PRELIMINARY ECONOMIC ASSESSMENT FOR THE KIENA MINE … · I, Jorge Torrealba, P. Eng., employed with BBA Inc., do hereby consent to the filing of the Technical Report prepared for

NI 43-101 Technical Report . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

PRELIMINARY ECONOMIC ASSESSMENT FOR THE

KIENA MINE COMPLEX PROJECT

Val-d’Or, Québec, CANADA

. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Prepared for:

Wesdome Gold Mines Ltd.

EFFECTIVE DATE: March 31, 2020

SIGNATURE DATE: June 17, 2020

By qualified persons:

J Thomas Corkal, P. Eng. ............. BBA Inc.

Colin Hardie, P. Eng. .................... BBA Inc.

Luciano Piciacchia, P. Eng. ........... BBA Inc.

Pierre-Luc Richard, P. Geo. ........... BBA Inc.

Jorge Torrealba, P. Eng. ................ BBA Inc.

IMPORTANT NOTICE

This report was prepared as a National Instrument 43-101 Technical Report (the “Report”) for Wesdome

Gold Mines Ltd. (Wesdome) by BBA Inc. (BBA), 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 Wesdome, subject to the respective terms and conditions of its contracts with the

individual Report Authors. Except for the purposes legislated under Canadian provincial and territorial

securities law, any other uses of this Report by any third party is at that party’s sole risk. The

responsibility for this disclosure remains with Wesdome. The user of this document should ensure that

this is the most recent technical report for the Property as it is not valid if a new technical report has been

issued.

CAUTIONARY STATEMENT

This Preliminary Economic Assessment (PEA) is preliminary in nature and is based on numerous

assumptions and inferred mineral resources. Inferred mineral resources are considered too speculative

geologically to have economic considerations applied to them that would enable them to be categorized

as mineral reserves except as allowed for by Canadian Securities Administrators' National Instrument

43-101 in PEA studies. No mineral reserves have been estimated. There is no guarantee that Inferred

resources can be converted to Indicated or Measured resources and, as such, there is no guarantee that

the Project economics described herein will be achieved.

Cover photo credit: © 2016 Jean-Philippe Richard


NI 43-101 – Technical Report

Preliminary Economic Assessment – Kiena Mine Complex

JUNE 2020 BBA Document No.: 3767002-000000-40-ERA-0002-R00

DATE AND SIGNATURE PAGE

This report is effective as of the 31st day of March 2020.

“Signed and sealed original on file” June 17, 2020

J Thomas Corkal, P. Eng. BBA Inc.

Date


Colin Hardie, P. Eng. BBA Inc.

Date


Luciano Piciacchia, P. Eng. BBA Inc.

Date


Pierre-Luc Richard, P. Geo. BBA Inc.

Date


Jorge Torrealba, P. Eng., Ph.D. BBA Inc.

Date

2020 Robert-Bourassa Blvd. Suite 300 Montréal, QC H3A 2A5 T +1 514.866.2111 F +1 514.866.2116

bba.ca

To: Wesdome Gold Mines Ltd.

And to: Autorité des marchés financiers (Québec)

Alberta Securities Commission

British Columbia Securities Commission

Ontario Securities Commission

And to: Toronto Stock Exchange

CONSENT OF QUALIFIED PERSON - FILED BY SEDAR

I, J Thomas Corkal, P. Eng., employed with BBA Inc., do hereby consent to the filing of the

Technical Report prepared for Wesdome Gold Mines Ltd. titled “NI 43-101 Technical Report for

the Preliminary Economic Assessment for the Kiena Mine Complex Project”, dated June 17, 2020

and an effective date of March 31, 2020, with the securities regulatory authorities referred to

above that is being filed in support of a press release dated May 27, 2020 (the “Press Release”).

I, the undersigned, hereby confirm that I have read the Press Release and that it fairly and

accurately represents the information in the sections for which I am responsible for in the

abovementioned Technical Report.

Signed this 17th day of June 2020.

“Signed and sealed original on file”

J Thomas Corkal, P. Eng.

BBA Inc.

1010 Lorne Street, Unit 101, Sudbury (ON) P3C 4R9 T + 1 705.265.1119 F + 1 450.464.0901 bba.ca

CERTIFICATE OF QUALIFIED PERSON

J Thomas Corkal, P. Eng.

This certificate applies to the NI 43-101 Technical Report titled "Preliminary Economic Assessment for the

Kiena Mine Complex Project, Val-d’Or, Québec, Canada" (the Technical Report), prepared for Wesdome

Gold Mines Ltd. (Wesdome) issued on June 17, 2020 and effective as of March 31, 2020.

I, J Thomas Corkal, P. Eng., do hereby certify that:

I am a Senior Mining Engineer with BBA Inc. located at 1010 Lorne Street, Unit 101, Sudbury (ON) P3C 4R9.

I am a graduate of Queen’s University at Kingston in Ontario, Canada and have a Bachelor of Science Degree from the Faculty of Applied Science specializing in Mining Engineering received in 1979.

I am a member of the Professional Engineers of Ontario (license #9379504).

During my 40 plus years in mining, I have worked at mining operations in Technical and Operations roles including Engineer, Chief Engineer, Principle Engineer and supervisory positions. My experience is predominantly in underground bulk and narrow vein mines.

I have read the definition of “qualified person” set out in the NI 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101) and certify that, by reason of my education, affiliation with a professional association, and past relevant work experience, I fulfill the requirements to be a qualified person for the purposes of NI 43-101.

I am independent of the issuer applying all the tests in Section 1.5 of NI 43-101.

I am author and responsible for the preparation of Chapter 16, and I am also responsible for the relevant portions of chapters 1, 2, 3, 21, 25, 26 and 27 of the Technical Report.

I personally visited the property that is the subject to the Technical Report on October 2nd and 3rd of 2019.

I have had no prior involvement with the property that is the subject of the Technical Report.

I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101.

As at 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 the portions of the Technical Report for which I am responsible not misleading.



J Thomas Corkal, P. Eng. BBA Inc.


bba.ca








I, Colin Hardie, P. Eng., employed with BBA Inc., do hereby consent to the filing of the Technical

Report prepared for Wesdome Gold Mines Ltd. titled “NI 43-101 Technical Report for the

Preliminary Economic Assessment for the Kiena Mine Complex Project”, dated June 17, 2020








Colin Hardie, P. Eng.

BBA Inc.

2020 Robert-Bourassa Blvd. Suite 300 Montréal, QC H3A 2A5 T +1 514.866.2111 F +1 514.866.2116 bba.ca


Colin Hardie, P. Eng.

This certificate applies to the NI 43-101 Technical Report titled "Preliminary Economic Assessment for the Kiena Mine

Complex Project, Val-d’Or, Québec, Canada" (the Technical Report), prepared for Wesdome Gold Mines Ltd.

(Wesdome) issued on June 17, 2020 and effective as of March 31, 2020.

I, Colin Hardie, P. Eng., do hereby certify that:

I am the National Director, Mining and Metals Business Line with the firm BBA Inc. located at 2020 Robert-Bourassa Blvd., Suite 300, Montréal, Québec, H3A 2A5, Canada.

I graduated from the University of Toronto, Ontario Canada, in 1996 with a BASc in Geological and Mineral Engineering. In 1999, I graduated from McGill University of Montréal, Québec Canada, with an M. Eng. in Metallurgical Engineering and in 2008 obtained a Master of Business Administration (MBA) degree from the University of Montréal (HEC), Québec Canada.

I am a member in good standing of the Professional Engineers of Ontario (PEO: 90512500) and of the Canadian Institute of Mining, Metallurgy, and Petroleum (Member No. 140556). I have practiced my profession

continuously since my graduation.

I have been employed in mining operations, consulting engineering and applied metallurgical research for over 20 years. My relevant project experience includes metallurgical testwork analysis, flowsheet development, cost estimation and financial modeling. Since joining BBA in 2008, I have worked as a senior process engineer and/or lead study integrator for numerous North American iron ore, precious metal, industrial mineral, and base metal projects.

I have read the definition of “qualified person” set out in NI 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101) and certify that, by reason of my education, affiliation with a professional association, and past relevant work experience, I fulfill the requirements to be a qualified person for the purposes of NI 43-101.


I am responsible for Chapters 1, 2, 3, 15, 19 to 22, and 24 to 27 of this NI 43-101 Technical Report.

I did not visit the Kiena Property that is the subject of the Technical Report.






Colin, Hardie, P. Eng. BBA Inc.


bba.ca








I, Luciano Piciacchia, P. Eng., employed with BBA Inc., do hereby consent to the filing of the










Luciano Piciacchia, P. Eng.

BBA Inc.


bba.ca



This certificate applies to the NI 43-101 Technical Report titled "Preliminary Economic Assessment for the Kiena Mine

Complex Project, Val-d’Or, Québec, Canada" (the Technical Report), prepared for Wesdome Gold Mines Ltd.

(Wesdome) issued on June 17, 2020 and effective as of March 31, 2020.

I, Luciano Piciacchia, P. Eng., do hereby certify that:

I am an engineer and the director of Waste Management with BBA Inc. located at 2020 Robert-Bourassa Blvd., Suite 300, Montréal, Québec H3A 2A5, Canada.

I am a graduate of mining engineering from McGill University in 1981 and a Masters’ and Ph.D. focusing in soil and rock geotechnics, also from McGill in 1983 and 1988.

I am a member of the order of engineers in, Quebec, Ontario, Newfoundland & Labrador, British Columbia and Nunavut.

I have over 30 years of experience in geotechnical engineering with a focus on mining. I have applied my geotechnical / civil background to mine waste management, including waste rock, tailings and water.



I am responsible for the preparation of Chapter 18, and I am also responsible for the relevant portions of chapters 1, 2, 3, 25, 26 and 27 of the Technical Report.

I did visit the Kiena Property that is the subject of this Technical Report.








bba.ca








I, Pierre-Luc Richard, P. Geo., employed with BBA Inc., do hereby consent to the filing of the










Pierre-Luc Richard, P. Geo.

BBA Inc.

2020 Robert-Bourassa Blvd. Suite 300 Montréal, QC H3A 2A5 T +1 514.866.2111 F +1 514.866.2116 bba.ca


Pierre-Luc Richard, P. Geo.




I, Pierre-Luc Richard, P. Geo., do hereby certify that:

I am a Principal Geologist with BBA Inc. located at 2020 Robert-Bourassa Blvd, Suite 300, Montréal, Québec, Canada, H3A 2A5.

I am a graduate of Université du Québec à Montréal in Resource Geology in 2004. I also obtained a M.Sc. from Université du Québec à Chicoutimi in Earth Sciences in 2012.

I am a member in good standing of the Ordre des Géologues du Québec (OGQ Member No. 1119), the Association of Professional Geoscientists of Ontario (APGO Member No. 1714), and the Northwest Territories Association of Professional Engineers and Geoscientists (NAPEG Member No. L2465).

I have worked in the mining industry for more than 15 years. My exploration expertise has been acquired with Richmont Mines Inc., the Ministry of Natural Resources of Québec (Geology Branch), and numerous companies through my career as a consultant. My mining expertise was acquired at the Beaufor mine and several other producers through my career. I managed numerous technical reports, mineral resource estimates and audits as a consultant for InnovExplo from February 2007 to March 2018 and as a consultant for BBA since.



I am author and responsible for the preparation of Chapters 4 to 12, 14 and 23 of the Technical Report and co-author of the relevant sections of Chapters 1, 2, 3, 25, 26 and 27 of this NI 43-101 Technical Report.

I have visited the Kiena Property that is the subject of this Technical Report on August 6-8 and on other occasions during the course of other mandates.

I have had no prior involvement with the property that is the subject of the Technical Report, except for authoring Technical Reports in the past.


As at 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 the portions of the Technical Report for which I am responsible not misleading



Pierre-Luc Richard, P. Geo. BBA Inc.


bba.ca








I, Jorge Torrealba, P. Eng., employed with BBA Inc., do hereby consent to the filing of the










Jorge Torrealba, P. Eng.

BBA Inc.


bba.ca


Jorge Torrealba, P. Eng.




I, Jorge Torrealba, P. Eng., Ph.D., do hereby certify that:

I am employed as an engineer by and carried out this assignment for BBA Inc. – Consulting Firm in Engineering,

located at 2020 Robert-Bourassa Blvd., Suite 300, Montréal, Québec, Canada, H3A 2A5.

I graduated with a B.Eng. and M.Sc. in Metallurgy from Santiago de Chile University (Santiago, Chile) in 1998. I

obtained a Ph.D. degree in Metallurgy from McGill University (Montreal, Quebec) in 2005.

I am a member in good standing of the Association of Professional Engineers and Geoscientists of New Brunswick (APEGNB licence No. M7957) and a member of the Canadian Institute of Mining Metallurgy and Petroleum.

I have worked as an engineer for a total of 22 years since graduating from University in 1998. My expertise in Mineral processing has been acquired with Santiago de Chile University in Chile, with Chile University in Chile, with McGill University in Quebec. I have been a consulting process engineer for BBA Inc. since February 2005.



I am author and responsible for Chapters 13 and 17 of the Technical Report and co-author of the relevant sections of chapters 1, 2, 25, 26 and 27.

I have visited the Kiena Property that is the subject of the Technical Report on August 6 and 7, 2019.

I have had no prior involvement with the properties that are the subject of the Technical Report.





Jorge Torrealba, P. Eng., Ph.D. BBA Inc.




JUNE 2020 i

TABLE OF CONTENTS

1.1 Property Description and Ownership ....................................................................................... 1-1

1.2 Geology and Mineralization ...................................................................................................... 1-1

1.3 Status of Exploration and Drilling ............................................................................................. 1-3

1.4 Mineral Processing and Metallurgical Testing ......................................................................... 1-3

1.5 Mineral Resource Estimate ...................................................................................................... 1-5

1.6 Mineral Reserve Estimate ........................................................................................................ 1-7

1.7 Mining Methods ........................................................................................................................ 1-7

1.7.1 Introduction ............................................................................................................................... 1-7

1.7.2 Existing Infrastructure ............................................................................................................... 1-8

1.7.3 Mining Method ........................................................................................................................... 1-8

1.7.4 New Infrastructure ................................................................................................................... 1-10

1.7.5 Development ........................................................................................................................... 1-10

1.7.6 Production Plan ....................................................................................................................... 1-11

1.8 Recovery Methods ................................................................................................................. 1-14

1.9 Project Infrastructure .............................................................................................................. 1-20

1.10 Environmental and Permitting ................................................................................................ 1-21

1.11 Capital and Operating Costs Estimates ................................................................................. 1-22

1.11.1 Capital Costs ........................................................................................................................... 1-22

1.11.2 Operating Costs ...................................................................................................................... 1-24

1.12 Project Economics ................................................................................................................. 1-25

1.13 Interpretations and Conclusions............................................................................................. 1-28

1.14 Recommendations ................................................................................................................. 1-29

2.1 Wesdome Gold Mines Ltd. ....................................................................................................... 2-1

2.2 Basis of Technical Report ........................................................................................................ 2-1

2.3 Report Responsibility and Qualified Persons ........................................................................... 2-2

2.4 Effective Dates and Declaration ............................................................................................... 2-3

2.5 Currency, Units of Measure and Calculations .......................................................................... 2-4

2.6 Sources of Information ............................................................................................................. 2-5

2.6.1 General ..................................................................................................................................... 2-5

2.6.2 BBA ........................................................................................................................................... 2-5

2.7 Site Visits .................................................................................................................................. 2-7

2.8 Acknowledgement .................................................................................................................... 2-8




JUNE 2020 ii

3.1 Introduction ............................................................................................................................... 3-1

3.2 Mineral Tenure and Surface Rights ......................................................................................... 3-1

3.3 Taxation .................................................................................................................................... 3-1

3.4 Commodity Price Projections ................................................................................................... 3-2

4.1 Property Description and Location ........................................................................................... 4-1

4.2 Mineral Tenure ......................................................................................................................... 4-3

4.3 Royalties, Agreement and Encumbrances ............................................................................... 4-4

4.4 Environmental Liabilities .......................................................................................................... 4-6

4.5 Permitting ................................................................................................................................. 4-9

5.1 Accessibility .............................................................................................................................. 5-1

5.2 Climate ..................................................................................................................................... 5-2

5.3 Local Resources and Infrastructure ......................................................................................... 5-3

5.3.1 Airports, Rail Terminals and Bus Services ................................................................................ 5-3

5.3.2 Local Work Force ...................................................................................................................... 5-3

5.3.3 Additional Support Services ...................................................................................................... 5-3

5.4 Physiography............................................................................................................................ 5-4

5.5 Mine Infrastructure ................................................................................................................... 5-4

5.5.1 Crown Pillar – Kiena Mine ......................................................................................................... 5-6

6.1 Wesdome Principal History in the Project ................................................................................ 6-1

6.2 Detailed Historical Work by Area ............................................................................................. 6-4

6.2.1 Kiena Mine Area ........................................................................................................................ 6-4

6.2.2 Wisik Shaft Area ........................................................................................................................ 6-8

6.2.3 Shawkey Mine Area .................................................................................................................. 6-8

6.2.4 Elmac Shaft Area .................................................................................................................... 6-11

6.2.5 Joubi Mine Area ...................................................................................................................... 6-12

6.2.6 Dorval-Siscoe/Wesdome Deposit Area ................................................................................... 6-16

6.2.7 Siscoe Mine Area .................................................................................................................... 6-18

6.2.8 Siscoe Extension Area ............................................................................................................ 6-21

7.1 Regional Geology ..................................................................................................................... 7-1

7.1.1 Abitibi Greenstone Belt ............................................................................................................. 7-1




JUNE 2020 iii

7.1.2 Pontiac Subprovince ................................................................................................................. 7-6

7.2 Geology of the Kiena Mine Complex Area ............................................................................... 7-6

7.2.1 Stratigraphy ............................................................................................................................... 7-8

7.2.2 Intrusive Rocks ........................................................................................................................ 7-13

7.2.3 Structural Elements from Kiena Complex ............................................................................... 7-13

7.2.4 Large-scale Fault Zones ......................................................................................................... 7-14

7.2.5 Mineralization Types ............................................................................................................... 7-16

7.3 Mineralized Zones .................................................................................................................. 7-18

7.3.1 Kiena Mine .............................................................................................................................. 7-18

7.3.2 Siscoe Mine............................................................................................................................. 7-28

7.3.3 Shawkey Mine ......................................................................................................................... 7-28

8.1 Archean Greenstone-Hosted Orogenic Gold Deposits ............................................................ 8-1

8.2 Gold Mineralization in the Val-d'Or District .............................................................................. 8-3

9.1 Surface Exploration .................................................................................................................. 9-1

9.2 Underground Exploration ......................................................................................................... 9-1

10.1 Drilling Methodology ............................................................................................................... 10-1

10.1.1 Drillhole Location/Set-up ......................................................................................................... 10-1

10.1.2 Drillhole Orientation during Operation ..................................................................................... 10-1

10.1.3 Drilling ..................................................................................................................................... 10-2

10.1.4 Core Logging and Measurement ............................................................................................. 10-2

10.1.5 Core Storage ........................................................................................................................... 10-4

10.2 Recent Diamond Drilling ........................................................................................................ 10-4

11.1 Wesdome Data....................................................................................................................... 11-1

11.1.1 Core Handling, Sampling and Security ................................................................................... 11-1

11.1.2 Methods of Preparation, Processing and Analysis .................................................................. 11-3

11.1.3 Sample Shipping and Security ................................................................................................ 11-4

11.2 Quality Assurance and Quality Control (QA/QC) ................................................................... 11-5

11.2.1 Duplicates ............................................................................................................................... 11-6

11.2.2 Blanks ..................................................................................................................................... 11-9

11.2.3 Certified Reference Materials (Standards) ............................................................................ 11-11

11.2.4 Check Assays ....................................................................................................................... 11-13

11.3 Conclusion ............................................................................................................................ 11-16




JUNE 2020 iv

12.1 Site Visit ................................................................................................................................. 12-1

12.2 Sample Preparation, Analytical, QA/QC and Security Procedures........................................ 12-1

12.2.1 Drillhole Location ..................................................................................................................... 12-3

12.2.2 Downhole Survey .................................................................................................................... 12-3

12.2.3 Assays .................................................................................................................................... 12-4

12.3 Underground Voids ................................................................................................................ 12-4

12.4 Conclusion .............................................................................................................................. 12-4

13.1 CTRI 2018 Metallurgical Tests ............................................................................................... 13-1

13.1.1 Testwork Program ................................................................................................................... 13-1

13.1.2 Sample Preparation ................................................................................................................ 13-1

13.1.3 Cyanidation Tests ................................................................................................................... 13-2

13.2 BBA 2019 Metallurgical Testwork .......................................................................................... 13-3

13.2.1 Testwork Samples ................................................................................................................... 13-5

13.2.2 Comminution Tests ................................................................................................................. 13-7

