5 |

| Feasibility Study – Dikes | Advance Dikes design to feasibility design level. | 0.8 |

| Feasibility Study – CDF | Advance CDF design to feasibility level design. Scope includes water balance, water management, geochemical assessment, quarry design, and closure design. | 3.5 |

| Total | | 11.3 |

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26.7 Environmental

Building on the knowledge and the work completed in the final EIS/EA, First Mining will proceed into obtaining environmental permits for construction and operation of the Project upon successful resolution of the EIS/EA process. The available environmental data will support permit applications, but there are opportunities to add value to the Project generally as listed below:

·	Store and manage all project environmental data in an appropriate database.
·	Continue hydrometric and lake level monitoring on waterbodies around the Project area.
·	Continue water quality monitoring on waterbodies around the Project area.
·	Continue the collection of hydrogeological and hydrogeochemical data, with allocation for additional data collection for identified areas needed for monitoring during permitting.
·	Investigate the potential for relocation of the effluent discharge location into Birch Lake.
·	Continue the engagement of local Indigenous communities in the preparation of environmental permits and closure planning.

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Roberts, A.R., 2010. “Springpole Project Summary Report on Exploration Activities Fall 2009 through Winter 2010 for Gold Canyon Resources Inc.”

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Saunders, R., and McIntosh, A., 2009. “Petrographic Analysis of Selected Core Samples, Springpole Property, Ontario.”

Saunders, R., and McIntosh, A., 2010. “Petrographic Analysis of Selected Drill Core Samples, Springpole 2010 Winter Core Program.”

Slate Falls Nation, 2024. “Health, Socio-Economic, Indigenous Knowledge and Land Use Baseline Study.”

Springpole Gold Project	Page 398

| NI 43-101 Technical Report and Pre-Feasibility Study | December 1, 2025 |

Smith, G., 2008a. “Winter Drilling Reports 2006 and 2007, Springpole Gold Project, Red Lake Mining District, Ontario for Gold Canyon Resources Inc.”

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SRK Consulting (Canada) Inc., 2013b. “Structural Model of the Springpole Project.”

SRK Consulting (Canada) Inc., 2019. “Preliminary Economic Assessment Update for the Springpole Gold Project for First Mining Gold Corp.”

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Wood, 2022. “Springpole Gold Project, Preliminary Acid Base Accounting Analogue – Draft Memo for First Mining Gold Corp.”

WSP Canada Inc, 2004. “Environmental Impact Statement/Environmental Assessment Summary English Springpole Gold Project, ONS2104.”

Zabev, B., 2004. “Technical Report on the Springpole Lake Property, Red Lake Mining Division, NW Ontario for Gold Canyon Resources Inc.”

Springpole Gold Project	Page 399

| NI 43-101 Technical Report and Pre-Feasibility Study | December 1, 2025 |

Appendix A Springpole Gold Project Mineral Tenure Map

Springpole Gold Project	Page 400

| NI 43-101 Technical Report and Pre-Feasibility Study | December 1, 2025 |

Appendix B List of the Patents, Mining Leases and Mining Claims

Springpole Gold Project	Page 401

| NI 43-101 Technical Report and Pre-Feasibility Study | December 1, 2025 |

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim ID	Tenure Type	Status	Township ** / Area**	Owner	Anniversary Date	Area (Ha)	Claim Area

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary Date	Area (Ha)	Claim Area

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary Date	Area (Ha)	Claim Area

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 1: GOLD CANYON - MINING CLAIMS

Claim **** ID	Tenure ** Type**	Status	Township ** / Area**	Owner	Anniversary ** Date**	Area ** (Ha)**	Claim ** Area**