13.2.3 Metallurgical Tests .................................................................................................................. 13-7

13.2.4 Reagent Consumptions ......................................................................................................... 13-20

13.2.5 Solid and Liquid Separation Tests ......................................................................................... 13-22

13.3 Summary .............................................................................................................................. 13-22

13.4 Conclusions .......................................................................................................................... 13-23

13.5 Recommendations ............................................................................................................... 13-25

13.6 Future Steps ......................................................................................................................... 13-26

14.1 Introduction ............................................................................................................................. 14-1

14.2 Methodology ........................................................................................................................... 14-1

14.3 Resource Database ............................................................................................................... 14-4

14.4 Geological Interpretation and Modelling ................................................................................ 14-6

14.4.1 Geological Model .................................................................................................................... 14-7

14.4.2 Voids Model .......................................................................................................................... 14-10

14.4.3 Overburden and Topography ................................................................................................ 14-10

14.5 Data Analysis - Block Model MRE ....................................................................................... 14-10

14.5.1 Raw Assay Statistics ............................................................................................................. 14-10

14.5.2 Compositing .......................................................................................................................... 14-11

14.5.3 Outlier Handling .................................................................................................................... 14-12

14.5.4 Density .................................................................................................................................. 14-24

14.5.5 Variogram Analysis ............................................................................................................... 14-24




JUNE 2020 v

14.6 Data Analysis - Polygonal MRE ........................................................................................... 14-26

14.6.1 Outlier Handling .................................................................................................................... 14-26

14.6.2 Density .................................................................................................................................. 14-27

14.7 Block Modelling .................................................................................................................... 14-27

14.7.1 Block Model Parameters ....................................................................................................... 14-27

14.7.2 Search Ellipsoid Strategy ...................................................................................................... 14-27

14.7.3 Interpolation Parameters ....................................................................................................... 14-31

14.7.4 Interpolation Methodology ..................................................................................................... 14-32

14.8 Block Model Validation ......................................................................................................... 14-32

14.8.1 Visual Validation .................................................................................................................... 14-32

14.8.2 Statistical Validation .............................................................................................................. 14-33

14.8.3 Swath Plots ........................................................................................................................... 14-35

14.9 Polygonal Mineral Resource Estimate methodology ........................................................... 14-37

14.10 Mineral Resource Classification ........................................................................................... 14-38

14.10.1 Mineral Resource Definition .................................................................................................. 14-39

14.10.2 Mineral Resource Classification for the Block Model MRE ................................................... 14-40

14.10.3 Mineral Resource Classification for the Polygonal MRE ....................................................... 14-41

14.11 Cut-off Grade........................................................................................................................ 14-44

14.12 Kiena Mine Complex Mineral Resource Estimate ................................................................ 14-45

16.1 Historical Mining ..................................................................................................................... 16-1

16.2 Existing Material/Waste Handling Facilities ........................................................................... 16-3

16.2.1 Shaft ........................................................................................................................................ 16-3

16.2.2 Loading and Haulage .............................................................................................................. 16-6

16.2.3 Crushing .................................................................................................................................. 16-6

16.2.4 Waste Disposal ....................................................................................................................... 16-6

16.2.5 Existing Services Facilities ...................................................................................................... 16-7

16.3 Geotechnical ........................................................................................................................ 16-13

16.3.1 Structure and Lithology ......................................................................................................... 16-13

16.3.2 Rock Strength ....................................................................................................................... 16-13

16.3.3 Rock Mass Characteristics .................................................................................................... 16-14

16.3.4 Stress Assumptions .............................................................................................................. 16-15

16.3.5 Anticipated Rock Mass Behaviour ......................................................................................... 16-16

16.3.6 Ground Support Requirements ............................................................................................. 16-17

16.4 Mining Method ...................................................................................................................... 16-18

16.4.1 Longhole Mining .................................................................................................................... 16-18

16.4.2 Underhand Longhole Mining ................................................................................................. 16-19




JUNE 2020 vi

16.4.3 Mineable Resource ............................................................................................................... 16-20

16.5 New and Upgraded Material Handling Facilities .................................................................. 16-27

16.5.1 Shaft ...................................................................................................................................... 16-27

16.5.2 Loading and Hauling ............................................................................................................. 16-28

16.5.3 Crushing ................................................................................................................................ 16-28

16.5.4 Waste Disposal ..................................................................................................................... 16-28

16.5.5 New and Upgraded Service Facilities ................................................................................... 16-29

16.6 Development ........................................................................................................................ 16-33

16.7 Production Plan .................................................................................................................... 16-37

17.1 Process Description ............................................................................................................... 17-1

17.2 Process Plant Design Criteria ................................................................................................ 17-3

17.3 Process Plant Facilities Description ....................................................................................... 17-7

17.3.1 Crushing Circuit ....................................................................................................................... 17-7

17.3.2 Grinding Circuit ....................................................................................................................... 17-7

17.3.3 Cyanidation ............................................................................................................................. 17-8

17.3.4 Carbon-in-pulp (CIP) Process ................................................................................................. 17-8

17.3.5 Acid Wash and Carbon Regeneration ..................................................................................... 17-9

17.3.6 Cyanide Destruction Circuit and Tailings Treatment ............................................................... 17-9

17.3.7 Reagents Systems ................................................................................................................ 17-10

17.4 Energy, Water and Consumable Requirements .................................................................. 17-10

17.4.1 Energy Requirements ........................................................................................................... 17-10

17.4.2 Water Requirements ............................................................................................................. 17-11

17.4.3 Consumables Requirements ................................................................................................. 17-11

18.1 Introduction ............................................................................................................................. 18-1

18.2 Site Access ............................................................................................................................. 18-1

18.3 Service and Administration Buildings ..................................................................................... 18-1

18.4 Personnel and Accommodation ............................................................................................. 18-1

18.5 Power and Electrical .............................................................................................................. 18-2

18.6 Communications .................................................................................................................... 18-2

18.7 Cyanide Destruction ............................................................................................................... 18-2

18.8 Filter Plant .............................................................................................................................. 18-3

18.9 Existing Tailings Storage Facility ........................................................................................... 18-3

18.10 Additional Tailings Management Facility ................................................................................ 18-4

18.10.1 Design Considerations ............................................................................................................ 18-5

18.10.2 Facility Configuration ............................................................................................................... 18-6




JUNE 2020 vii

18.11 Water Management ................................................................................................................ 18-9

18.12 Specific Infrastructure Upgrades for Restart ........................................................................ 18-11

19.1 Market Studies ....................................................................................................................... 19-1

19.2 Assumptions ........................................................................................................................... 19-1

19.3 Contracts ................................................................................................................................ 19-2

20.1 Environmental Studies and Issues ......................................................................................... 20-1

20.1.1 Environmental Studies ............................................................................................................ 20-1

20.1.2 Environmental Issues .............................................................................................................. 20-1

20.2 Waste Rock, Mineralized Material, Tailings and Water Management Strategy ..................... 20-2

20.2.1 Waste Rock Management ....................................................................................................... 20-2

20.2.2 Mineralized Material Management .......................................................................................... 20-3

20.2.3 Tailings Management .............................................................................................................. 20-3

20.2.4 Water Management ................................................................................................................. 20-3

20.3 Legal Aspects and Permitting ................................................................................................ 20-4

20.3.1 Federal .................................................................................................................................... 20-4

20.3.2 Provincial................................................................................................................................. 20-5

20.3.3 City of Val-d’Or ........................................................................................................................ 20-8

20.4 Social and Communities Issues ............................................................................................. 20-8

20.5 Closure and Rehabilitation ..................................................................................................... 20-8

20.5.1 Concepts ................................................................................................................................. 20-8

20.5.2 Cost Estimates ...................................................................................................................... 20-11

21.1 Capital Costs .......................................................................................................................... 21-1

21.1.1 Summary ................................................................................................................................. 21-1

21.1.2 Scope of Capital Cost Estimate ............................................................................................... 21-2

21.1.3 Structure and Basis of Capital Cost Estimate.......................................................................... 21-2

21.1.4 Mining Capital Costs ............................................................................................................... 21-4

21.1.5 Pre-production Capital Costs .................................................................................................. 21-6

21.1.6 Sustaining Capital Costs ......................................................................................................... 21-7

21.2 Operating Costs ................................................................................................................... 21-10

21.2.1 Summary ............................................................................................................................... 21-10

21.2.2 Basis of Operating Cost Estimate ......................................................................................... 21-12

21.2.3 Mining ................................................................................................................................... 21-12

21.2.4 Processing and Laboratory ................................................................................................... 21-12

21.2.5 Surface Operations ............................................................................................................... 21-13




JUNE 2020 viii

21.2.6 Technical Services ................................................................................................................ 21-13

21.2.7 HSE & Training ..................................................................................................................... 21-14

21.2.8 Administration ....................................................................................................................... 21-14

21.2.9 Tailings Management ............................................................................................................ 21-14

21.3 Site Personnel Summary – All Areas ................................................................................... 21-15

22.1 Assumptions and Basis .......................................................................................................... 22-1

22.2 Gold Production ..................................................................................................................... 22-3

22.3 Capital and Sustaining Costs ................................................................................................. 22-4

22.4 Royalties ................................................................................................................................. 22-4

22.5 Taxation .................................................................................................................................. 22-4

22.6 Financial Analysis Summary .................................................................................................. 22-5

22.7 Production Costs .................................................................................................................... 22-8

22.8 Sensitivity Analysis ................................................................................................................. 22-9

23.1 Canadian Malartic Property .................................................................................................... 23-1

23.2 Dubuisson JV Property .......................................................................................................... 23-1

23.3 Agnico Eagle Mines Ltd. Properties ....................................................................................... 23-2

23.4 Tarmac Project Property ........................................................................................................ 23-3

23.5 Harricana River Mining (O3 Mining) Property ........................................................................ 23-3

23.6 Knick Exploration Property ..................................................................................................... 23-3

23.7 Metanor Resources (Bonterra) Property ................................................................................ 23-3

23.8 Marban Block Property ........................................................................................................... 23-3

23.9 Siscoe East Property .............................................................................................................. 23-5

24.1 Project Execution Plan ........................................................................................................... 24-1

24.1.1 Project Organization ................................................................................................................ 24-1

24.1.2 Construction Management ...................................................................................................... 24-3

24.2 Project Execution Schedule ................................................................................................... 24-3

25.1 Overview ................................................................................................................................ 25-1

25.2 Data Verification and Mineral Resources ............................................................................... 25-1

25.3 Mining Methods ...................................................................................................................... 25-3

25.4 Metallurgy and Processing ..................................................................................................... 25-4

25.4.1 Metallurgy................................................................................................................................ 25-4




JUNE 2020 ix

25.4.2 Process Flowsheet .................................................................................................................. 25-5

25.5 Environmental Studies ........................................................................................................... 25-6

25.6 Infrastructure .......................................................................................................................... 25-6

25.7 Capital and Operating Costs .................................................................................................. 25-6

25.8 Indicative Economic Results .................................................................................................. 25-8

25.9 Project Risks and Opportunities ............................................................................................. 25-8

Drilling and Geology ............................................................................................................... 26-2

Metallurgical Testwork ........................................................................................................... 26-2

Mining ..................................................................................................................................... 26-3

Recovery Methods ................................................................................................................. 26-3

Infrastructure .......................................................................................................................... 26-4

Environment and Permitting ................................................................................................... 26-4

APPENDICES

Appendix A: Detailed list of mineral claims (verified on July 2, 2019)

Appendix B: List of drillholes on the Kiena Mine Complex Property

Appendix C: Polygonal resources summary

Appendix D: Tailings Storage Facility (TSF)

Appendix E: CPM Independent Gold Price Projections (March 30, 2020)




JUNE 2020 x

LIST OF TABLES

Table 1-1: Overall Au recovery for each testwork program ....................................................................... 1-4

Table 1-2: Underground Indicated and Inferred Mineral Resource Estimate ............................................ 1-6

Table 1-3: Total resource for mining .......................................................................................................... 1-8

Table 1-4: Resource removed from mining plan ........................................................................................ 1-9

Table 1-5: Resource recovery by zone ...................................................................................................... 1-9

Table 1-6: Resource scheduled ................................................................................................................. 1-9

Table 1-7: Capital development schedule ................................................................................................ 1-11

Table 1-8: Life of mine production ........................................................................................................... 1-12

Table 1-9: Project capital cost summary .................................................................................................. 1-23

Table 1-10: Operating costs summary ..................................................................................................... 1-25

Table 1-11: Financial analysis summary .................................................................................................. 1-26

Table 2-1: Qualified Persons and areas of report responsibility ................................................................ 2-2

Table 4-1: List of historical properties with their applicable royalties ......................................................... 4-6

Table 4-2: Wesdome certificates of authorization details ........................................................................ 4-10

Table 5-1: Kiena Mine Complex infrastructure ........................................................................................... 5-5

Table 6-1: Kiena mine production from 1981 to 2002 ................................................................................ 6-6

Table 6-2: Kiena mine production from 2006 to 2013 ................................................................................ 6-7

Table 6-3: Shawkey mine production from 1936 to 1964 ........................................................................ 6-11

Table 6-4: Joubi mine production from 1990 to 1999 .............................................................................. 6-15

Table 6-5: Siscoe mine yearly production from 1929 to 1949 ................................................................. 6-20

Table 10-1: Summary of the drilling completed on the Property during the 2018-2019 Program ........... 10-4

Table 11-1: Samples submitted to the laboratories for analysis during the 2018 to 2019 drilling

campaigns ........................................................................................................................... 11-6

Table 11-2: Standard Certified Reference Materials used at the Kiena Mine Complex during the

2018-2019 drilling campaign ............................................................................................. 11-12

Table 13-1: Composite splits and weights ............................................................................................... 13-1

Table 13-2: Composite feed assays ........................................................................................................ 13-2

Table 13-3: Cyanidation test results ........................................................................................................ 13-3

Table 13-4: Testwork samples ................................................................................................................. 13-6

Table 13-5: Comminution test results ...................................................................................................... 13-7




JUNE 2020 xi

Table 13-6: Whole ore leaching test conditions (three master composites) ............................................ 13-8

Table 13-7: WOL test results of all three master composites .................................................................. 13-9

Table 13-8: Knelson/Mozley gravity test results .................................................................................... 13-10

Table 13-9: Gravity tails leaching test ID codes..................................................................................... 13-11

Table 13-10: Gravity tails leaching test conditions for each master composite ..................................... 13-11

Table 13-11: Gravity tails leaching test results ...................................................................................... 13-12

Table 13-12: Overall recoveries of gravity and gravity tails leaching .................................................... 13-20

Table 13-13: Reagent consumptions per test type ................................................................................ 13-21

Table 13-14: Flocculant scoping test results summary .......................................................................... 13-22

Table 13-15: Overall Au recovery for each testwork program ............................................................... 13-23

Table 14-1: Mineralized zones of the 2019 MRE ..................................................................................... 14-6

Table 14-2: Basic statistics on raw assays for each domain ................................................................. 14-10

Table 14-3: Basic statistics on composites and high-grade capping value for each deposit ................ 14-23

Table 14-4: Summary of the density measurements ............................................................................. 14-24

Table 14-5: Variogram model parameters for each mineralized zone ................................................... 14-25

Table 14-6: Kiena Mine area block model parameters .......................................................................... 14-27

Table 14-7: Search ellipsoid ranges by interpolation passes ................................................................ 14-30

Table 14-8: Restricted search ellipsoid parameters .............................................................................. 14-31

Table 14-9: Interpolation parameters ..................................................................................................... 14-31

Table 14-10: Comparison of the block and composite mean grades at a zero cut-off grade for

Inferred and Indicated blocks ............................................................................................ 14-34

Table 14-11: Underground Indicated and Inferred Mineral Resource Estimate .................................... 14-45

Table 14-12: Indicated and Inferred block model Mineral Resource Estimate per lens ........................ 14-46

Table 14-13: Indicated and Inferred Polygonal Mineral Resource Estimate per zone .......................... 14-47

Table 14-14: Kiena Mine Complex Block Model Indicated and Inferred Mineral Resource cut-off

grade sensitivity table ........................................................................................................ 14-47

Table 16-1: Surface fresh air fans ............................................................................................................ 16-7

Table 16-2: Return air fan ........................................................................................................................ 16-7

Table 16-3: Rock strength assumptions ................................................................................................ 16-13

Table 16-4: Kiena Deep Zone, typical rock mass characteristics .......................................................... 16-15

Table 16-5: S50/Deep B Zone, typical rock mass characteristics ......................................................... 16-15

Table 16-6: Stress assumptions, Kiena Deep and S50/B zones ........................................................... 16-16




JUNE 2020 xii

Table 16-7: Wireframe evaluated using Deswik Stope Optimizer ......................................................... 16-21

Table 16-8: Resource wireframes utilized .............................................................................................. 16-22

Table 16-9: DSO key parameters .......................................................................................................... 16-23

Table 16-10: Potential Inferred Resource .............................................................................................. 16-23

Table 16-11: Potential Indicated Resource ............................................................................................ 16-23

Table 16-12: Total Resource.................................................................................................................. 16-24

Table 16-13: Resource removed ........................................................................................................... 16-24

Table 16-14: Resource recovered by zone ............................................................................................ 16-26

Table 16-15: Resource for schedule ...................................................................................................... 16-27

Table 16-16: New fleet of mobile equipment ......................................................................................... 16-28

Table 16-17: New service facilities ........................................................................................................ 16-29

Table 16-18: Capital development schedule .......................................................................................... 16-34

Table 16-19: Development sizes ........................................................................................................... 16-34

Table 16-20: Life of mine production ..................................................................................................... 16-38

Table 17-1: Project process design criteria for different scenarios .......................................................... 17-4

Table 17-2: Reagent mixing system ...................................................................................................... 17-10

Table 17-3: Process plant power demand by phase and area .............................................................. 17-11

Table 17-4: Estimated grinding media consumption .............................................................................. 17-12

Table 17-5: Reagents and consumables – Applications and consumption ........................................... 17-12

Table 18-1: Surface tailings production – TSF required capacity ............................................................ 18-6

Table 18-2: Project design precipitation data......................................................................................... 18-10

Table 19-1: NSR assumptions used in the economic analysis ................................................................ 19-1

Table 19-2: Existing contracts at the Kiena Mine Complex ..................................................................... 19-2

Table 21-1: Project capital costs summary .............................................................................................. 21-1

Table 21-2: BBA work breakdown structure (WBS) ................................................................................. 21-3

Table 21-3: Mining capital cost expenditures........................................................................................... 21-5

Table 21-4: Project pre-production capital cost summary ....................................................................... 21-6

Table 21-5: Project sustaining capital cost summary .............................................................................. 21-8

Table 21-6: Sustaining capital cost summary by year ............................................................................. 21-9

Table 21-7: Operating costs summary ................................................................................................... 21-10

Table 21-8: Operating costs breakdown per year for LOM .................................................................... 21-11




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Table 21-9: Cyanide destruction reagents cost ..................................................................................... 21-13

Table 21-10: Summary of personnel – All areas ................................................................................... 21-15

Table 22-1: Financial model parameters ................................................................................................. 22-2

Table 22-2: Financial analysis summary (pre-tax and after-tax).............................................................. 22-6

Table 22-3: Kiena Mine Complex project financial model summary ........................................................ 22-7

Table 22-4: Production cost summary ..................................................................................................... 22-9

Table 22-5: NPV sensitivity results (after-tax) for metal price and exchange rate variations ................ 22-10

Table 22-6: NPV sensitivity results (after-tax) for capital (LOM) and operating costs variations .......... 22-10

Table 22-7: IRR sensitivity results (after-tax) for metal price and exchange rate variations ................. 22-10

Table 22-8: IRR sensitivity results (after-tax) for capital (LOM) and operating costs variations ............ 22-11

Table 22-9: NPV sensitivity results (after-tax) for discount rate ............................................................. 22-11

Table 24-1: Key milestones (preliminary) ................................................................................................ 24-4

Table 25-1: Underground Indicated and Inferred Mineral Resource Estimate ........................................ 25-2

Table 25-2: Comparison of block model MRE and mining shapes .......................................................... 25-3

Table 25-3: Project capital cost summary ................................................................................................ 25-7

Table 25-4: Operating costs summary ..................................................................................................... 25-7

Table 25-5: Project risks and opportunities .............................................................................................. 25-9

Table 26-1: Work program budget ........................................................................................................... 26-1

LIST OF FIGURES

Figure 1-1: Mining zones included in this PEA (looking north) .................................................................. 1-7

Figure 1-2: Mining production - tonnes by year for each zone ................................................................ 1-13

Figure 1-3: Mining gold - ounces by year for each zone .......................................................................... 1-13

Figure 1-4: Kiena Mine process plant flowsheet (December 2011) ......................................................... 1-16

Figure 1-5: Average plant throughput per year based on LOM production plan ..................................... 1-17

Figure 1-6: Phase 1 (years 2021-2023) block flow diagram .................................................................... 1-18

Figure 1-7: Phase 2 (year 2024 and beyond) block flow diagram ........................................................... 1-19

Figure 1-8: Proposed tailings management facility layout ....................................................................... 1-20

Figure 1-9: Annual and cumulative project capital costs ......................................................................... 1-24

Figure 1-10: Sensitivity of the net present value (after-tax) to financial variables ................................... 1-27

Figure 1-11: Sensitivity of the internal rate of return (after-tax) to financial variables ............................. 1-28




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Figure 4-1: Property overview map of the Kiena Mine Complex ............................................................... 4-2

Figure 4-2: Kiena Mine Complex titles as of October 29, 2019 ................................................................. 4-3

Figure 4-3: Kiena Mine Complex mining royalties and historical property names ..................................... 4-5

Figure 4-4: Surface projection of the Kiena Mine Complex underground work, production and

exploration shafts as of end of 2018...................................................................................... 4-8

Figure 5-1: Location of the Kiena Mine Complex Property ........................................................................ 5-1

Figure 5-2: Gravel road leading to the property with an overview of the infrastructure ............................. 5-2

Figure 7-1: Location of the Abitibi Greenstone Belt within the Superior province ..................................... 7-2

Figure 7-2: Geologic map of the southern Abitibi Greenstone Belt and location of the Kiena Complex ... 7-4

Figure 7-3: Kiena Mine Complex property geology with historic and active mines and mineralized

zones ..................................................................................................................................... 7-7

Figure 7-4: Stratigraphic chart of the Kiena Complex region with relevant isotopic ages ....................... 7-12

Figure 7-5: Geological cross-section of S50 Zone ................................................................................... 7-19

Figure 7-6: S50 deep extension mineralized zones ................................................................................. 7-21

Figure 7-7: VC1 and VC6 mineralized zones........................................................................................... 7-22

Figure 7-8: Zone South mineralized zones .............................................................................................. 7-24

Figure 7-9: Kiena Deep A zones folded within the schist ........................................................................ 7-26

Figure 7-10: Typical cross-section of the Kiena Deep A Zones (1615NE) .............................................. 7-27

Figure 7-11: Kiena Deep B Zone ............................................................................................................. 7-28

Figure 8-1: Inferred crustal levels of gold deposition showing the different types of gold deposits and

the inferred deposit clan ........................................................................................................ 8-3

Figure 8-2: Schematic diagram of the geometric relationships between the structural elements of

veins and shear zones and the deposit-scale strain axes ..................................................... 8-5

Figure 10-1: Location of drillholes throughout the Property with their status ........................................... 10-5

Figure 10-2: Kiena Mine Complex diamond drillhole locations, close-up view of the 2019 MRE zone ... 10-6

Figure 11-1: Zoomed in scatterplot with linear trend of the coarse duplicates and original samples’

results from ALS for the 2018-2019 drilling program (n=1,611) .......................................... 11-7

Figure 11-2: Zoomed in scatterplot with linear trend of the pulp duplicates and original samples

results from ALS for the 2018-2019 drilling program (n=1,613) .......................................... 11-8

Figure 11-3: Zoomed in scatterplot with linear trend of the pulp duplicates and original samples’

results from Actlabs for the 2018-2019 drilling program (n=822) ........................................ 11-9

Figure 11-4: Results for blanks used by Wesdome during the 2018-2019 drilling program .................. 11-10

Figure 11-5: Scatterplot showing results from the check assay program for the metallic sieve

methodology ...................................................................................................................... 11-14




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Figure 11-6: Scatterplot showing results from the check assay program for the gravimetric finish

methodology ...................................................................................................................... 11-15

Figure 11-7: Scatterplot showing results from the check assay program for the atomic absorption

methodology ...................................................................................................................... 11-16

Figure 12-1: Core review in the core logging facility during the August 6-8, 2019 site visit .................... 12-2

Figure 12-2: Sampling procedures review during the August 6-8, 2019 site visit .................................... 12-2

Figure 12-3: Storage review during the August 6-8, 2019 site visit ......................................................... 12-3

Figure 13-1: Testwork program flowsheet ............................................................................................... 13-5

Figure 13-2: Samples collected to prepare composites A, A1 and A2 for 2019 testwork program ......... 13-6

Figure 13-3: WOL test kinetic curve ......................................................................................................... 13-9

Figure 13-4: Zone A, effect of pre-leaching and lead nitrate at 75 microns ........................................... 13-13

Figure 13-5: Zone A, effect of pre-leaching and lead nitrate at 100 microns ......................................... 13-14

Figure 13-6: Zone A, the effect of different CN addition on gold recovery ............................................ 13-15

Figure 13-7: A1 Zone, effect of pre-leaching and lead nitrate at 75 microns ......................................... 13-16

Figure 13-8: A1 Zone, effect of pre-leaching and lead nitrate at 100 microns ....................................... 13-16

Figure 13-9: A1 Zone, the effect of different CN addition amounts ....................................................... 13-17

Figure 13-10: Zone A2, effect of pre-leaching and lead nitrate at 75 microns ....................................... 13-18

Figure 13-11: Zone A2, effect of pre-leaching and lead nitrate at 100 microns ..................................... 13-18

Figure 13-12: Zone A2, the effect of different CN addition amounts ..................................................... 13-19

Figure 14-1: 2019 MRE block model and polygonal resources location ................................................. 14-3

Figure 14-2: 3D view looking west of the 3D model and of the drillholes included in this resource

estimate ............................................................................................................................... 14-5

Figure 14-3: 3D view showing the mineralized zones and undergrounds voids in the Kiena Mine

area looking west ................................................................................................................. 14-8

Figure 14-4: Plan view showing the mineralized zones and undergrounds drifts in the Kiena Mine

area looking down................................................................................................................ 14-9

Figure 14-5: Sample length distribution within the mineralized zones ................................................... 14-12

Figure 14-6: Graphs supporting capping threshold decisions on composites for the S50 deposit ........ 14-13

Figure 14-7: Graphs supporting capping threshold decisions on composites for the VC1 deposit ....... 14-14

Figure 14-8: Graphs supporting capping threshold decisions on composites for the VC6 Zone .......... 14-15

Figure 14-9: Graphs supporting capping threshold decisions on composites for the South Zone ........ 14-16

Figure 14-10: Graphs supporting capping threshold decisions on composites for the B Zone ............. 14-17




JUNE 2020 xvi

Figure 14-11: Graphs supporting capping threshold decisions on composites for the Kiena Deep A

Zone A subzones ............................................................................................................... 14-18


Zone A1 ............................................................................................................................. 14-19


Zone A2 – upper domain ................................................................................................... 14-20


Zone A2 – lower domain .................................................................................................... 14-21

Figure 14-15: Graphs supporting capping threshold decisions on composites for the H1ZA Zone ...... 14-22

Figure 14-16: Example of the variography study for the South Zone 132 ............................................. 14-26

Figure 14-17: Example of search ellipsoids for the S50 zones for the three interpolation passes ........ 14-29

Figure 14-18: Comparative example of the grade distribution between the blocks and the composites

in section (A) and longitudinal (B) views ........................................................................... 14-33

Figure 14-19: Block model validation swath plot along strike (X-direction) for the ZA Zone ................. 14-35

Figure 14-20: Block model validation swath plots across strike (Y-direction) for the ZA Zone ............. 14-36

Figure 14-21: Block model validation swath plots along elevation (Z-direction) .................................... 14-37

Figure 14-22: Longitudinal section 12200 N +/- 50m looking north of the Dubuisson zone showing

the polygonals for the 2019 MRE ...................................................................................... 14-38

Figure 14-23: Mineral Resource classification example for the S50_102 Zone .................................... 14-41

Figure 14-24: Polygonal MRE and discarded area example for the 2019 MRE .................................... 14-43

Figure 14-25: Polygonal MRE and of discarded area example for the 2019 MRE ................................ 14-44

Figure 14-26: Example of 3D views showing grade distribution and classification of the Kiena Deep A

Zone A ............................................................................................................................... 14-48

Figure 16-1: Historical production 1981 to 2019 ...................................................................................... 16-2

Figure 16-2: Mining zones included in this PEA ...................................................................................... 16-3

Figure 16-3: Shaft No. 1 compartments ................................................................................................... 16-4

Figure 16-4: Hoist drives and motors ....................................................................................................... 16-4

Figure 16-5: Surface mineralized material bins ....................................................................................... 16-5

Figure 16-6: Crusher ................................................................................................................................ 16-6

Figure 16-7: Repair facilities .................................................................................................................... 16-8

Figure 16-8: Second fueling station ......................................................................................................... 16-8

Figure 16-9: Dewatering pumps ............................................................................................................... 16-9

Figure 16-10: Surface compressor room ............................................................................................... 16-10




JUNE 2020 xvii

Figure 16-11: Surface compressors ....................................................................................................... 16-10

Figure 16-12: Process water distribution ............................................................................................... 16-11

Figure 16-13: Transformers T7 and T11 ................................................................................................ 16-12

Figure 16-14: Trend of the principal stress tensor (Azimuth 035°) relative to the geometry of Kiena

Deep and S50/B zones ...................................................................................................... 16-16

Figure 16-15: Excavation behaviour matrix as a function of rock mass quality and stress (depth) ....... 16-17

Figure 16-16: Longhole mining .............................................................................................................. 16-18

Figure 16-17: Crown recovery ............................................................................................................... 16-20

Figure 16-18: Mining zones.................................................................................................................... 16-22

Figure 16-19: Resource removed from South Zone .............................................................................. 16-25

Figure 16-20: Resource removed from VC Zone ................................................................................... 16-25

Figure 16-21: Resource removed from South Deep (S50) .................................................................... 16-26

Figure 16-22: VC Zone booster fan location .......................................................................................... 16-30

Figure 16-23: Kiena Deep Zone booster fan location ............................................................................ 16-31

Figure 16-24: South Zone access .......................................................................................................... 16-35

Figure 16-25: VC Zone access .............................................................................................................. 16-35

Figure 16-26: South Deep Zone access ................................................................................................ 16-36

Figure 16-27: Kiena Deep Zone access ................................................................................................ 16-36

Figure 16-28: Kiena Deep Mining fronts ................................................................................................ 16-39

Figure 16-29: Mining production - tonnes by year for each zone .......................................................... 16-40

Figure 16-30: Mining gold ounces by year for each zone ...................................................................... 16-40

Figure 16-31: Mining grade by year for each zone ................................................................................ 16-41

Figure 17-1: Kiena Mine process plant flowsheet (December 2011) ....................................................... 17-2

Figure 17-2: Throughput versus year based on LOM .............................................................................. 17-3

Figure 17-3: Phase 1 (years 2021 to 2023) block flow diagram and mass balance ................................ 17-5

Figure 17-4: Phase 2 (year 2024 and beyond) block flow diagram and mass balance ........................... 17-6

Figure 17-5: SAG mill at the Kiena Mine Complex .................................................................................. 17-7

Figure 18-1: Kiena mine site - Existing surface layout ............................................................................. 18-4

Figure 18-2: Proposed tailings management facility layout ..................................................................... 18-5

Figure 18-3: TSF configuration and typical cross-section ........................................................................ 18-8

Figure 18-4: Berm location and typical cross-section .............................................................................. 18-9




JUNE 2020 xviii

Figure 18-5: TSF water management infrastructure .............................................................................. 18-11

Figure 18-6: Drainage ditch conceptual cross-section ........................................................................... 18-11

Figure 19-1: CPM’s gold price projections to 2029 .................................................................................. 19-1

Figure 21-1: Distribution of pre-production capital costs (%) ................................................................... 21-7

Figure 21-2 Project sustaining capital cost summary (%) ........................................................................ 21-8

Figure 22-1: Annual payable gold production (oz) ................................................................................... 22-3

Figure 22-2: Overall Kiena Mine Complex project capital cost profile ..................................................... 22-4

Figure 22-3: Life of mine cash flow projection (cumulative, pre-tax and after-tax) .................................. 22-8

Figure 22-4: Sensitivity of the net present value (after-tax) to financial variables ................................. 22-12

Figure 22-5: Sensitivity of the internal rate of return (after-tax) to financial variables ........................... 22-12

Figure 23-1: Kiena Mine Complex adjacent properties ............................................................................ 23-6




JUNE 2020 xix

LIST OF ABBREVIATIONS AND UNITS OF MEASURE

TABLE OF ABBREVIATIONS

Abbreviation Description

σC uniaxial compressive testing

2D Two dimensional

3D Three dimensional

Axb SMC parameters

a Annum (year)

AA Atomic absorption

ADR Adsorption-desorption-recovery

AGB Abitibi greenstone belt

Ai Abrasion index

AISC All-in sustaining cost

Au Gold

B Billion

BBA BBA Inc.

BWi Bond Work index

C Carbon

Ca Calcium

CAD or $ Canadian dollar (examples of use: CAD 2.5M / $2.5M)

CaO Lime

CF Callahan Fault

CG Cadillac Group

CIM Canadian Institute of Mining, Metallurgy and Petroleum

CIP Carbon-in-pulp

CMT Construction management team

CND Cyanide destruction

CRMs Certified Reference Materials

CTRI Centre Technologique des Résidus Industriels

Cu Copper

CWi Crushing Work index

CSZ Callahan Shear Zone

DDH Diamond drillhole

DF Dubuisson Formation

DPFZ Destor-Porcupine fault zone

DSO Deswik Stope Optimizer

EA Environmental assessment

EAC Eurasian Conformity mark




JUNE 2020 xx



EBIT Earnings before interest and tax

et al. and others

FA Fire assay

Fe Iron

FS Feasibility study

GPa gigapascal

GRG Gravity recoverable gold

GTL Gravity tails leaching

HF Héva Formation

ID Identification

ID2 Inverse distance squared

IP Induced polarization

IRR Internal rate of return

ISRM International Society for Rock Mechanics

JF Jacola Formation

KCN Potassium cyanide

KNA Kriging Neighbourhood Analysis

KSZ K Shear Zone

LHD Load Haul Dump

LLCFZ Larder Lake-Cadillac fault zone

LVF La Motte-Vassan

m a.s.l. Metres above sea level

M Million

Ma Mega annuum (Million years)

MELCC Ministère de l’Environnement et de la Lutte contre les changements climatiques

MEND Mine Environment Neutral Drainage Program

MF Marbenite Fault

MFFP Ministère des Forêts, de la Faune et des Parcs

MG Malartic Group

MRE Mineral Resource Estimate

MS Metallic sieve

MTO Material take-offs

NaCN Sodium cyanide

NF Norbenite Fault

Ni Nickel

NN Nearest neighbour

No. Number




JUNE 2020 xxi



NPI Net profit interest

NPV Net present value

NQ NQ-Caliber drillhole

NSR Net smelter return

NTS National topographic system

O2 Oxygen

O/F Overflow

OK Ordinary kriging

P80 80% passing - Product size

Pb Lead

PEA Preliminary economic assessment

PF Parfouru Fault

PFS Prefeasibility study

PG Piché Group

pH Potential of hydrogen

PhD Doctor of philosophy

PO Pontiac Group

PS Pontiac Subprovince

QA/QC Quality Assurance / Quality Control

QP Qualified Person

RHF Rivière Héva Fault

ROM Run of mine

RQD Rock Quality Designation

RWi Rod Work index

S Sulphur

SAG Semi-autogenous (mill)

SCSE SAG circuit specific energy

SD Standard deviation

SEDAR System for electronic document analysis and retrieval

SG Specific gravity

SMC SAG mill comminution (test)

SPLP Synthetic Precipitation Leaching Procedure

Std Standard

ta Abrasion characteristics of the sample

TBC To be confirmed

TCLP Toxicity Characteristic Leaching Procedure

TiO2 Titanium dioxide




JUNE 2020 xxii



TMF Tailings management facility

TSF Tailings storage facility

TSS Total suspended solids

U Uranium

UCS Unconfined compressive strength

U-Pb Uranium-lead dating

USD or US$ United States dollar (examples of use: USD2.5M)

UTM Universal Transverse Mercator

VDF Val-d’Or Formation

vs. Versus

WBS Work breakdown structure

WOL Whole ore leach




JUNE 2020 xxiii

TABLE OF ABBREVIATIONS – UNITS OF MEASURE

Unit Description

Imperial

deg. or ° angular degree

ft2 square feet

ft2/d square feet per day

d day (24 hours)

°F Degrees Fahrenheit

Ø diameter

ft feet (12 inches)

gal gallon

ha Hectare

h hour (60 minutes)

in. or ” inch

in2 square inch

K Thousand (000)

lb pound

mi. miles

mPa megapascal

mph miles per hour

M Million

mesh US Mesh

min minute (60 seconds)

oz Troy ounce

oz/t Troy ounces per tonne

oz/y Troy ounces per year

ppm parts per million

psi pounds per square inch

% percent

%solids percent solids by weight

s second

st short ton (2,000 lbs)

Wk Week

wt% weight percent

yd. yard (36 inches)

y year (365 days)




JUNE 2020 xxiv


Unit Description

Metric

m3 cubic metre

d day (24 hours)

°C degree Celsius

Ø diameter

$/t Dollar per metric tonne

G Giga

g gram

g/t gram per (metric) tonne

h hour (60 minutes)

hp Horsepower

KCFM Thousand cubic feet per minute

kg kilogram

kg/t kilogram per tonne

km kilometre

km2 square kilometre

kPa kilopascal

kt kilotonne

L litre

m metre

mg milligram

MW Megawatt

ml millilitre

µm micron

mm millimetre

M Million

Mt Million tonne

Mtpa Million tonne per annum

ppm parts per million

% percent

SG specific gravity

m2 square metre

mm2 square millimetres

K Thousand (000)

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

tpa tonne per annum

tpd tonne per day

tph tonne per hour




JUNE 2020 xxv


Unit Description

Metric

tpy tonne per year

W Watt

WG water gauge

wt% weight percent

y year (365 days)

Wesdome Gold Mines



JUNE 2020 1-1

SUMMARY

This report was prepared as a Preliminary Economic Assessment-level (PEA) National Instrument

43-101 (NI 43-101) Technical Report (this "Technical Report" or "Report") for Wesdome Gold Mines

Ltd. (Wesdome) by BBA Inc. (BBA) on the Kiena Mine Complex Gold project (the “Project”).

1.1 Property Description and Ownership

The Project is in the Province of Quebec in the Abitibi-Temiscamingue administrative region within

the limits of the municipality of Val-d’Or and 100 km east of Rouyn-Noranda. It lies to the northwest

of the urban centre of Val-d’Or and covers 7,047 ha.

The coordinates for the approximate centre of the Project are latitude 48°08' N and longitude 77°54'

W (284105E and 5335715N: NAD 83 / UTM Zone 18).

As of October 29, 2019, the Kiena Mine Complex property consists of a contiguous group of 183

contiguous electronic map designated mining claims and one mining concession. From the 183

mining claims, Wesdome has 169 claims registered in their name at 100%. Six claims are held by

Mines Dynacor (50%) and Wesdome (50%), which represents the Maufort property. Eight claims

are held by Wesdome (75%) and 9264-7890 Québec inc. (25%), which represents the Siscoe

Extension property.

Some of the mining titles comprising the Project are subject to certain agreements and royalties.

The Project includes the milling and tailings facilities of the Kiena mine, nine shafts including the

Parker Shaft, and related underground workings from past producers and exploration projects, and

various surface facilities. Other than the exploration offices and underground exploration

development, the principal infrastructure of the Project has been under care and maintenance since

mid-2013.

1.2 Geology and Mineralization

The Kiena Mine Complex area is located in northwestern Quebec, straddling the southern part of

the Abitibi greenstone belt (AGB) and the northern part of the Pontiac Subprovince:

▪ The AGB comprises E-trending successions of folded and faulted volcanic and sedimentary

rocks and intervening domes of intrusive rocks. An important geologic feature of the AGB is

the occurrence of major, E-trending ductile-brittle fault zones. These zones cut across the

entire belt from the Kapuskasing structural zone in the west to the Grenville front in the east,

dividing the supracrustal rocks and intervening domes into distinct lozenge-shaped domains.

The most two important fault zones in the southern AGB are Destor-Porcupine fault zone

(DPFZ) in the north and Larder Lake-Cadillac fault zone (LLCFZ) in the south.

▪ The Pontiac Subprovince (PS) consists principally of a turbiditic succession composed of

graywacke and mudstone with minor intercalated conglomerate and basalt.

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JUNE 2020 1-2

The southern Abitibi greenstone belt in the Kiena Complex area consists of 2714–2700 Ma volcano-

plutonic assemblages, including the Malartic and Louvicourt groups, intruded by calc-alkaline

plutonic rocks. The Malartic Group comprises mainly komatiitic and tholeiitic basalt flows and sills,

with minor sedimentary rocks, which are interpreted as an oceanic floor in an extensional

environment related to mantle plumes, whereas the Louvicourt Group is mainly composed of mafic

to felsic volcanic rocks that formed in a subduction-related arc setting. From south to north, the

Kiena Complex area is underlain by the lithologies of the Pontiac Group (PO), the Piché Group

(PG), the Cadillac Group (CG), and formations belonging to the Louvicourt Group and the Malartic

Group.