TABLE 2: GOLD CANYON RESOURCES - MINING LEASES

Lease ID	Claim ** ID(s)**	Owner	Township/ ** Area**	Area (ha)	Property

| 108953 | KRL562895 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 18 | Springpole |

| 108954 | KRL562896 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 7.7 | Springpole |

| 108955 | KRL562897 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 13 | Springpole |

| 108956 | KRL562898 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 12 | Springpole |

| 108957 | KRL562899 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 15.4 | Springpole |

| 108958 | KRL562900 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 17 | Springpole |

| 109846 | CLM 527 (KRL720373, KRL720374, KRL720375, KRL818859, KRL818714, KRL818715, KRL818855, KRL818858, KRL818893, KRL856298, KRL856301, KRL856302, KRL856305, KRL870237, KRL903559, KRL977224, KRL977247, KRL977248, KRL977249, KRL977250, KRL1185277, KRL1201986, KRL1201987, KRL1201988, KRL1201991, KRL1201992, KRL3004712, KRL3018701 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA, KEIGAT LAKE AREA, SEAGRAVE LAKE AREA | 708.4 | Springpole |

TABLE 3: GOLD CANYON RESOURCES - PATENTS

Claim	Owner	Township ** / Area**	Patent **** ID	Area ** (ha)**	Property

| KRL11229 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 7562 | 14 | Springpole |

| KRL11230 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 7563 | 18.2 | Springpole |

| KRL11231 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 7564 | 24.9 | Springpole |

| KRL12868 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 7565 | 18.8 | Springpole |

| KRL12869 | Gold Canyon Resources Inc. | CASUMMIT LAKE AREA | 7566 | 20.7 | Springpole |

| KRL12909 | Kenneth Gaarder | CASUMMIT LAKE AREA | 8876 | 26.9 | Springpole |

| KRL12903 | Neil Gaarder | CASUMMIT LAKE AREA | 8877 | 28.4 | Springpole |

| KRL12907 | Neil Gaarder | CASUMMIT LAKE AREA | 8878 | 17.3 | Springpole |

| KRL11233 | R&S Legacy | CASUMMIT LAKE AREA | 8879 | 13.7 | Springpole |

| KRL11234 | R&S Legacy | CASUMMIT LAKE AREA | 8880 | 23.7 | Springpole |

| KRL11235 | R&S Legacy | CASUMMIT LAKE AREA | 8881 | 19.8 | Springpole |

| KRL12896 | R&S Legacy | CASUMMIT LAKE AREA | 8882 | 15.6 | Springpole |

| KRL12897 | R&S Legacy | CASUMMIT LAKE AREA | 8883 | 22.1 | Springpole |

| KRL12898 | R&S Legacy | CASUMMIT LAKE AREA | 8884 | 17.7 | Springpole |

| KRL12899 | R&S Legacy | CASUMMIT LAKE AREA | 8885 | 16.5 | Springpole |

| KRL12900 | R&S Legacy | CASUMMIT LAKE AREA | 8886 | 18.7 | Springpole |

| KRL12901 | R&S Legacy | CASUMMIT LAKE AREA | 8887 | 19.9 | Springpole |

| KRL13043 | R&S Legacy | CASUMMIT LAKE AREA | 8888 | 14.9 | Springpole |

| KRL12905 | Everett Williams | CASUMMIT LAKE AREA | 8965 | 15.2 | Springpole |

| KRL12908 | Everett Williams | CASUMMIT LAKE AREA | 8966 | 21.9 | Springpole |

| KRL12904 | Everett Williams | CASUMMIT LAKE AREA | 8967 | 16.7 | Springpole |

| KRL12874 | Everett Williams | CASUMMIT LAKE AREA | 8968 | 23.7 | Springpole |

| KRL11236 | Douglas Hamblin | CASUMMIT LAKE AREA | 8970 | 17.4 | Springpole |

| KRL12872 | Douglas Hamblin | CASUMMIT LAKE AREA | 8971 | 24.6 | Springpole |

| KRL12867 | Walter Howard | CASUMMIT LAKE AREA | 8979 | 19.2 | Springpole |

| KRL12873 | Tim Howard | CASUMMIT LAKE AREA | 8980 | 25.2 | Springpole |

| KRL12871 | The Springpole Co. | CASUMMIT LAKE AREA | 48767 | 22.7 | Springpole |

| KRL12870 | The Springpole Co. | CASUMMIT LAKE AREA | 48768 | 10.6 | Springpole |

| KRL12902 | Lillian Hamblin | CASUMMIT LAKE AREA | 51239 | 24.6 | Springpole |

| KRL12906 | Lillian Hamblin | CASUMMIT LAKE AREA | 51240 | 16.3 | Springpole |