The Kiena Complex area has a series of large-scale shear zones and related subsidiary faults

trending ESE-WNW to SE-NW, subparallel to stratigraphy and dipping steeply to the north. They

are, from south to north: the Larder Lake-Cadillac Fault Zone (LLCFZ), the Parfouru Fault (PF), the

Marbenite Fault (MF), the Norbenite Fault (NF), the Callahan Fault (CF), the K Shear Zone (KSZ)

and the Rivière Héva Fault (RHF). The Kiena Complex area is cut by all of them. The shear zones

contain dykes or stocks of monzonitic or tonalitic composition that vary widely in age (pre-, syn- or

post-tectonic) and are spatially associated with gold mines (Norlartic, Marban, Kiena, Sullivan,

Goldex, Siscoe, Joubi, Sigma and Lamaque). The observed diversity in the styles and ages of gold

mineralization related to these large-scale shear zones demonstrates that several distinct episodes

of mineralization occurred.

Gold mineralization on the Kiena Mine Complex shares many geological attributes with other vein-

type gold deposits of the Val-d’Or district and with orogenic gold deposits (also known as lode gold,

greenstone-hosted quartz-carbonate vein, or mesothermal deposits) in terms of host rock

composition, mineralogy and hydrothermal alteration.

Gold mineralization in the property occurs in all rock types except the Proterozoic dykes but is more

common in intrusive bodies and basalt as these acted as competent rock units that promoted

fracturing during deformation. Gold mineralization is concentrated where there is a marked

competency contrast between these competent units and the adjacent deformed komatiite and/or

chlorite-talc schists. There are at least two main gold mineralizing events in the region: young

deposits in which the gold mineralization did not experience much deformation after its

emplacement; and early mineralization in which mineralized bodies are commonly affected by D1

asymmetric folds, are highly strained and are locally dismembered. In a few deposits, both

generations are present. Gold-bearing veins in the region exhibit a great variety of orientations,

mineralogy and crosscutting relationships.

At least 63 mineralized zones have been observed on the property. In general, mineralized zones

on the property occur near a large-scale fault. They are often associated with a subsidiary shear

zone that may be proximal, adjacent or host to the mineralization. Alteration minerals are

dominantly albite, carbonates and pyrite with lesser chlorite and silica.

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JUNE 2020 1-3

A total of 48 mineralized zones were interpreted for the purpose of the 2019 MRE. This MRE formed

the basis for the development of the LOM plan.

The Kiena Deep A Zone was first intersected in December 2007 and is localized within the

Marbenite Fault deformation corridor and is divided into three main lenses and a fourth smaller

lens. The lenses occur along an isoclinal fold associated with the MF and a subsidiary fault.

1.3 Status of Exploration and Drilling

The resource database for the Project, as of August 6, 2019, consisted of 6,616 DDH (976,170.3 m)

with a cumulative length of 884,329.5 m including 187 DDH from the 2018-2019 drilling program.

As of March 24, 2020, Wesdome had completed an additional 164 DDH for 47,861 m that are not

included in the herein MRE. The Qualified Person (QP) is of the opinion that while the addition of

these new holes would increase knowledge and confidence on the Project, it would not materially

affect the MRE presented in this Report.

The QP, Pierre-Luc Richard, reviewed the drilling, sample preparation, analytical and security

procedures, as well as insertion rates and the performance of blanks, standards and duplicates for

the 2018-2019 drilling programs, and concluded that the observed failure rates are within expected

ranges and that no significant assay biases are present. The QP is of the opinion that the protocols

in place are adequate and followed. The database for the Project is of good overall quality and

adequate to industry standards. The QP is of the opinion that the database is appropriate for the

purpose of the MRE and that the sample density allows for a reliable estimate to be made of the

size, tonnage and grade of the mineralization in accordance with the level of confidence established

by the Mineral Resource categories in the CIM Standards.

1.4 Mineral Processing and Metallurgical Testing

In 2018, Wesdome mandated CTRI (Noël, 2019) to perform some preliminary gold leaching tests

on mineralized material from the Kiena Deep A Zone and S50 Zone. The objective of the testwork

was to benchmark the leaching performance of the Kiena Deep A Zone to the previously mined

Zone S50. In July 2019, BBA was selected by Wesdome to design and manage the 2019 testwork

program for the Kiena Mine Complex project. The testwork program started in August 2019 at SGS

Lakefield and was completed in November 2019. The objective of the testwork program was to

obtain preliminary design information and validate gold recovery for composites selected by

Wesdome and BBA.

The testwork consisted of chemical characterization, a preliminary evaluation of comminution

characteristics, a series of gravity, leaching (whole ore leach (WOL) and gravity tails leach (GTL)

as well as preliminary settling tests. The best overall gold recovery at comparable testwork

conditions is presented in Table 1-1.

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JUNE 2020 1-4

Table 1-1: Overall Au recovery for each testwork program

Composite ID Program Protocol P80

microns

Au calc

g/t

Au rec

%

NaCN cons.

kg/t

CaO cons.

kg/t

Kiena A, Zone 1 CTRI WOL 75 19.51 98.8 0.15 1.22




S50 Zone CTRI WOL 75 3.11 95.7 0.25 1.56

A SGS WOL(1) 75 19.60 99.0 0.49 0.93

A1 SGS WOL(1) 75 11.80 100.0 0.47 0.84

A2 SGS WOL(1) 75 19.60 99.0 0.45 0.89

A1 SGS GTL(2) 73 19.90 99.3 0.20 0.89

A2 SGS GTL(2) 75 10.70 99.0 0.20 0.84

A2 SGS GTL(2) 70 16.40 99.1 0.44 0.92

(1) 1 kg/t lead nitrate

(2) 150 g/t lead nitrate (in leaching).

The comminution test results have shown that all samples are moderately soft. Furthermore, the

Bond Abrasion test indicates that the samples are in the range of slightly abrasive to moderately

abrasive.

In Knelson/Mozley gravity tests, high gold recoveries of 61%, 67% and 59% were achieved for

samples A, A1 and A2, respectively. The addition of a gravity circuit to the existing process has the

potential to recover gold faster and reduce the amount of coarse gold particles reporting to the

leach circuit for improved process economics. More testwork is required to validate this observation.

The GTL tests resulted in extremely high gold recoveries; like WOL test results, the gold recoveries

varied between 95.7% and 100%. When gravity recoveries were added to the GTL recoveries, the

best overall gravity and gravity tails leaching recovery ranged from 99% to 99.3%. Additionally, the

impacts regarding the grind size, NaCN content, pre-aeration and lead nitrate addition (pre-

conditioning) on overall gold recoveries and leach kinetics were analyzed. Slightly increased gold

recovery was achieved by finer grinding (75 microns) and higher amount of NaCN addition. Pre-

conditioning was found to have minimal effect on the overall gold recoveries, but it improved the

leach kinetics. The majority of GTL test samples reached their maximum gold recoveries in

approximately 24 hours.

NaCN consumptions in WOL and GTL tests were less than 0.6 kg/t, which is considered low.

However, in GTL, sample A consumed 54%, sample A1 consumed 46% and sample A2 consumed

26% less cyanide than the amounts consumed in WOL. The amount of consumed CaO was found

to be slightly lower than 1 kg/t in each test for all samples. CaO consumptions of all samples were

found to be average.

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JUNE 2020 1-5

SNF Flomin 913 VHM was found to be the most effective and efficient flocculant among the

flocculants tested. The dosage range of 9-12 g/t was found to be relatively low.

Based on the results, there is no economic benefit to installing a gold gravity recovery circuit.

Therefore, the current gold processing circuit at Kiena appears adequate to maximize the gold

recovery, however this will be investigated further during the PFS.

1.5 Mineral Resource Estimate

The October 2019 Kiena Mine Complex Mineral Resource Estimate (the “2019 MRE”) was

prepared by Karine Brousseau, P. Eng., Wesdome, and audited by Pierre-Luc Richard, P. Geo.,

BBA, using all available information including historical and recent diamond drillholes.

The 2019 MRE used for this PEA includes drill data as of August 6, 2019 and includes an additional

140 drillholes, for a total of 36,050 m drilled since the close-out date of the last Kiena MRE

announced on December 12, 2018.

The 2019 MRE combines two different approaches:

▪ A block model mineral resource estimate for the zones in the former Kiena Mine area (the

"block model MRE") was prepared by Karine Brousseau, P. Eng., Senior Engineer of

Wesdome and has been reviewed and audited by Pierre-Luc Richard, P. Geo., Qualified

Person of BBA. Geological wireframes were constructed in Leapfrog Geo 4.5™. Geovia®

GEMS 6.8.2.2 was used for the compositing, 3D block modelling, interpolation and

classification. Statistical studies were conducted using Excel and Snowden Supervisor

v.8.11;

▪ A polygonal mineral resource estimate for the zones outside of the Kiena Area (the

“polygonal MRE”) was prepared by Turcotte et al. (2015) and reviewed and modified by

Pierre-Luc Richard, P. Geo., Qualified Person of BBA with the following steps:

- Confirmation of the lack of new material information;

- Review and validation of the resource with Excel;

- Review of the classification in long section;

- Review of the underground cut-off grade.

The block model resource database for the Kiena Mine Area, as of August 6, 2019, consisted of

349 surface and 6,267 underground drillholes with a cumulative length of 893,318.52 m. The

polygonal resource database consisted of 216 drillholes cutting across the mineralized zones. No

new information was added to the polygonal resource database since 2015.

The mineral resources are not mineral reserves as they do not have demonstrated economic

viability. The estimate is categorized as Indicated and Inferred Resources based on data density,

geological and grade continuity, search ellipse criteria, drillhole density and specific interpolation

parameters. The effective date of the 2019 MRE is September 25, 2019 based on the compilation

status and cut-off grade parameters.

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JUNE 2020 1-6

The total Indicated and Inferred Mineral Resource Estimate for the block model MRE and the

polygonal MRE is presented in Table 1-2:

Table 1-2: Underground Indicated and Inferred Mineral Resource Estimate

Indicated Resources Inferred Resources

Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Block Model MRE 968,900 14.46 450,400 1,121,200 11.02 397,100

Polygonal MRE 1,859,300 5.65 337,800 1,796,900 6.94 401,000

Total 2,828,200 8.67 788,100 2,918,100 8.51 798,100

Notes to Table 1-2:

The independent qualified person for the 2019 MRE, as defined by NI 43 101, is Pierre Luc- Richard,

P. Geo., of BBA. The effective date of the estimate is September 25, 2019.

These mineral resources are not mineral reserves as they do not have demonstrated economic viability.

The mineral resource estimate follows CIM definitions and guidelines for mineral resources.

Results are presented in situ and undiluted and considered to have reasonable prospects for economic

extraction, below 100 m crown pillar.

The estimation combined two estimation methods, ordinary kriging in the Kiena Complex and polygonal

for other deposits on the property.

The Kiena Complex resources encompass 20 zones with a minimum true thickness of 3.0 m using the

grade of the adjacent material when assayed or a value of zero when not assayed. High-grade capping

varies from 20 g/t to 200 g/t Au (when required) was applied to composited assay grades for interpolation

using an Ordinary Kriging interpolation method based on 1.0 m composite and block size of 5 m x 5 m x

5 m, with bulk density values of 2.8 (g/cm3). In addition, a high grade limit or second capping value was

used for the second and third pass grade interpolation to restrict high grade impact at greater distance

from the drillhole intersect. Indicated resources are manually defined and enclose areas where drill

spacing is generally less than 25 metres, blocks are informed by a minimum of three drillholes, and

reasonable geological and grade continuity is shown.

The zone outside the Kiena Complex encompasses eight zones with a minimum true thickness of 1.5 m

using a polygonal estimation method. Indicated resources were estimated from drillhole results using the

mid distance between drillhole or a maximum of 30 m, 12.5 m in some areas. The high-grade capping

was fixed at 34.28 g/t Au with a bulk density value of 2.8 (g/cm3).

The estimate is reported for potential underground scenario at cut-off grades of 3.0 g/t Au (>40° dip) and

4.0 g/t Au (<40° dip, Wesdome Zone). The cut-off grades were calculated using a gold price of USD 1,300

per ounce, a CAD:USD exchange rate of 1.31 (CAD 1,700); mining cost $110/t (>40° dip); $150/t (<40°

dip); processing cost $35/t; G&A $15/t. The cut-off grades should be re-evaluated in light of future

prevailing market conditions (metal prices, exchange rate, mining cost, etc.).

The number of metric tonnes and ounces were rounded to the nearest hundred and the metal contents

are presented in troy ounces (tonne x grade / 31.10348).

The QP, Pierre-Luc Richard, P. Geo., is not aware of any known environmental, permitting, legal, title-

related, taxation, socio-political or marketing issues, or any other relevant issue not reported in this

Technical Report that could materially affect the mineral resource estimate.

Wesdome Gold Mines



JUNE 2020 1-7

1.6 Mineral Reserve Estimate

Since this Report summarizes the results of a Preliminary Economic Assessment (PEA), no Mineral

Reserves have been estimated for the Kiena Mine Complex project as per NI 43-101 guidelines.

1.7 Mining Methods

1.7.1 Introduction

Since the suspension of operations, Wesdome has completed extensive underground diamond

drilling to enhance the mineral resource in four zones of the mine (South, South Deep, VC and

Kiena Deep). Figure 1-1 illustrates the four zones that are considered within the scope of this PEA.

Figure 1-1: Mining zones included in this PEA (looking north)

Wesdome Gold Mines



JUNE 2020 1-8

1.7.2 Existing Infrastructure

The key existing infrastructure includes the 930 m deep shaft and its hoist, with an average hoisting

capacity of 2,000 tpd; the underground crusher; and the ventilation system with two fresh air fans

and one return air fan. Also existing are surface shops, underground repair facilities, fueling,

backfill, dewatering, compressed air, process water, and electrical facilities.

1.7.3 Mining Method

Future mining will continue utilizing the historical longhole mining method, and development will be

performed utilizing standard development methods. When mining under previously mined areas,

underhand longhole mining will be utilized.

Resource wireframes and Block Models were evaluated using Deswik Stope Optimizer (DSO) to

create mineable shapes above cut-off and determine potential Indicated and Inferred Resource that

is available to be used for design purposes. A summary of all resources considered for mining is

provided in Table 1-3. It should be noted that none of the Polygonal MRE Resource shown in

Table 1-2 were included in mine designs as they are outside the Kiena mine area.

Table 1-3: Total resource for mining

Class Zone DSO total

(tonne) Contained gold in DSO shapes

(ounces) grade (g/t)

All

South 173,439 21,633 3.88

VC 208,047 34,057 5.09

S50 & B 228,091 29,187 3.98

Kiena Deep 1,772,954 719,935 12.63

Total 2,382,530 804,813 10.51

A review of the mineable shapes and their location, with respect to existing infrastructure and

planned infrastructure, was completed and DSO shapes that did not contain enough value for

development and mining costs have been removed (refer to Table 1-4). Two lenses have been

removed entirely; the rest of the removed DSO shapes are in the remaining lenses.

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JUNE 2020 1-9

Table 1-4: Resource removed from mining plan


(tonne) Contained gold in DSO shapes


All

South 39,157 4,283 3.40

VC 9,996 1,214 3.78

S50 & B 19,140 2,122 3.45

Kiena Deep 1,240 161 4.04

Total 69,532 7,780 3.48

A recovery factor was applied to the remaining resources specific to each zone and is illustrated in

Table 1-5. This recovery accounts for losses expected within each zone during the mining process.

The dilution reported in Table 1-5 is the percentage of waste (at 0 g/t) included in the resource

calculations.

Table 1-5: Resource recovery by zone

Zone Resource recovery

(%) Dilution (%)

South 80 7

VC 80 12

South Deep 80 8

Kiena Deep 90 21

Table 1-6 illustrates that a total of 2.07 Mt grading 10.64 g/t resource is included in the mine

schedule. Based on a processing plant recovery of 97%, this will produce 688,000 ounces of gold.

Table 1-6 has an additional 4,111 tonnes at 4.08 g/t from the South Zone that was not included in

Table 1-7 as a result of the 2020 change in plans.

Table 1-6: Resource scheduled

Class Scheduled (tonnes) Gold in schedule


Indicated 1,031,059 383,131 11.56

Inferred 1,043,296 326,473 9.73

Total 2,074,354 709,604 10.64

Wesdome Gold Mines



JUNE 2020 1-10

1.7.4 New Infrastructure

A new fleet of mobile equipment will be purchased to support the mine plan. Development and

production drills are diesel driven and electric/hydraulic operated. All prime movers (LHDs and

trucks) are diesel as well as all miscellaneous equipment. During the PFS a trade-off on the use of

electric battery versus diesel will be completed.

All existing infrastructure is considered operational and fit for use, however, new facilities will be

required such as refuge, fueling, backfill, and electrical stations, latrines, and a new repair bay, fuse

mag, and explosives mag.

The current surface fresh air and return air fan installations should not require upgrading but will

require a full inspection. The current mine air heater systems may require upgrading. Two additional

return air booster fan installations will be required, one in the VC Zone, and one in the Kiena Deep

Zone. A heat study will need to be conducted to determine whether a cooling system is required.

1.7.5 Development

Development is scheduled to start on the face locations identified by Kiena Mine personnel and will

commence in 2021. During the year 2020, some development is planned for exploration purposes

and replacement of a section of Return Air Raises that will be completed by a contractor. The year

2021 will be a year of transition for the mine as it will move from contractor development to

Company development crews. This transition period will take place over 18 months as equipment

arrives for the mine crews. Based on the zone wireframes, ramps and level access drifts have been

designed to access the identified potential mineable resource. In the South and Kiena Deep zones,

some mineable shapes can be accessed from existing ramps and levels. Table 1-7 indicates the

required lateral ramp and access development required to support the Mine Production Schedule.

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JUNE 2020 1-11

Table 1-7: Capital development schedule

Zone Unit Maximum monthly

Total 2021 2022 2023 2024 2025 2026 2027 2028 2029

South m 84 3,803 985 986 986 847 0 0 0 0 0

Avg. m/d 2.8 2.7 2.8 2.8 2.8 2.6 0.0 0.0 0.0 0.0 0.0

VC m 84 4,513 985 985 986 988 569 0 0 0 0

Avg. m/d 2.8 2.8 2.8 2.8 2.8 2.8 2.7 0.0 0.0 0.0 0.0

S50 & B m 84 1,846 986 841 19 0 0 0 0 0 0

Avg. m/d 2.8 2.5 2.8 2.4 0.6 0.0 0.0 0.0 0.0 0.0 0.0

Kiena m 167 6,565 762 1,548 1,710 1,658 883 0 0 0 0

Avg. m/d 5.6 3.9 2.1 4.3 4.8 4.6 3.7 0.0 0.0 0.0 0.0

Total m 395 16,727 3,718 4,360 3,701 3,493 1,452 0 0 0 0

Avg. m/d 13.2 9.9 10.4 12.2 10.4 9.8 6.1 0.0 0.0 0.0 0.0

1.7.6 Production Plan

The production plan is based on a mill start-up of mid-2021; therefore, only sill production is planned

in the first half of 2021, which will be stockpiled. Each zone was reviewed, and a composite mining

rate developed in tonnes per day (tpd) that accounts for sill development and the mining activities

(drill, blast, muck and fill). This composite rate was applied to the zone tonnage to produce the LOM

production schedule (refer to Table 1-8). Kiena Deep being the largest of the four zones has been

divided into Seven Mining Fronts (A to G) for scheduling purposes. Each Mining Front consists of

five levels to enable flexibility in the mining plan. (Note: Table 1-6 has an additional 4,111 tonnes

at 4.08 g/t from the South Zone that was not included in Table 1-8).

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Table 1-8: Life of mine production

Kiena Deep Zone Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029

Trammed Mineralized Material 1,631,353 tonne 16,084 88,974 202,665 217,815 223,276 243,451 270,852 260,401 107,835

Mining Rate 585 tpd 45 250 569 610 627 684 761 729 609

Mined Grade Kiena Deep 12.35 g/t 9.77 11.42 12.23 11.40 11.01 12.62 13.36 12.54 14.81

Mined 647,797 ounce 5,052 32,672 79,676 79,848 79,025 98,812 116,351 105,010 51,351

VC Zone Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029

Trammed Mineralized Material 163,157 tonne 4,689 26,794 35,405 35,405 35,405 25,459 0 0 0

Mining Rate 93 tpd 13 75 99 99 99 72 0 0 0

Mined Grade VC 5.01 g/t 4.86 5.69 4.94 4.40 5.43 4.69 0.00 0.00 0.00

Mined Gold 26,275 ounce 733 4,900 5,621 5,004 6,179 3,839 0 0 0

S50 & B Zone Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029


Mining Rate 80 tpd 80 80 80 80 80 79 0 0 0

Mined Grade South Deep 3.95 g/t 3.71 3.91 3.42 3.92 4.22 4.52 0.00 0.00 0.00

Mined Gold 21,652 ounce 3,393 3,577 3,133 3,601 3,862 4,087 0 0 0

South Zone Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029

Trammed Mineralized Material 105,102 tonne 0 28,235 28,470 28,548 19,849 0 0 0 0

Mining Rate 79 tpd 0 79 80 80 56 0 0 0 0

Mined Grade South 3.95 g/t 0.00 3.82 3.93 4.05 4.01 0.00 0.00 0.00 0.00

Mined Gold 13,341 ounce 0 3,468 3,595 3,718 2,560 0 0 0 0

All Zones Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029

Operating Days 3,027 day 356 356 356 357 356 356 356 357 177


Mining Rate 684 tpd 138 484 829 869 862 834 761 729 330

Mined Grade 10.65 g/t 5.80 8.05 9.70 9.24 9.28 11.18 13.36 12.54 14.81

Mined Gold 709,065 ounce 9,177 44,616 92,025 92,171 91,626 106,737 116,351 105,010 51,351

Backfill Required 1,219,930 tonne 29,018 101,636 173,845 182,865 180,911 175,050 159,609 153,451 63,546

Backfill (from Tails) 790,804 tonne 2,902 25,409 121,692 128,005 126,638 122,535 111,726 107,415 44,482

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JUNE 2020 1-13

Figure 1-2 and Figure 1-3 show the tonnes and ounces by year for each zone for the LOM Plan.

Figure 1-2: Mining production - tonnes by year for each zone

Figure 1-3: Mining gold - ounces by year for each zone

Wesdome Gold Mines



JUNE 2020 1-14

1.8 Recovery Methods

The Kiena Mine processing plant became operational in September 1984. A conventional gold

recovery process involving cyanidation and conventional CIP was used. A cyanide destruction

system was initially installed and operated for about eight years before being taken offline. The

principal process steps included crushing, grinding, leaching by cyanidation, gold adsorption and

desorption, electrolysis, melting and casting of doré bars. Figure 1-4 shows the process flowsheet

of the Kiena plant as at December 2011.

The Kiena Mine Complex was designed with a plant throughput of 1,250 tpd, however crushing

capacity was added, and the plant has run as high as 2,000 tpd in the past. A new tailings storage

facility at the Kiena Mine Complex will be available in year 2024. It is envisioned that, during the

first three years of operation (namely Phase 1), the process plant will operate “as-is” at a maximum

throughput of 829 tpd until the new TSF is operative. In year 2024 (Phase 2), the throughput will

increase (maximum 869 tpd) based on the latest LOM and a new tailings plant facility will be

operative (see Section 18.10). As part of the dry stack initiative for storing process plant tailings, a

filter plant will be installed in Phase 2. The processing plant flowsheet will be modified to include

the addition of a cyanide destruction circuit and a filtration plant.

The design criteria are based on a process plant design throughput of 829 tpd (years 2021 to 2023)

and 869 tpd (year 2024 and beyond). The design values were selected from the highest throughput

from each phase (Figure 1-5).

The process plant consists of a jaw-crusher (located at the mine), ore handling equipment (30 in x

10 ft vibrating feeder and a 30-inch belt conveyor) and storage (35-tonne capacity hopper, two

coarse rock silos, both with a capacity of 600 tonnes), a secondary crusher with a 1,800-tonne silo,

a 1,000 hp semi-autogenous (SAG) mill (11'6" x 18'8") followed by a 900 hp ball mill (10'6" x 13')

operating in closed circuit and two stages of cyclones for classification. Design grinding product

was 75 microns (P80).

Based on potential future gravity testwork, a gravity circuit (currently not shown on the PFDs) may

be installed and fed by a portion of the feed or underflow of the cyclone (to be confirmed by

testwork). The cyclone overflow will be fed to a vibrating 20-mesh screen. The screen underflow

will be directed to a 65-ft diameter process thickener. Thickener overflow will be pumped to a series

of three carbon columns tanks. Thickener underflow will be forwarded to a series of three leach

tanks (retention time in the carbon columns and leach tanks is approximately 11 minutes and

36 hours, respectively), based on a plant feed of 1,250 tpd. The leach tails are pumped to five CIP

tanks in series. The discharge of the CIP tanks is pumped to a 28-mesh safety screen.

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JUNE 2020 1-15

Figure 1-4 presents the tailings handling configuration (as of December 2011), where the safety

screen undersize can be treated to produce backfill or sent to tailings, but there was no cyanide

destruction option (Section 17.3.6 presents an alternative proposed configuration). In Phase 1, the

slurry is sent to primary and secondary backfill cyclones. The cyclones' underflow (coarse particles)

is then sent to the backfill plant. The fine particle overflow from the cyclones is directed to the 75-ft

diameter waste-thickener, from where the fine slurry is sent to the tailings. The dilution water is

recycled toward the grinding circuit.

Block flow diagrams for the processing plant during Phase 1 (plant “as-is”) and Phase 2 (plant

including cyanide destruction and dry stacking) are presented in Figure 1-6 and Figure 1-7. In

Phase 1, the existing TSF is utilized, and in Phase 2, the deposition of tailings switches to a new

facility. This two-phased approach optimizes cashflow by delaying the capital investment for the

filter plant, cyanide destruction circuit, and new tailings storage facility.

According to the current production plan, 38% of the tailings will be used for backfilling on average

over the life of the mine, with the remaining being stored at the TSF. This value is lower for the

years 2021 and 2022 (6% and 15% respectively) as more waste is available. From 2023 to 2029,

it is estimated that 41% of the tailings will be used for backfilling. There may be an opportunity to

store most of the tailings for the first two years underground behind hydrostatic plugs (to be

investigated in the PFS).

A 2-t carbon per day Zadra processing adsorption-desorption-recovery (ADR) circuit and gold room

will be used to recover the gold and produce doré. The payable metal Au recovery is estimated to

be 95.7% to 99.7% based on the current testwork program. The plant also includes a reagent

preparation area and two process water circuits, i.e., cyanide bearing and cyanide free, to service

the entire plant.

The connected load for the plant was estimated at 1.4 MW and 2.1 MW (Phase 1 and 2 respectively)

with an annual power consumption of 9.4 GWh and 13.9 GWh (Phase 1 and 2 respectively).

Wesdome Gold Mines



JUNE 2020 1-16

Figure 1-4: Kiena Mine process plant flowsheet (December 2011)

Wesdome Gold Mines



JUNE 2020 1-17

Figure 1-5: Average plant throughput per year based on LOM production plan

Wesdome Gold Mines



JUNE 2020 1-18

Figure 1-6: Phase 1 (years 2021-2023) block flow diagram

Wesdome Gold Mines



JUNE 2020 1-19

Figure 1-7: Phase 2 (year 2024 and beyond) block flow diagram




JUNE 2020 1-20

1.9 Project Infrastructure

The Kiena Complex produced gold up until June 2013 and has been under care & maintenance

since then. It is therefore expected that most of the surface infrastructure can be reused as is to

support the planned production rate as the future operation will typically be at a rate below historical

operating rates. An exception is the power distribution system, which will need refurbishing prior to

restarting operations.

The existing tailings storage facility, which consists of two storage cells and a polishing pond, is

expected to reach capacity within 50 months after the restart of operations. A new facility is

therefore required to store the tailings that will be generated by the future operations. In order to be

conservative and allow for risk mitigation and operational flexibility at this early stage of design, the

new TSF has been scheduled to come online in 2024, roughly 30 months after start-up. The

opportunity to optimize the TSF design and timing will be explored during the prefeasibility study.

This new tailings management system will be of the dry stack type, and will include a filter plant,

stacking area, and water treatment plant. It will also require the refurbishment of the existing

cyanide destruction system and the relocation of the core shack. The tailings will be pumped from

the processing plant to the filter plant. The dry product at 85% solids will be trucked to the tailings

storage facility, and the filtrate will be recirculated to the processing plant to be reused as process

or fresh water. If the drainage water from the tailings storage facility does not meet the discharge

quality criteria, it will be pumped to the existing tailings pond. Figure 1-8 provides an overview of

the proposed system.

Figure 1-8: Proposed tailings management facility layout




JUNE 2020 1-21

1.10 Environmental and Permitting

The mineralized material production rate of the Project will be approximately 800 tpd to 900 tpd and

therefore the Project is not subject to the Canadian Impact Assessment Act Process. However, the

Metal and Diamond Mining Effluent Regulations will apply to the Project.

The project is not subject to the Quebec Environmental Impact Assessment and Review Process.

In fact, most of the project infrastructure have been permitted under the 1984 Certificate of

Authorization which allows for the mining and milling of 2,200 tpd. However, the new Tailings

Management Facilities (TMF) and the relocation of the waste rock storage dumps must be permitted

by new Certificates of Authorization under the 2012 version of Directive 019. Modifications to the

existing Certificate of Authorization will also be required for various topics including cyanide

destruction for tailings used for backfilling or sent to the new TMF.

Approximately 62 % of the tailings produced at the mineralized material treatment plant will be sent

to the TMF. The technology selected for the TMF is the dry stack method and the selected site is

located east of the existing tailings storage area at the existing location of the waste rock dump.

Various environmental and geotechnical baseline studies must be carried out before establishment

of the new TMF.

The main challenge for the establishment of the new TMF is the lack of available surface area due

to the proximity of the lake and the CN railroad. Social resistance from some stakeholders is a

possibility.

The waste rock produced will be used for backfilling or stored in underground workings. The waste

rock stored on the existing pile will be used for water management at the dry-stack area.

The new project includes the mining of new deposit zones. Therefore, it is possible that mineralized

material from the new zones could be different from the mineralized material mined during the

previous phases of the Kiena project. In this context, a geochemical characterization study must be

carried out on representative mineralized material samples or ideally on representative tailings

samples produced during metallurgical studies.

A new version of the Closure and Rehabilitation Plan must be produced for the new phase of the

Project. Rehabilitation works will include buildings and infrastructure dismantling, site safety,

ground ripping and revegetation of impacted area such as the infrastructure’s footprints and some

roads, as well as revegetation of tailings management facilities and waste rock piles.

Buildings and infrastructure will be dismantled to retrofit the sites to a state compatible with the

surrounding environment. Other infrastructure may be maintained for the benefit of the local

communities.

The closure and rehabilitation costs, considering that all buildings will be dismantled, are estimated

at $10.0M. A closure and restoration plan for the existing Kiena project has been submitted to

MERN and approved June 19, 2017. The total cost of reclamation was estimated at $7.0M.

Wesdome has contributed to a financial guarantee fund for this amount. The additional reclamation

costs not covered by the current financial fund is therefore estimated at $3.0M.




JUNE 2020 1-22

The following permits and approvals will be required for the Project:

▪ Certificate of Authorization for the new Tailings Management Facilities (TMF) as per Section

22 of the Environmental Quality Act;

▪ Certificate of Authorization for the waste rock storage as per Section 22 of the Environmental

Quality Act;

▪ Modification of the existing Certificate of Authorization for various topics including cyanide

destruction;

▪ Update Closure and Rehabilitation Plan as per Mining Act;

▪ Approval of the TMF location as per the Mining Act;

▪ Approval of the waste rock dump location as per the Mining Act.

1.11 Capital and Operating Costs Estimates

1.11.1 Capital Costs

The total pre-production capital cost for the Kiena Gold Project is estimated to be $43.7M including

allowances for indirect costs and contingency of $2.7M and $2.9M respectively. The pre-production

capital expenditures include refurbishments in the mine and the process plant, mine development,

mining equipment, engineering, and field programs.

The cost estimate meets the AACE class 4 requirements and has an expected accuracy of +/-35%

of the final Project cost. The capital cost estimate was compiled using historical development costs,

budgetary quotations, database costs, and database factors. Items such as sunk costs, taxes,

permitting, licensing, and financing costs are not included in the cost estimate. The pre-production

mine development costs for the year 2020 are considered as sunk costs. The contingency was set

at 25% but has not been applied to items whose scope is considered well defined, such as mine

development, mining equipment, and a number of mine infrastructure items.

Costs are expressed in Q4 2019 Canadian dollars with an exchange rate of CAD 1.00 for USD 0.76

with no allowances for escalation, currency fluctuation or interest during construction.

The cumulative life of mine capital expenditure including costs for pre-production and sustaining is

estimated to be $164.5M. The cumulative life of mine ‘Forecast to Spend’ amount, which includes

reclamation and closure bonding costs and excludes sunk costs, is estimated to be $158.7M.

Table 1-9 provides a summary of the capital costs.




JUNE 2020 1-23

Table 1-9: Project capital cost summary

Area Cost area description Pre-production

capital cost ($M) Sustaining

capital cost ($M) Total cost

($M)

2000 Administration and Services 0.5 0.0 0.5

3000 Mine 29.4 92.8 122.2

5000 Stockpiling and Conveying 0.1 0.0 0.1

6000 Processing Plant 2.4 1.2 3.6

7000 Tailings Storage Facility & Water Management

0.02 16.65 16.7

8000 Owner's Costs (8900 excluded) 3.4 0.6 4.0

9000 Project Indirect Costs (9800 excluded) 2.7 4.0 6.7

9800 Contingency 2.9 5.6 8.5

Pre-production Operating Costs 2.2 0.0 2.2

Total 43.7 120.8 164.5

Less Sunk Costs -8.9 - -8.9

8900 Site Reclamation and Closure 1.5 1.5 3.0

Total - Forecast to Spend 36.3 122.3 158.7

All capital costs for the Project have been distributed against the development schedule to support

the economic cash flow model. Figure 1-9 presents the planned annual and cumulative LOM

‘Forecast to Spend’ capital cost profile.




JUNE 2020 1-24

Figure 1-9: Annual and cumulative project capital costs

1.11.2 Operating Costs

The operating cost estimate (OPEX) is mainly based on historical costs provided by Wesdome, as

well as reference projects and studies as appropriate for a PEA study. The target accuracy of the

operating cost estimate is +/-35%. No cost escalation or contingency has been included within the

operating cost estimate.

The operating cost estimate in this Study includes the costs to mine and process the mineralized

material to produce gold doré. It also includes costs for general and administration expenses and

tailings management.

The average operating cost over the 8-year mine life is estimated to be $162.66/t mined. Table 1-10

provides a summary of the projected operating costs for the Kiena Mine Complex project.

0

20

40

60

80

100

120

140

160

180

0

10

20

30

40

50

60

2020 2021 2022 2023 2024 2025 2026 2027 2028

Cum

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tive C

apital C

osts

(m

illio

n C

AD

)

Annual C

apital C

osts

(m

illio

n C

AD

)

Pre-production Capital Costs Sustaining Capital Costs

Reclamation and Closure Costs Cumulative Capital Costs




JUNE 2020 1-25

Table 1-10: Operating costs summary

Cost area LOM ($M) Annual average

cost ($M) Average LOM

($/t mined) Average

LOM ($/oz) OPEX (%)

U/G Mining 214.1 26.8 104.15 315.7 64.0

Processing & Lab 48.6 6.1 23.62 71.6 14.5

Surface Operations 17.4 2.2 8.47 25.7 5.2

Technical Services 12.3 1.5 6.00 18.2 3.7

HSE & Training 11.5 1.4 5.61 17.0 3.4

Administration 24.8 3.1 12.07 36.6 7.4

Tailings Management (new facility)

5.6 0.7 2.75 8.3 1.7

Total 334.4 41.8 162.66 493.1 100.0

1.12 Project Economics

The economic/financial assessment of the Kiena Mine Complex project was carried out using a

discounted cash flow approach on a pre-tax and after-tax basis, based on independent long-term

projections for the gold price in United States currency (USD) and cost estimates capital

expenditure (CAPEX) and operating expenditure (OPEX) in Canadian (CAD or $) currency. Inflation

or cost escalation factors were not considered. The base case gold price is USD 1,532/oz, which

is the average price projection for the period 2020-2029 provided by CPM Group.

The economic analysis presented in this section contains forward-looking information about the

mineral resource estimates, commodity prices, exchange rates, proposed mine production plan,

projected recovery rates, operating costs, construction costs and project schedule. The results of

the economic analysis are subject to several known and unknown risks, uncertainties and other

factors that may cause actual results to differ materially from those presented here. The reader is

cautioned that this PEA is preliminary in nature and includes the use of Inferred mineral resources

that are considered too speculative geologically to have the economic considerations applied to

them that would enable them to be categorized as mineral reserves and, as such, there is no

certainty that the PEA economics will be realized.

The input parameters used, and results of the financial analysis are presented in Table 1-11 (all

amounts are in CAD unless noted otherwise).

On an after-tax basis, the base case financial model resulted in an IRR of 102% and an NPV of

$416.1M using a 5% discount rate. The after-tax payback period after start of operations is

1.7 years.

The pre-tax base case financial model resulted in an IRR of 126% and an NPV of $620.4M using

a 5% discount rate. The pre-tax payback period after start of operations is 1.6 years.




JUNE 2020 1-26

The all-in sustaining costs (AISC) over the LOM are USD 512/oz. Sunk costs are not included in

the financial analysis.

Table 1-11: Financial analysis summary

Description Unit Value

Total Tonnes Mined M tonne (Mt) 2.1

Average Diluted Gold Grade g/t 10.65

Total Gold Contained oz 709,065

Total Gold Payable oz 687,449

Average Annual Gold Produced Au oz per year 85,931

Total Preproduction Capital Cost $M 34.8

Sustaining Capital $M 120.8

Site Restoration Cost $M 3.0

Operating Costs $/t mined 162.66

All-in Sustaining Costs (AISC) USD/oz 512

Total LOM NSR Revenue $M 1376.4

Total LOM Operating Costs $M 334.4

Total LOM Pre-tax Cash Flow $M 883.3

Average Annual Pre-tax Cash Flow $M 110.4

LOM Royalties $M 0.0

LOM Mining Taxes $M 101.7

LOM Income Taxes $M 186.3

Total LOM After-tax Free Cash Flow $M 595.3

Average Annual After-tax Free Cash Flow $M 74.4

Pre-tax Summary

Pre-tax NPV (@ 5% Discount Rate) $M 620.4

Pre-tax IRR % 126

Pre-tax Payback (after start of operations) year 1.6

After-tax Summary

After-tax NPV (@ 5% Discount Rate) $M 416.1

After-tax IRR % 102

After-tax Payback (after start of operations) year 1.7




JUNE 2020 1-27

A financial sensitivity analysis was conducted on the Project’s after tax NPV and IRR using the

following variables: capital costs (pre-production and sustaining) operating costs, USD:CAD

exchange rate, and the price of gold.

The graphical representations of the financial sensitivity analysis on NPV and IRR are depicted in

Figure 1-10 and Figure 1-11. The sensitivity analysis reveals that the price of gold has the most

significant influence on both the NPV and IRR compared to the other parameters, based on the

range of values evaluated. After the price of gold, the NPV and IRR were most impacted by changes

in the USD:CAD exchange rate and then, to a lesser extent, by variations in operating costs and

capital costs. It should be noted that the economic viability of the Project will not be significantly

negatively impacted by variations in the capital or operating costs, within the margins of error

associated with the PEA cost estimates.

Overall, the NPV and IRR of the Project are positive over the range of values used for the sensitivity

analysis when analyzed individually.

Figure 1-10: Sensitivity of the net present value (after-tax) to financial variables

0

100

200

300

400

500

600

700

-40% -30% -20% -10% 0% 10% 20% 30% 40%

Aft

er-

Tax N

PV

@ 5

% (

CA

D-m

illio

ns)

% Change in Variable

Gold Price USD:CAD CAPEX OPEX




JUNE 2020 1-28

Figure 1-11: Sensitivity of the internal rate of return (after-tax) to financial variables

1.13 Interpretations and Conclusions

This PEA was prepared by a group of independent QPs to demonstrate the economic viability of

restarting operations at the Kiena Complex. This Report provides a summary of the results and

findings from each major area of investigation. Standard industry practices, equipment and

processes were used. To date, the QPs are not aware of any unusual or significant risks or

uncertainties that could materially affect the reliability or confidence in the Kiena Mine Complex

Gold project based on the information available.

The results of the Study indicate that the proposed Project has technical and financial merit using

the base case assumptions. It has also identified additional field work, metallurgical testwork, trade-

off studies and analysis required to support more advanced mining studies. The QPs consider the

PEA results sufficiently reliable and recommend that the Kiena Mine Complex Gold project be

advanced to next stage of development through the initiation of a prefeasibility study.

0%

20%

40%

60%

80%

100%

120%

140%

160%

180%

-40% -30% -20% -10% 0% 10% 20% 30% 40%

Aft

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R (

%)






JUNE 2020 1-29

An analysis of the results of the investigations has identified a series of risks and opportunities

associated with each of the technical aspects considered for the development of the Kiena Mine

Complex project. A preliminary rating of the risks listed in Chapter 25 indicates that the most

significant risks to the Project are the following:

▪ The existing underground backfill system may be found to be inadequate, not usable, or

could fail (due to piping system failure, piping or equipment plugging, etc.);

▪ The plant restart could take longer than anticipated, and additional costs would be incurred

to support extended restart efforts and provide for refurbishments that were not anticipated.

Opportunities also exist:

▪ The calculation of the mineable resource was based on a gold price of USD1,300 per oz.

Using a higher gold price would result in current resource below cut-off becoming

economical, thus increasing resource tonnes;

▪ The remaining capacity in the existing tailings storage facility may be greater than assumed

in this study, which may allow the delaying or even elimination of the new dry stack tailings

storage facility.

1.14 Recommendations

The QPs recommend that the Project proceed to the prefeasibility study phase. It is also

recommended that environmental and permitting continue as needed to support the Project’s

development plans and Project schedule.

An extensive work program including additional exploration drilling and the prefeasibility study has

been developed based on QP recommendations. The work program is estimated to cost

approximately $7.6M including a $1.2M contingency.




JUNE 2020 2-1

INTRODUCTION

This Report was prepared and compiled by BBA Inc. (BBA) at the request of Wesdome Gold Mines

Ltd. (Wesdome or the "Company"). The purpose of this Report is to summarize the results of the

Preliminary Economic Assessment (PEA) of the Kiena Mine Complex project (the "Project") in

accordance with the guidelines of the Canadian Securities Administrators National Instrument

43-101 (NI 43-101) and Form 43-101F1.

BBA is an independent engineering consulting firm headquartered in Mont-Saint-Hilaire, Quebec,

with mining groups based in Montreal, Vancouver, Toronto, Val-d’Or, and Sudbury. The Toronto

team led the overall integration of this Report.

2.1 Wesdome Gold Mines Ltd.

Wesdome Gold Mines Ltd. is a mining, exploration and development company based in Toronto,

Ontario, focused primarily on gold. The Company has three ongoing projects in Canada: the Eagle

River Complex in Wawa, Ontario, with two deposits and a processing plant currently operating; the

Moss Lake deposit in Thunder Bay, Ontario, at the exploration stage; and the Kiena Mine Complex

in Val-d’Or, Quebec. The Kiena Mine Complex is an underground mine that has been on care and

maintenance since mid-2013, with a 2,000 tpd mill and a 930 m shaft. A total of 1.75 million ounces

of gold has been produced at the Kiena Mine Complex since 1981. Following new gold discoveries,

Wesdome is looking to restart operations at the Kiena Mine Complex.

Wesdome has mandated BBA to manage the Kiena Mine Complex project PEA.

2.2 Basis of Technical Report

The following Report presents the results of the PEA for the development of the Kiena Mine

Complex project. As of the date of this Report, Wesdome is a Canadian gold producer trading on

the TSX under the trading symbol (WDO), with its head office situated at:

220 Bay St, Suite 1200

Toronto, Ontario

M5J 2W4

This Report, titled “Preliminary Economic Assessment for the Kiena Mine Complex Project” (the

"Report"), was prepared by Qualified Persons (QPs) following the guidelines of the NI 43-101 and

in conformity with the guidelines of the Canadian Institute of Mining, Metallurgy and Petroleum

(CIM) Standards on Mineral Resources and Reserves.




JUNE 2020 2-2

2.3 Report Responsibility and Qualified Persons

The following individuals, by virtue of their education, experience and professional association, are

considered Qualified Persons as defined in the NI 43-101, and are members in good standing of

appropriate professional institutions.

▪ Colin Hardie, P. Eng. BBA Inc.

▪ Pierre-Luc Richard, P. Geo. BBA Inc.

▪ Jorge Torrealba, P. Eng. BBA Inc.

▪ Tom Corkal, P. Eng. BBA Inc.

▪ Luciano Piciacchia, P. Eng. BBA Inc.

The preceding QPs have contributed to the writing of this Report and have provided QP certificates,

included at the beginning of this Report. The information contained in the certificates outlines the

sections in this Report for which each QP is responsible. Each QP has approved the contributed

figures, tables and portions of Chapters 1 (Summary), 25 (Interpretation and Conclusions), and

26 (Recommendations). Table 2-1 outlines the responsibilities for the various sections of the Report

and the name of the corresponding Qualified Person.

Table 2-1: Qualified Persons and areas of report responsibility

Chapter Description Qualified Person

Company Comments and exceptions

1. Executive Summary C. Hardie BBA All QPs contributed based on their respective scope of work and the Chapters/Sections under their responsibility.

2. Introduction C. Hardie BBA All QPs contributed based on their respective scope of work and the Chapters/Sections under their responsibility.

3. Reliance on other Experts C. Hardie BBA All QPs contributed based on their respective scope of work and the Chapters/Sections under their responsibility.

4. Project Property Description and Location

P-L. Richard BBA All Chapter 4

5. Accessibility, Climate, Local Resource, Infrastructure and Physiography


6. History P-L. Richard BBA All Chapter 6

7. Geological Setting and Mineralization P-L. Richard BBA All Chapter 7

8. Deposit Types P-L. Richard BBA All Chapter 8

9. Exploration P-L. Richard BBA All Chapter 9

10. Drilling P-L. Richard BBA All Chapter 10

11. Sample Preparation, Analyses and Security





JUNE 2020 2-3

Chapter Description Qualified Person

Company Comments and exceptions

12. Data Verification P-L. Richard BBA All Chapter 12

13. Mineral Processing and Metallurgical Testing

J. Torrealba BBA All Chapter 13

14. Mineral Resource Estimate P-L. Richard BBA All Chapter 14

15. Mineral Reserve Estimate C. Hardie BBA All Chapter 15

16. Mining Methods T. Corkal BBA All Chapter 16

17. Recovery Methods J. Torrealba BBA All Chapter 17

18. Project Infrastructure L. Piciacchia BBA All Chapter 18

19. Market Studies and Contracts C. Hardie BBA All Chapter 19

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

C. Hardie BBA All Chapter 20

21. Capital and Operating Costs C. Hardie BBA All Chapter 21 except mining capital and operating costs, under the responsibility of T. Corkal, QP

22. Economic Analysis C. Hardie BBA All Chapter 22

23. Adjacent Properties P-L. Richard BBA All Chapter 23

24. Other Relevant Data and Information C. Hardie BBA All Chapter 24

25. Interpretation and Conclusions C. Hardie BBA All QPs contributed based on their respective scope of work and the Chapters/Sections under their responsibility.

26. Recommendations C. Hardie BBA All QPs contributed based on their respective scope of work and the Chapters/Sections under their responsibility.

27. References C. Hardie BBA All QPs contributed based on their respective scope of work and the Chapters/Sections under their responsibility.

2.4 Effective Dates and Declaration

The Report has several cut-off dates for information:

▪ Effective date of the Kiena Mine Complex project Mineral Resource Estimate used as the

basis for the LOM Plan: September 25, 2019;

▪ Date of last supply of laboratory testwork and investigations: September 2019;

▪ Date of the financial analysis: March 31, 2020.




JUNE 2020 2-4

This Report was prepared as National Instrument 43-101 (NI 43-101) Technical Report for

Wesdome Gold Mines Ltd. by QPs from BBA Inc., 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 Wesdome subject to the

terms and conditions of its respective contracts with the Report Authors. Except for the purposes

legislated under Canadian provincial and territorial securities law, any other use of this Report by

any third party is at the sole risk of that party.

As of the effective date of this Report, the QPs are not aware of any known litigation potentially

affecting the Project. The QPs did not verify the legality or terms of any underlying agreement(s)

that may exist concerning the Project ownership, permits, off-take agreements, license agreements,

royalties or other agreement(s) between Wesdome and any third parties.

The results of this Report are not dependent upon prior agreements concerning the conclusions to

be reached, nor are there any undisclosed understandings concerning any future business dealings

with Wesdome and the QPs. The QPs are being paid a fee for their work in accordance with the

normal professional consulting practice.

The opinions contained herein are based on information collected throughout the course of the

investigations by the QPs, which in turn reflect various technical and economic conditions at the

time of writing. Given the nature of the mining business, these conditions can change significantly

over relatively short periods of time. Consequently, actual results can be significantly more or less

favourable.

2.5 Currency, Units of Measure and Calculations

Unless otherwise specified or noted, the units used in this Report are metric. Every effort has been

made to clearly display the appropriate units being used throughout this Report.

▪ Currency is in Canadian dollars (CAD or $), unless otherwise noted;

▪ All ounce units are reported in troy ounces, unless otherwise stated:

1 oz (troy) = 31.1 g = 1.1 oz (Imperial);

▪ All metal prices are expressed in US dollars (USD);

▪ A Canadian dollar (CAD) to United States dollar (USD) exchange rate of 0.76 USD for

1.00 CAD was used;

▪ All cost estimates have a base date of the fourth quarter (Q4) of 2019.

This Report includes technical information that required subsequent calculations to derive

subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding

and consequently introduce a margin of error. Where these occur, the QPs consider them

immaterial.




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2.6 Sources of Information

2.6.1 General

This Report is based in part on internal company reports, maps, published government reports,

company letters and memoranda, and public information, as listed in Chapter 27 “References” of

this Report.

This PEA has been completed using available information contained in, but not limited to, the

following reports, documents and discussions:

▪ Technical discussions and emails with Wesdome personnel;

▪ QPs’ personal inspection of the Project site(s);

▪ BBA project team members inspection of the Project site(s);

▪ Report of mineralogical, metallurgical and grindability characteristics of the Kiena Mine

Complex deposits, conducted by industry recognized metallurgical testing laboratories on

behalf of Wesdome;

▪ Internal and commercially available databases and cost models;

▪ Various reports covering site hydrology, hydrogeology, geotechnical, and geochemistry;

▪ Various reports covering site physical and biological environment;

▪ Additional information from public domain sources.

The QPs have no known reason to believe that any of the information used to prepare this Report

and evaluate the mineral resources presented herein is invalid or contains misrepresentations.

The authors have sourced the information for this Report from the collection of documents listed in

Chapter 27 (References).

2.6.2 BBA

The following individuals provided specialist input to Colin Hardie, QP:

▪ Frank Gagnon, Environment Coordinator of Wesdome at the Kiena Mine Complex, and

Patrick Frenette, Project Superintendent, Wesdome, provided information regarding the

Project’s environmental status, permits, and Social and Community Impact;

▪ Patrick Frenette, Project Superintendent, Wesdome, provided historical operating costs as

well as project capital costs used in the development of the Project’s cost estimates found in

Chapter 21 (Capital and Operating Costs);

▪ Jocelyn Marcoux, BBA, provided the industrial standards, norms and factors for the various

material, manpower and construction costs used in the development of the process plant

capital costs (Chapter 21).




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The following individuals provided specialist input to Pierre-Luc Richard, QP:

▪ Karine Brousseau, Senior Engineer - Mineral Resources, Wesdome, provided the 3D

modeling and block models to be audited and described in Chapter 14 (Mineral Resource

Estimate);

▪ Bruno Turcotte, Senior Project Geologist, Wesdome, provided information about the

polygonal Mineral Resource Estimate to be audited and described in Chapter 14 (Mineral

Resource Estimate);

▪ Patrick Frenette, Project Superintendent, Wesdome, provided historical operating costs as

well as project capital costs used in the development of the cut-off grade found in Chapter 14

(Mineral Resource Estimate).

The following individuals provided specialist input to Tom Corkal, QP:

▪ Wesdome staff and other experts retained by Wesdome supplied factual data on as-built

mine openings, existing mine infrastructure capacity, design and operational status and the

geological models;

▪ John Henning, BBA, provided expertise with respect to geomechanical work;

▪ Tim Paquin, BBA, provided expertise with respect to the mine ventilation.

The following individuals provided specialist input to Jorge Torrealba, QP:

▪ Patrick Frenette, Project Superintendent, Wesdome, provided historical information such as

process design criteria, flowsheets, equipment lists, operating costs as well as project capital

costs used in the development of the Project’s mass and water balances, Chapter 17 and

Chapter 21 (Recovery Methods, and Capital and Operating Costs, respectively);

▪ Tom Corkal, BBA, provided the LOM throughput schedule and backfill capacity for the

Project used in the mass and water balance (Chapter 17);

▪ Yves Thomassin, BBA, provided factors used in the water balance (Chapter 17).

The following individuals provided specialist input to Luciano Piciacchia, QP:

▪ Stantec provided the background information and data to perform a conceptual design of the

tailings management infrastructure. The following information have been reviewed:

- PowerPoint presentation from Stantec "Complexe Kiena, Gestion intégrée des résidus

miniers – Différence de conception entre les deux parcs";

- Avis au MELCC concernant l’emplacement de la nouvelle aire d’accumulation au

Complexe Kiena afin de réaliser une codéposition des stériles et des résidus miniers,

Stantec, mars 2019;

- Geotechnical study of the Kiena Complex, Val-d’Or, Québec, Stantec, September 2018;




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- Supplementary Geotechnical Investigation – Eastern Area of the Kiena Complex, Val-

d’Or, Québec, Stantec, November 2019;

- Communications on January 20, 2020 between Fortunato Copola (Stantec) and Yves

Thomassin (BBA).

These specialists are not considered as QPs for the purposes of this NI 43-101 Report.

2.7 Site Visits

The following bulleted list describes which Qualified Persons visited the Kiena Mine Complex site,

the date of the visit, and the general objective of the visit:

▪ Pierre-Luc Richard (BBA) has visited the Kiena Property that is the subject of this Report on

August 6 to 8, 2019, and on different other occasions as part of the current mandate. The

purpose of the visits was to review the Project with the Wesdome team. The visits included

an overview of the general geological conditions, a tour of the core storage facility, visual

inspections of select mineralized drill core samples and an underground visit of the site. A

review of assaying, QA/QC and drillhole procedures was also completed.

▪ Jorge Torrealba (BBA) has visited the Kiena Property that is the subject of this Report on

August 6 and 7, 2019. The purpose of the visit was to review the Kiena Mine Complex

project with Kiena’s superintendent (Patrick Frenette) and former mill superintendent

(Bernard Belley). The visit included an overview of the operation’s history, process operation,

a tour of the general site (from grinding area to the backfill plant) and a visit to the core shack

where samples were selected by J. Torrealba (with the help of BBA’s geologist at site,

Charlotte Athurion) for the 2019 metallurgical testwork program.

▪ Tom Corkal (BBA) has visited the Kiena Property that is the subject of this Report on

October 2 and 3, 2019. The objective of the visit was to inspect the surface facilities

associated with mining, tour the underground mine, and gather technical data on the

underground mine.

As of the effective date of this Report, the following QPs have not visited the Project site(s):

▪ Colin Hardie (BBA). A site visit was not deemed necessary for the purposes of supporting

Mr. Hardie’s responsibility as a Qualified Person or for the technical report integration

activities.

▪ Luciano Piciacchia (BBA). A site visit was not deemed necessary for the purposes of

supporting Mr. Piciacchia’s responsibility as a Qualified Person or for the technical report

integration activities.




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2.8 Acknowledgement

BBA and the other contributors to the study would like to acknowledge the following individuals for

their general support provided during this assignment:

The Project benefitted from the specific input of Marc-André Pelletier, Patrick Frenette, Ben Au,

Frank Gagnon, Sean McCormak, Michael Michaud, John Henning, Navin Gangadin, Tim Paquin,

Antonio Vides, Yves Thomassin, Hakan Tunc, Harold Bon, Wilbur Wong, Jocelyn Marcoux, Sylvain

Boily, Stéphane Landry, David Dobney and Manon Dussault. Their contributions are greatly

appreciated.




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RELIANCE ON OTHER EXPERTS

3.1 Introduction

The Qualified Persons (QPs) relied on reports, information sources and opinions provided by

Wesdome and external experts for certain aspects of the Project, such as the Project’s mineral

rights, 3rd party agreements, surface rights, property agreements, royalties, environmental status,

and fiscal situation.

As of the date of this Report, Wesdome indicates that there are no known litigations potentially

affecting the Kiena Mine Complex project.

A draft copy of the Report has been reviewed for factual errors by Wesdome. Any changes made

as a result of these reviews did not involve any alteration to the conclusions made. Hence, the

statements and opinions expressed in this document are given in good faith and in the belief that

such statements and opinions are neither false nor misleading at the date of this Report.

3.2 Mineral Tenure and Surface Rights

Wesdome supplied information regarding mining titles, options’ agreements, royalty agreements,

environmental liabilities and permits. Pierre-Luc Richard, QP, consulted the GESTIM online claim

management system via:

https://gestim.mines.gouv.qc.ca/MRN_GestimP_Presentation/ODM02101_login.aspx

for the latest status regarding ownership and mining titles. Although Mr. Richard has reviewed the

option agreements and available claim status documents, he is not qualified to express any legal

opinion with respect to the property titles, current ownership, or possible litigations. A description

of such agreements, the property, and ownership thereof, is provided for general information

purposes only. In this regard, Mr. Richard has relied on information supplied by Wesdome and the

work of experts they understand to be appropriately qualified.

This information is used in Chapter 4 of the Report. The information is also used in support of the

Mineral Resource Estimate in Chapter 14.

3.3 Taxation

Colin Hardie, QP, has fully relied upon, and disclaims responsibility for, information supplied by

Wesdome staff and experts retained by Wesdome for information related to taxation as applied to

the financial model. This information is used in support of the financial analysis in Chapter 22.

https://gestim.mines.gouv.qc.ca/MRN_GestimP_Presentation/ODM02101_login.aspx




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3.4 Commodity Price Projections

Colin Hardie, QP, has relied upon independent gold price projections supplied by CPM Group

(CPM, 2020) to prepare the financial model and project valuation. This information is used in

support of the financial analysis in Chapter 22.

CPM Group (CPM) is a commodities research, consulting, financial advisory and commodities

management firm providing independent research, analysis and advisory services related to

commodities markets, corporate and project finance, and the financial management of exposure to

commodity-oriented investments.

Analysts from CPM have been studying the gold market, along with other commodities markets,

since 1980. CPM has provided independent gold price projections for use in several projects for

companies around the world, such as preliminary economic assessments, prefeasibility studies,

bankable full feasibility studies, prospectuses for initial public offerings, secondary offerings, rights

offerings, and other applications. CPM's clients include Anglogold, Ashanti Goldfields, Anglogold

Ashanti, Barrick, Newmont, Normandy, Placer Dome, Penoles, Luismin, Goldcorp, Lonmin, Norilsk

Nickel, Novagold, Wheaton Precious Metals, and others. In addition to mining companies, CPM’s

long-term price projections are used by institutional investors, banks and financial institutions,

central banks and monetary authorities, intergovernmental organizations such as the United

Nations, World Bank, and International Monetary Fund, and others.




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PROPERTY DESCRIPTION AND LOCATION

4.1 Property Description and Location

The Kiena Mine Complex project is located in the Province of Quebec in the Abitibi-

Temiscamingue administrative region within the limits of the municipality of Val-d’Or and 100 km

east of Rouyn-Noranda (Figure 4-1). It lies to the northwest of the urban centre of Val-d’Or and

covers 7,047 Ha.

The coordinates for the approximate centre of the Project are latitude 48°08' N and longitude

77°54' W (284105E and 5335715N: NAD 83 / UTM Zone 18). The Project lies in the townships of

Dubuisson and Vassan on NTS map sheets 32D/01 and 32C/04.




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Figure 4-1: Property overview map of the Kiena Mine Complex




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4.2 Mineral Tenure

Pierre-Luc Richard verified the status of the mineral claims using the Quebec government online

claim management tool GESTIM. As of October 29, 2019, the Kiena Mine Complex property

consists of a contiguous group of 183 contiguous electronic map designated mining claims and

one mining concession (Figure 4-2).

From the 183 mining claims, Wesdome has 169 claims registered in their name at 100%. Six

claims are held be Mines Dynacor (50%) and Wesdome (50%), which represents the Maufort

property. Eight claims are held by Wesdome (75%) and 9264-7890 Québec inc. (25%), which

represents the Siscoe Extension property.

The Project is located in the Dubuisson and Vassan Township, near the town of Val-d’Or in the

Province of Quebec. The total area of the Project is 7,047 hectares. A detailed list of the Project

mineral claims is shown in Appendix A.

Figure 4-2: Kiena Mine Complex titles as of October 29, 2019




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4.3 Royalties, Agreement and Encumbrances

Some of the mining titles comprising the Project are subject to certain agreements and royalties.

Figure 4-3 shows the historical properties that were amalgamated to form the current Project,

some of which have active royalties. Table 4-1 provides the details of these royalties.




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Figure 4-3: Kiena Mine Complex mining royalties and historical property names




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Table 4-1: List of historical properties with their applicable royalties

Name Royalties Date of agreement Details

Audet Block 2% NSR to Huguette Audet N/A

Callahan 1% NSR to Placer Dome (CLA) Ltd. N/A

Dubuisson No royalty

Elmac 2% NOP to Albert Audet and Daniel Audet N/A

Kiena No Royalty

Kiena South Block

No Royalty

Kiena West 1% NSR to Jack Stoch N/A

Lac De Montigny No Royalty

Lac Dubuisson No Royalty

Lamothe 1% NSR to Robert Lamothe January 15, 1998

Maufort 10% NPR to Charlim Exploration Inc. N/A

Rosenbaum 2% NSR to Pierre-André Bigué N/A 1% can be bought back for $1,000,000

Roy Option CAD0.25 per tonne of ore milled to Marie-Louis Roy (6 of 8 claims)

N/A

School Mine No Royalty

Shawkey No Royalty

Shawkey South 1% NSR to Léo Audet N/A

Siscoe and Siscoe Extension

3% NSR to Dynacor Mines Inc. November 9, 1999 1% can be bought back for $500,000

0.5% NSR to Demontigny Resources Inc. N/A 0.5% can be bought for $500,000

Vassan No Royalty

Wesdome 1% NSR to Dome Mines Ltd. November 21, 1997

Yankee Clipper 2% NPR to Jacques Duval and Kenneth Alexander Wheeler

February 25, 1981

4.4 Environmental Liabilities

Several studies were conducted before and during operations on portions of the Project related to

former operations and other projects within its boundaries. The Project includes the milling and

tailings facilities of the Kiena mine, nine shafts including Parker Shaft, and related underground

workings from past producers and exploration projects (Figure 4-4), and various surface facilities.

The remaining infrastructure components at the former Kiena mine site now constitute the

principal infrastructure of the Project; these components are listed below:

▪ The former collar of the Parker Shaft is located close to the actual mill but is not accessible;

▪ A functional shaft of 930 m depth;




JUNE 2020 4-7

▪ A dewatered underground ramp system from 170 m (level 17) to 1,090 m (level 109);

▪ A 2,000-tpd processing plant;

▪ A tailing management facility;

▪ Core shacks;

▪ Offices;

▪ A guard house and security gate;

▪ Workshops and warehouses;

▪ A Hazmat storage facility.

Other than the exploration offices and underground exploration development, the principal

infrastructure of the Project has been under care and maintenance since mid-2013. Effluent is

being controlled and analytical results are submitted to federal and provincial authorities who

have determined that the effluent, based on the available data, complies with regulation

requirements. A recent study conducted by Stantec (Stantec, 2017) to assess the chemical

behaviour of the waste rock pile concluded that the waste rocks are neither acid generating nor

do they have metal leaching potential under the conditions present at the time of the study.

The environmental permits and the social acceptability of the Project, regardless of current status,

will need to be reviewed should any operations resume. Moreover, both federal and provincial

authorities have amended their regulations since 2013.

Any mining related work would be subject to environmental studies and permit applications.

Access is easy as the Project lies within an inhabited area and some data may already have been

collected through previous studies. The Project is located near a lake since the 1980s and

Wesdome’s representatives mentioned to the QP that there has been no social resistance or

counteraction.

The last version of the Kiena Closure and Rehabilitation Plan was accepted by the MERN in

September 2015. The total cost was estimated at $7.0M. The financial guarantee was completed

on September 28, 2017.

In the summer of 2018, a geotechnical study (Stantec, 2018) was conducted to investigate the

subsurface conditions of the area surrounding the Kiena waste rock pile. Stratigraphic profiles,

cone penetration tests, shear resistance profiles, test pits, and boreholes were done for the

characterization.

As of December 2018, Wesdome has kept the municipal authorities informed on their mining and

exploration activities but no official consultation has been conducted with other interested

stakeholders. The main acknowledged concern is groundwater quality as it is the source of

drinking water for nearby dwellings. Wesdome continued to monitor final effluent, surface water

and groundwater wells after cessation of operations in 2013.




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Figure 4-4: Surface projection of the Kiena Mine Complex underground work, production and exploration shafts as of end of 2018




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4.5 Permitting

Wesdome has various authorization certificates for their property. Table 4-2 shows the list of

certificates they currently hold with the conditions to respect. All certificates of authorization were

provided by Frank Gagnon, Environment Coordinator from Wesdome at the Kiena Mine Complex,

and are valid and active according to Mr. Gagnon.




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Table 4-2: Wesdome certificates of authorization details

Certificate of authorization number

Title Type Date Conditions

7610-08-01-70065-23

Reopening of "Les mines d'or Kiena Ltée"

Authorization certificate 1981-07-29

▪ Extraction of 1,000 tpd of ore;

▪ Truck transportation to the Lamaque Mill;

▪ Hydraulic backfill with the Lamaque Mill waste;

▪ Enlargement of the island with the waste rock;

▪ Disposal of the waste rock in a sequence;

▪ Dust control of the access roads and unpaved traffic areas;

▪ Final effluent monitoring according to the Directive 019 with additionally cadmium at 0.05 mg/l and mercury at 0.001 mg/l;

▪ All mitigation measures of the:

▪ Environmental Impact Assessments (Nov. 21, 1980):

▪ Letter of January 14, 1981;

▪ Letter of May 27, 1981.

Cession of the authorization certificate from "Les mines d'or

Kiena Ltée" to "Mc Watters" 1997-09-12

▪ Same conditions as the authorization certificate of July 29, 1981.

Modification 2001-02-23 ▪ Extraction change of 1,000 tpd to 2,040 tpd of ore.

Cession of the authorization certificate from "Mc Watters" to

"Wesdome inc." 2004-05-25 ▪ Extraction of 2,040 tpd of ore.

Expansion of the waste rock pile

Modification 2004-12-20

▪ From 160,000 t to 418,000 t;

▪ On lots 24B and 25B, road A, Dubuisson township;

▪ Extraction of 10,000 t per month of waste rock;

▪ Kiena Mine waste rock and development of the ramp towards Shawkey and Wesdome;

▪ Waste rock will be haul by truck towards the waste pile;

▪ Actual waste pile of 160,000 t, 11,800 m2 with a height of 8 m;




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▪ Temporary extension of the waste pile to 258,000 t for 31,357 m2 at 4.5 m high;

▪ Some waste rock will be used for raising the tailings pond, building berms and roads;

▪ Slope of 2H/1V.

7610-08-01-70065-24 Operation of the ore plant

at the Kiena mine Authorization certificate 1984-08-02

▪ Located on lots 19B to 25B, road B, Dubuisson township;

▪ Treatment plant of 1,250 tpd;

▪ Direct cyanidation and activated carbon recovery;

▪ Hydrocarbon sensor installation where they are likely to be present;

▪ Installation of leak proof structures to store a volume equivalent to the largest tank of the same group for all tanks of toxic and dangerous products;

▪ Installation of a dry or wet scrubber to capture the emissions of gas or dust generated by the treatment operations;

▪ Installation of equipment to eliminate or reduce dust emissions from storage, disposal, transhipment and transportation areas in accordance with the Regulation on the Quality of the Atmosphere;

▪ Construction of a tailings transportation line from the plant to the park;

▪ Development of a tailings pond;

▪ Construction of a polishing pond;

▪ Digging of diversion channels for drainage water at the periphery of the tailings pond;

▪ Installation of a continuous recording system for pH and flow;

▪ Cyanide treatment at the mill effluent;

▪ Effluent monitoring according to the standards established in the authorization certificate.




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Modification of the authorization certificate of

Aug. 2, 1984 for the tailings pond construction


▪ Lots 19B to 25B, road A, Dubuisson township;

▪ Use of the waste rocks from "Malartic Goldfields" to rise the pond;

▪ Installation of a geotextile into the structure of the pond embankment.

Modification at the final effluent standards

Modification 1989-04-04 ▪ Modification of the effluent discharge standards for the

tailings pond pursuant to guidelines from Directive 019.

Operation of the ore plant at the Kiena mine

Modification 1996-11-27 ▪ Addition of two polishing ponds of 32,500 m2 at the south

end of the current pond.


Kiena Ltée" to "Mc Watters" 1997-09-12

▪ All the conditions of the authorization certificate remain unchanged.

Modification 2001-03-19 ▪ Change of 1,250 tpd to the treatment of 2,040 tpd.


"Wesdome inc." 2004-05-25


7610-08-01-70065-28 Tailings pond expansion


▪ Natural degradation of the cyanide with the SO2-Air process;

▪ Technical description of the cyanide content of the hydraulic backfill;

▪ Installation of five piezometers at the tailings pond (P-6 to P-10) and a control piezometer for background (P-11);

▪ Commitment to comply with Directive 019;

▪ Regular inspections of embankments and pipes;

▪ Pumping of 125,000 m3 per year of fresh water from Demontigny Lake;

▪ The overburden will be kept.

Cession of the authorization certificate from ''Les mines d'or Kiena Ltée.'' to ''Mc Watters.''

1997-09-12 ▪ All the conditions of the authorization certificate remain

unchanged.




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"Wesdome inc."

2004-05-25 ▪ All the conditions of the authorization certificate remain

unchanged.


▪ Upgrading of the northern cell of the tailings pond from elevation 3,048 m to 3,050 m;

▪ Lots 18B to 22B, road A, Dubuisson township.

7610-08-01-70065-29 Clay quarry exploitation


▪ Exploitation area of 40,000 m2, with an average thickness of 1.5 m and maximum of 2 m;

▪ Lots 21B, 22B and 23B, rang A, Dubuisson township.


Kiena ltée" to ''Mc Watters" 1997-09-12





7610-08-01-70065-33 Clay borrow pit exploitation


▪ Topsoil will be stored on the edge of the exploitation for use during restoration (8,550 m3);

▪ Exploitation above the water table.



▪ Exploitation of 12,000 m3 of clay on an area of 29,000 m2, an average thickness of 1 m and maximum of 2 m on lots 18B half East and half West, road A, Dubuisson township.

7610-08-01-70065-00 Septic system Authorization certificate 1981-02-10

▪ For the services building with a maximum capacity of 135 persons;

▪ Installation of the septic system with a capacity of 5,744 US gallons;

▪ Installation of a pumping station with two pumps;

▪ Installation of a leaching field with an area of 3,200 ft2.




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ACCESSIBILITY, CLIMATE, LOCAL RESOURCES AND INFRASTRUCTURE, PHYSIOGRAPHY

5.1 Accessibility

The Project is easily accessible via paved highways from local communities, such as Val-d’Or.

The Project is located 10 km west of the Val-d’Or Township (Figure 5-1) along the provincial

Highway 117 and turning north on the Chemin Kienawisik gravel road. The property can be

accessed and operated on a year-round basis. The nearest airport with daily flights connecting

Montreal is in Val-d’Or.

Figure 5-1: Location of the Kiena Mine Complex Property




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Figure 5-2: Gravel road leading to the property with an overview of the infrastructure

(© Jean-Philippe Richard, 2016)

5.2 Climate

The Val-d’Or area experiences a continental subarctic subhumid climate, characterized by short,

cool summers and long, cold winters. The nearest permanent weather monitoring station

maintained by Environment Canada (climat.meteo.gc.ca) is the Amos station, approximately

50 km north of the Property. According to the available data collected from 1981-2010 at this

weather station, the daily average temperature for January was -17.2°C and the daily average

temperature in July was 17.4°C. The record low during this period was -52.8°C, and the record

high was 37.2°C.

Data collected from the Amos weather station from 1981 to 2010 indicates that the total annual

precipitation was 929.0 mm, with peak rainfall occurring during July (112.1 mm average), August

(98.3 mm average) and September (106.7 mm average). Snowfall is light to moderate, with an

annual average of 253.3 cm. Snow typically accumulates from October to April, with a peak

snowfall occurring in November (45.0 cm average), December (58.5 cm average) and January

(55.6 cm average); during this period, snowpack averages 39 cm depth, with a maximum depth of

approximately 142 cm. On average, the Property is frost-free for 97 days, though discontinuous

permafrost exists in the area. Hours of sunlight vary from 15.5 hours at the summer solstice in

June to 8.1 hours at the winter solstice in December.




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The climatic conditions at the Property do not significantly impede the Project or hinder

exploration or mining activities, beyond seasonal considerations for certain work. The operating

season is year-round since the majority of the work is underground. The parts of the Project

located on islands are accessible by boat during the summer season and by ice bridges in winter.

5.3 Local Resources and Infrastructure

5.3.1 Airports, Rail Terminals and Bus Services

The town of Val-d’Or, with a population of approximately 32,900 residents, is located 10 km east

of the Project along the provincial Highway 117. Val-d’Or has been a mining service centre since

its foundation in the 1920s. Val-d’Or is one of the largest communities in the region and has all

major services including an airport with scheduled service from Montreal. CN railway line crosses

the southern part of the property, connecting east through to Montreal and west to the North

American rail network. Val-d’Or is a six-hour drive from Montreal, and there are daily bus services

between Montreal and the other cities in the Abitibi-Temiscamingue region.

5.3.2 Local Work Force

According to the 2016 census prepared by Statistics Canada, the population of the MRC of La

Vallée-de-l’Or was 43,226 people, with 66% of the residents aged 15-64, and an average of 41

years old. Male population accounts for 51% of the population, 49% is female, and 8.5% is

Aboriginal. In 2016, 64.4% of the population participated in the labour force, with 14.2% of the

labour force employed in mining, quarrying, and oil and gas extraction. This portion of the

workforce is experienced in mining operations, as they are currently employed at exploration and

gold mines located elsewhere in the region. Local resources also include commercial

laboratories, drilling companies, exploration service companies, engineering consultants,

construction contractors and equipment suppliers.

5.3.3 Additional Support Services

Additional services within the town of Val-d’Or include the Val-d’Or Hospital, grocery stores, fuel

stations, financial institutions and hotels. Val-d’Or has a Canada Post office and additional

shipping/freight services by several providers. Landline telephone, mobile service, high-speed

internet and satellite internet are available in town and the nearby vicinity.




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5.4 Physiography

The Project is located in the Abitibi Greenstone Belt of the Superior Province. The topography of

the area is characterized by low ridges and hills flanked by generally flat areas of glacial outwash,

swamps and numerous lakes and bogs. The average elevation varies from 300 masl to 320 masl

with some areas in the southern part of the project rising to 360 masl. Overburden varies between

0 m and 15 m and consists of stratified clays as well as glacial and glaciofluvial Pleistocene

deposits (MDDELCC). Most of the project is covered by Lac De Montigny, which includes several

islands.

The Project is located in the southern part of the boreal bioclimatic domain. The dominant

vegetation is fir and white spruce forest, with occurrences of white birch, black spruce, jack pine,

larch and aspen (MFFP website).

Two sources of water were used at the Project: surface water from Lac De Montigny and an

underground source from level 17 in the Kiena mine at a vertical depth of 170 m. Surface water is

used in case of emergencies (e.g., fire). Two large pumps are located in a building behind the

plant. Underground water flows naturally on level 17. A pumping system was installed to bring

water to the surface at the pumping facilities near the No.1 shaft. This water is used as clean

water for showers and toilets. A small storage tank (holding water from level 17) is also located on

level 38 for industrial underground use.

Electricity is available from Hydro-Québec through an above-ground power line. This source of

power is used to heat the surface buildings and run the mill. A generator is also available in case

of a power outage to run the emergency lighting, the underground pumps and the silo truck at the

surface. A 10,000 L capacity diesel tank is located near the generator, linked to a tank with a

capacity of 1,135 L. Four diesel tanks are located underground at mine levels 27, 33, 48 and 64.

These tanks have a capacity of 4,500 L.

A Natural Gas line is also used to feed the mill, shaft, warehouse, backfill plant and air heating of

the vent raise.

5.5 Mine Infrastructure

The Project is an amalgamation of 20 former properties. Figure 4-4 shows all underground

development for the Project. Table 5-1 presents a summary of the remaining infrastructure in

each part (former site) of the Kiena Mine Complex.




JUNE 2020 5-5

Table 5-1: Kiena Mine Complex infrastructure

Historical mine Underground infrastructure Surface infrastructure

Kiena Mine ▪ Access to underground work is through the No.1 shaft to a depth of 930 m. It provides access from level 12 (120 m below surface) to level 93 (930 m below surface). Levels 17 to 109 are accessible by a ramp that extends from 170 m to 1,090 m below surface. A total of 49 levels was excavated. A northwest exploration drift of 947 m is present on level 12. A second exploration drift of about 3,500 m extends to the east on level 33.

▪ The shaft and head frame;

▪ A dry facility;

▪ A laboratory;

▪ A core shack;

▪ A hoisting room;

▪ Main electric substation (4 transformers);

▪ A secondary electrical substation (25 kV);

▪ Fuel reservoirs.

Wisik ▪ N/A ▪ Concrete slab and old foundation walls.

Shawkey Mine Shaft No. 1 Area

▪ A shaft is connected to 7 levels of underground drifts and raises at levels 125 ft (38 m), 225 ft (69 m), 325 ft (99 m), 450 ft (137 m), 575 ft (175 m), 625 ft (191 m) and 725 ft (221 m) below surface. The drifts amount to approximately 1,125 ft (343 m);

▪ Plan views for the first five levels are available.

▪ Concrete slab and old foundation walls.

Shawkey Mine Shaft No.2 Area

▪ A 743 ft (227 m) shaft provides access to four levels of drifting at 250 ft (76 m), 400 ft (122 m), 550 ft (168 m) and 700 ft (213 m) below the surface. Drifting and cross-cutting amounted to approximately

927 ft (283 m);

▪ Total underground lateral development in the form of drives, cross-cuts and drifts amounted to 2,337 ft (712.3 m). One drive on the 4th level was extended 926 ft southeast.

▪ Concrete slab and old foundation walls;

▪ Two core racks near the intersection of Highway 117 and the road to the site.

Elmac Shaft Area ▪ A 100 ft (30 m) shaft provides access to one level of drifting. Drifting and cross-cutting amounted to approximately 460 ft (140 m)

▪ Concrete slab and old foundation walls.




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Historical mine Underground infrastructure Surface infrastructure

Island No. 3 ▪ No underground infrastructure, 23 m shaft collar.

▪ Island No. 3 is accessible by boat. The shaft was collared in 1999 at a depth of 23 m, the 10-ft hoist and surface buildings were installed in 2000-2001 to access the Wesdome Deposit and the permitting process was initiated. The acquisition of the Kiena mine in 2003 provided an alternate access and work was suspended.

Dorval-Siscoe Shaft Area

▪ A historical three-compartment shaft of 343 ft (105 m) provides access to level 300 ft (91 m), where there are 850 m of drifts and cross-cuts.

▪ The Dorval-Siscoe Shaft is located on Island No. 6, which is east of Island No. 3. The Dorval-Siscoe Shaft was built to access part of the Wesdome Deposit. All that remains of the surface infrastructure are old foundation walls, concrete slabs and pieces of broken equipment.

Siscoe Mine Area ▪ A shaft of 2,475 ft (754 m) provides access to 19 levels below surface from 150 ft (46 m) to 2,475 ft (754 m). The development work amounts to more than 15,000 m of drifts and cross-cuts.

▪ Mine site rehabilitation is complete, flooded sink holes, caused by the collapse of near-surface underground excavations, were observed. These areas are protected

by fences.

Siscoe Extension Area ▪ A three-compartment shaft of 725 ft (221 m) provides access to two levels below surface at 350 ft (107 m) and 725 ft (221 m). The development work amounts to about 2,000 m of drifts and cross-cuts

▪ The area is restricted and blocked by a fence and a security camera system.

5.5.1 Crown Pillar – Kiena Mine

A crown pillar of approximately 100 m thick was left in place in the Kiena Mine. The crown pillar is

located under Parker Island (S50 Zone) and the lake.

In 2010, Itasca (Andrieux, 2010) performed a preliminary empirical analysis of crown pillar

stability in the North Zone. The analysis evaluated two scenarios: the first assumes that the pillar

is made of poor quality rock (schist), and the second assumes good quality (hard) rock. In the first

scenario, the pillar showed significant potential for instability. The second case showed a situation

of near-instability.




JUNE 2020 5-7

Three recommendations were made to monitor pillar stability. Any of these three would be an

adequate protocol.

Instrumentation:

▪ Manual probe to access any rock disentanglement (weekly readings); or

▪ Multi-point borehole extensometers installed toward the surface from level 12 (weekly

readings or datalogger).

Hydrostatic barricades;

Backfilling of open stopes:

▪ Backfill the majority of the void: in the advent of a pillar collapse, the loose material from

the pillar would not have a chance to expand and fill the rest of the opening.

Level 12 was equipped with two extensometers at the North Zone Area. During mining

operations, a monthly inspection was made of level 12 and measurements taken, and level 17 at

the North S50 Area was visually inspected every 6 months: once after the spring thaw and the

other time in fall.

Since the suspension of mining operation in June 2013, the inspections have been maintained on

a quarterly basis rather than on a monthly basis.

Hydraulic barricades are currently being constructed on the six levels allowing access to the

North Zone. These barricades are designed to withstand a water or backfill column. In the event

the mill is restarted, tailings could be stored underground behind these barricades.




JUNE 2020 6-1

HISTORY

The following information was mostly borrowed from a previous NI 43-101 technical report on the

property (Turcotte et al., 2015).This chapter is divided into two sections as follows: Wesdome

historical involvement in the Project and historical work on the different mineralized zones of the

Property.

6.1 Wesdome Principal History in the Project

1945: The origin of the Company’s business can be traced back to Western Quebec Mines Inc.,

incorporated in 1945. Western Quebec Mines began developing the Dorval-Siscoe property and

carried out various exploration works on the property until 1975.

1976: Wesdome Resources Limited (“Wesdome Resources”) was created as a joint venture in

1976 for the purpose of exploring and developing the Wesdome property (formerly the Dorval-

Siscoe property). The word “Wesdome” is a combination of the names Western Quebec Mines

and Dome Exploration Ltd. Wesdome Resources was held 30% by Western Quebec Mines and

70% by Dome Exploration.

1984: On November 13, 1984, Western Quebec Mines agreed to purchase a 40% interest in the

Joubi property from Valmag Inc.

1988-1989: The School Mine property, the Shawkey South property and a 35% interest in the

Shawkey property were acquired in 1988 and 1989 by Western Quebec Mines from Valmag Inc.

1990: Production started at the Joubi mine in 1990.

1992: On October 27, 1992, Western Quebec Mines acquired the Yankee Clipper property from

Goldhunter Explorations Inc.

1993: Western Quebec Mines completed its acquisition of the 100% interest in the Joubi property.

1994: Western Quebec Mines drilled 6 DDH for 2,958 m in the southwestern part of the School

Mine Property. Erratic values reached 77 g/t Au.

1996: Western Quebec Mines acquired the Dubuisson West property from Republic Goldfields

Inc. This property was merged with the Joubi property.

1997: On November 21, 1997, Western Quebec Mines acquired the 525,000 common shares of

Wesdome Resources that were held by Dome Exploration. The result was that Wesdome

Resources became wholly-owned by Western Quebec Mines. In November 1997, Western

Quebec Mines also acquired the 65% interest of the Shawkey property from Placer Dome. This

property hosts the past-producing Shawkey mine. The Shawkey and Shawkey South properties

were merged. During the period between 1936 and 1964, the Shawkey mine produced a total of

25,637 ounces of gold from 127,737 metric tons (t) of ore grading an average 6.24 g/t Au. On

December 1, 1997, Western Quebec Mines acquired the Callahan property from Placer Dome.




JUNE 2020 6-2

1998: Western Quebec Mines staked 3 claims (the Lamothe-Extension property) adjacent to the

Lamothe property. On January 15, 1998, the Lamothe property was acquired by Western Quebec

Mines from Robert Lamothe and Alphonse Beaudoin. On November 3, 1998, Western Quebec

transferred to Wesdome Resources all its interests in the Lamothe, Lamothe-Extension (now

Vassan), Yankee Clipper and Callahan properties.

1999: In October 1999, Dynacor Mines Inc. and Western Quebec Mines signed an agreement

whereby ownership of the contiguous Siscoe and Siscoe-Extension (Dynacor Mines) and

Wesdome, Lamothe, Lamothe-Extension, Yankee Clipper and Callahan (Wesdome Resources)

properties were to be pooled into a new company in order to develop them jointly. The new

company, Wesdome Gold Mines Inc., was created by Dynacor Mines, and the latter transferred

its 100% interest in the Siscoe property and its 75% interest in the Siscoe-Extension property.

Following this, Wesdome Gold Mines Inc. then acquired 100% of the share of Wesdome

Resources from Western Quebec. During the period between 1929 and 1949, the Siscoe mine

produced a total of 802,303 ounces of gold and 306,070 ounces of silver from 2,975,785 t of ore

grading an average 9.22 g/t Au and 3.20 g/t Ag.

The Joubi mine was closed in 1999 after a 10-year production history. The historical production

amounted to 62,283 ounces from 327,561 t of mined ore.

2003: In December 2003, Western Quebec Mines purchased the Kiena Complex and

subsequently placed the property into Wesdome Gold Mines Inc., thereby completing and

consolidating Wesdome’s land package around Lac De Montigny. As a part of this transaction,

Wesdome Gold Mines acquired a 100% interest in the Kiena, Kiena West, Lac Dubuisson,

Rosenbaum, Dubuisson, Audet Block, Elmac, South Block Kiena, Option Roy and Lac De

Montigny properties, and a 50% interest in the Maufort property.

Before this transaction, the Kiena mine produced a total of 1.56 million ounces of gold from

10.7 Mt of ore grading an average 4.54 g/t Au.

2006: On February 1, 2006, River Gold Mines Ltd. and Wesdome Gold Mines Inc. completed a

merger to form the current company called Wesdome Gold Mines Ltd. (Wesdome).

On April 4, 2006, Wesdome staked seven claims and added them to the Vassan property. The

Kiena mine was in the pre-production development stage until August 1, 2006, when commercial

production commenced.

2007: On July 10, 2007, a merger was completed with parent company Western Quebec Mines

on the basis of 1.45 shares of Wesdome for each share of Western Quebec Mines. Wesdome

was the surviving operating entity.




JUNE 2020 6-3

2013: Wesdome continuously operated the Kiena mine until its temporary shutdown on June 30,

2013. The mine was placed under a care and maintenance program. During the period between

August 2006 and June 2013, the Kiena mine produced a total of 198,708 ounces of gold from

1.826,500 t of ore averaging 3.38 g/t Au.

2014-2015: A total of 47 DDH was drilled on different areas to test extension and continuity of the

zones. Number of DDH and total length (m) are as follow: Dubuisson North Zone (14 DDH;

5,634 m), S50 Zone (25 DDH; 2,809 m), Siscoe C Vein Area (1 DDH; 300 m), Presqu’ile Zone

(7 DDH, 1,968 m).

2015: Wesdome released a technical report (Turcotte et al., 2015) to provide a technical

summary of Wesdome’s primary mining and exploration assets around the Kiena mine property,

known at the time as the Quebec Wesdome Project (now collectively the Kiena Mine Complex).

Measured and Indicated Resources of 2,500,600 t at 5.59 g/t Au below the 100-m crown pillar for

a total of 449,300 oz of gold, and 134,000 t at 5.48 g/t Au within 100-m crown pillar for a total of

23,600 oz of gold was estimated. Inferred Resources were estimated to be 1,563,300 t at 7.97 g/t

Au below the 100-m crown pillar for a total of 400,400 oz of gold and 747,600 t at 8.22 g/t Au

within the 100-m crown pillar for a total of 197,600 oz of gold.

2016-2018: Wesdome carried out an underground drilling program from June 2016 to December

2018 totalling 94,722 metres. In the fall of 2018, InnovExplo was retained to perform a mineral

resource estimate using 269 new DDH for a total of 58,646 m. Targets were drilled as follows:

VC1 (20 DDH), VC6 (18 DDH), Kiena Deep A Zone (39 DDH), South Zone (9 DDH), S50

(45 DDH), other targets (138 DDH). In August 2017, an exploration ramp starting at level 100 was

developed to provide additional underground drilling platforms. Total development as of

November 2018 was 2,200 m and allowed diamond drilling to be conducted from a better

direction to intersect the steeply plunging zones and provide drilling platforms for definition drilling

on the central area of Kiena Deep A Zone. In February 2018, a heliborne high-resolution

magnetic survey over the entire property was conducted.

2019: In January 2019, Wesdome released a mineral resource estimate (MRE) technical report

(Beausoleil et al., 2019) following its 2016-2018 drilling campaign. Measured and Indicated

Resources of 2,957,400 t at 5.83 g/t Au below the 100-m crown pillar for a total of 554,700 oz of

gold, and 162,800 t at 5.32 g/t Au within 100-m crown pillar for a total of 27,900 oz of gold was

estimated. Inferred Resources were estimated to be 3,025,300 t at 7.79 g/t Au below the 100-m

crown pillar for a total of 757,600 oz of gold and 1,113,200 t at 6.97 g/t Au within the 100-m crown

pillar for a total of 249,600 oz of gold. In 2019, Wesdome carried out an underground drilling

program totalling 59,468 metres.




JUNE 2020 6-4

6.2 Detailed Historical Work by Area

6.2.1 Kiena Mine Area

The following description of historical work in the Kiena Mine Area is mostly modified and

summarized from Mailhiot (1920), Cooke et al. (1931), Hawley (1931), Bell (1935; 1936;1937),

Auger (1947), Dresser and Denis (1949), Claveau et al. (1951), Salt (1960), Robinson (1961a;

1961b; 1962), Cormier (1986a; 1986b), Sauvé et al. (1993), Morasse (1998), Beauregard and

Gaudreault (2005) and the annual reports of Wesdome Gold Mines Inc. and Wesdome Gold

Mines Ltd. (2003–2013).

1911–1914: The first record of exploration along the shores of what was known then as Lac

Kienawisik (later Lac De Montigny) dates back to 1911–1914 when prospector Barney Parker

reported the discovery of native gold-bearing quartz veins in a shear zone at the northwestern

end of the island on which Kiena’s mill was later built.

1922–1927: Martin Gold Mines and Parker Island Gold Mines did some follow-up work on the

discovery outcrop between 1922 and 1927, and tested five quartz veins with trenches and

drillholes. Soon afterward, gold was also discovered in the "Wisik vein" on the eastern shore of

Moccasin Island, the larger of two nearby islands to the east of Parker Island.

1936–1940: In 1936 Kiena Gold Mines Ltd. was created. Under the control of Ventures Ltd.,

Kiena Gold Mines immediately initiated a major surface and exploration program on the property,

which resulted in the sinking of the Parker shaft at the western end of the island. Shaft-sinking to

a depth of 455 ft (-138 m) was followed by the development of exploration drifts at the 130-, 230-,

330-, and 430-ft levels in an effort to test the surface showing, but the discovery veins were found

to be of limited extent. During an extensive stratigraphic winter drilling program from lake surface

ice in 1937–1938, the S-21 Zone (later called the North Zone) was discovered 2,300 ft (701 m)

north of the island. A crosscut was excavated toward the S-21 Zone, intersecting four mineralized

veins. Mining operations ceased in 1940 due to limited ore reserves and wartime difficulties, and

exploration activities were suspended for 20 years.

1948: Ventures drilled two diamond drillholes (DDH) totalling 1,313 ft (400.2 m).

1961–1965: Ventures, which controlled the 1958 joint venture between Kiena Gold Mines and

Wisik Gold Mines Ltd., carried out a magnetometer and geological mapping survey on their

combined claim block. A 13-hole diamond drilling program was proposed on the basis of

favourable recommendations from these surveys, but R.W. Robinson cut this proposal down to

only three holes. In 1961, the third and last drillhole (S50) of this exploration program, targeting

the "nose of a fold" and a "magnetic low that could represent a siliceous intrusive", intersected

0.22 oz/t Au (7.54 g/t Au) over 50 ft (15.2 m) of core, at approximately 800 ft (243.8 m) below the

lake’s surface.




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Following the discovery of the S50 Zone, Falconbridge acquired Ventures in 1962, took over the

management of Kiena Gold Mines and its Kiena-Wisik property, and financed the subsequent

underground exploration. Based on 79,000 ft (24,079.2 m) of drilling from the surface, original

reserves of the S50 Zone were estimated at 5 million short tons (st) (4.53 Mt) averaging

0.185 oz/t Au (6.34 g/t Au; Cormier, 1986a).

In 1963, the No. 1 shaft was collared approximately 800 ft (243.8 m) east of the Parker shaft and

sunk to a depth of 1,324 ft (403.55 m). Extensive underground exploration and a definition

diamond drill program outlined reserves of 1.5 million short tons (1.36 Mt) grading 0.265 oz/t Au

(9.09 g/t Au; Cormier, 1986a) above the 27th level (270 m below surface).

This reserve estimate is considered to be historical in nature and should not be relied

upon; however, it does give an indication of mineralization on the property. It is included

in this section for illustrative purposes only and should not be used out of context.

But due to adverse ground conditions found in a test stope conditions much related to the

structural geology of the deposit, it was concluded that no large openings could be left unfilled

and low-cost mining could not be expected. In 1965, a feasibility study on the S50 orebody

showed that mining operations would be marginal at best, with gold selling for US$35.00/oz at

that time. The property was placed under a care and maintenance program, leaving the S50

orebody dormant for the next 14 years.

1979–1984: In 1979, Falconbridge re-evaluated the property and recommended to bring it into

production. The transition from an advanced exploration project to a mine operation was

successful, and mining officially started in October of 1981. Trackless and cut-and-fill methods

were employed. Custom milling at Teck’s nearby Lamaque Mill was used for the first three years

of production until September 1984 when Kiena’s own new CIP (carbon-in-pulp) mill was

operational.

1986–1994: Campbell Red Lake Mines Ltd. became Kiena’s major shareholder on January 25,

1986, when Falconbridge sold 56.7% of its interest in Kiena Gold Mines. Following the

amalgamation in 1987 of Placer Development Ltd., Dome Mines Ltd. and Campbell Red Lake

Mines Ltd., Placer Dome Inc. of Vancouver, became the owner and operator of the Kiena mine.

On January 1, 1994, Placer Dome Inc. changed its name to Placer Dome Canada Ltd.

1997–2003: Placer Dome Canada sold the mine (along with the neighbouring Sigma mine) to

McWatters Mining Inc., who officially became the new owner and operator of Kiena on September

12, 1997. McWatters continuously operated the Kiena mine until its closure in September 2002.




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During the period between October 1981 and September 2002, the Kiena mine produced a total

of 1.56 Moz of gold from 10.7 Mt of ore grading an average 4.54 g/t Au (Table 6-1). An

exploration program commenced in 2002 to further investigate five previously identified gold

targets. McWatters carried out a 5,012-m underground drill exploration program, the first phase of

the exploration program at the Kiena Complex to test the hanging wall mineralization of the S50

orebody. A follow-up drill program began on January 20, 2003 and was completed in March 2003.

The objective of this second phase of the exploration program, corresponding to approximately

8,200 m of surface drilling program, was to extend the hanging wall mineralization, and add to or

otherwise improve the resources of the Kiena Complex.

The Kiena mine complex was acquired by Wesdome Gold Mines Inc. in December 2003.

When mining operations were suspended in 2003, measured and indicated resources stood at

3,010,000 t grading 4.25 g/t Au or 410,000 ounces of contained gold (Wesdome, 2005).

This resource estimate is considered to be historical in nature and should not be relied



Table 6-1: Kiena mine production from 1981 to 2002

(Turcotte et al., 2015)

Year Metric tonnes (t)

milled Recovered grade

Au g/t Gold production

(oz)

1981 1982 1983 1984 1985 1986 1987 1988 1989 1990 1991 1992 1993 1994 1995 1996 1997 1998 1999 2000 2001 2002

101,231 287,916 307,661 378,014 381,376 453,793 478,752 477,947 470,705 473,602 486,217 501,827 496,401 504,873 534,330 608,701 631,606 594,000 647,933 719,363 745,391 415,400

4.61 6.81 6.19 5.48 5.71 4.98 4.36 3.79 4.38 4.33 4.55 5.03 4.95 5.12 5.03 4.32 4.59 4.39 4.16 3.74 3.41 2.82

15,018 63,038 61,193 66,658 70,035 72,694 67,113 58,219 66,235 65,953 71,112 81,195 79,034 83,044 86,375 84,609 93,169 83,807 86,602 86,610 81,631 37,626

TOTAL 10,697,039 4.54 1,560,970




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2004: Wesdome Gold Mines developed and explored its properties from underground via the

Kiena shaft and underground workings. Wesdome Gold Mines drove a drift 4.5 km to the north to

explore the Wesdome property on the 520 m level, and a second drift 2.0 km to the east on the

330 m level to explore the 22 Zone on the Shawkey property. Definition drilling and development

work proceeded on known zones of gold mineralization on the Kiena property with the intent of

preparing these for commercial production as soon as possible. Late in 2004, access was

established to the VC Zone, located 500 m north of the shaft, and fan drilling of the zone

commenced from the 520 m level.

2005: Wesdome Gold Mines continued the underground development at the Kiena Mine

Complex. The North Drift (at 520 m) was used to conduct a major exploration program on the VC

Zone approximately 500 m north of the shaft. Further work was completed on the North Zone and

the 388 Zone from the 330-m level. Progress was also made on the East Drift (at 330 m).

Development crews also drifted into the Martin Zone (from which Shawkey produced ore in the

1930s).

During the same year, Beauregard and Gaudreault (2005) estimated that the four zones (North,

VC, 388 and Martin) contained measured resources of 574,023 t at 4.45 g/t Au, and indicated

resources of 750,137 t at 4.52 g/t Au. Wesdome Gold Mines also completed a new resource

estimate for the VC1 to VC3 zones, calculating a total of 737,900 t at 5.05 g/t Au.




2006–2013: The Kiena mine was in pre-production development stage until August 1, 2006, when

commercial production commenced. Wesdome continuously operated the Kiena mine until its

shutdown in June 30, 2013. The mine was placed under a care and maintenance program.

During the period between August 2006 and June 2013, the Kiena mine produced a total of

198,708 ounces of gold from 1.826,500 t of ore at an average grade of 3.38 g/t Au (Table 6-2).

Table 6-2: Kiena mine production from 2006 to 2013





(oz)

2006 2007 2008 2009 2010 2011 2012 2013

94,200 284,757 241,641 302,034 285,527 255,311 265,872 97,158

3.07 3.87 5.19 3.65 3.50 2.38 2.20 2.49

9,300 35,404 40,344 35,398 32,162 19,516 18,814 7,770

TOTAL 1,826,500 3.38 198,708




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6.2.2 Wisik Shaft Area

The following description of historical work in the Wisik Shaft Area is mostly modified and

summarized from Bell (1935; 1936; 1937), Taschereau (1936), Denis (1937), Auger (1947),

Dresser and Denis (1949), and the diamond drill logs of Wisik Gold Mines Ltd.

1934–1937: The property around the Wisik deposit was owned by H. Klee who founded Wisik

Gold Mines. The property was held under option in 1935 by Teck Hughes interests, who carried

out some diamond drilling. The only visible mineralized showing of substantial size was a quartz

vein, exposed for only a few feet on the eastern shore of the largest Moccasin Island. In addition

to the diamond drilling in the vicinity of the vein, some holes were also put down along the water

stretch between the island and Shawkey Peninsula. The option was subsequently allowed to

lapse.

In 1936, the overburden was removed by Wisik Gold Mines for the purpose of sinking a shaft on

Moccasin Island. A three-compartment shaft was sunk below 300 ft (91.4 m). Drifts were

excavated along the principal vein or shear on the 200- and 300-ft levels, totalling 2,100 ft

(640.1 m). In 1937, metallurgical testwork was performed at the Ottawa facilities of the

Department of Mines and Resources of Canada on two samples of gold ore: one from the Wisik

North drift (237 lbs, 107.5 kg), and the other from Wisik South (200 lbs or 90.7 kg). The North rift

sample assayed 0.125 oz/t Au (4.29 g/t Au) and 0.19 oz/t Ag (6.51 g/t Ag), whereas the South

Drift sample returned 0.675 oz/t Au (23.14 g/t Au) and 0.09 oz/t Ag (3.09 g/t Ag). In September

1937, work was suspended and the mine flooded.

1958–1959: In July 1958, Ventures Ltd., who controlled the 1958 joint venture between Kiena

Gold Mines and Wisik Gold Mines, drilled one hole (W-1) totalling 350 ft (106.7 m) on Moccasin

Island, northwest of the Wisik shaft. In January 1959, another hole (W-2), totalling 750 ft

(228.6 m), was drilled to the north of hole W-1 to test the eastern extension of the No. 1 Zone

located on the Kiena mine property.

Later, Kiena Gold Mines took over management of the Wisik property and eventually formed the

Kiena-Wisik property.

6.2.3 Shawkey Mine Area

The following description of historical work in the Shawkey Mine Area is mostly modified and

summarized from Mailhiot (1920), Cooke et al. (1931), Hawley (1931), Bell (1935; 1936; 1937),

Dresser and Denis (1949), Ingham (1950), Claveau et al. (1951), Ingham (1953), Bourret et al.

(1956), Sauvé (1985), Chevalier (1989b), Sauvé et al. (1993), Beauregard and Gaudreault

(2005), and the annual information forms of Western Quebec Mines Inc., Wesdome Gold Mines

Inc. and Wesdome Gold Mines Ltd. (1997–2013).




JUNE 2020 6-9

1911: Fred La Palme discovered a gold-bearing vein (No. 1 Vein) on the south shore of Lac De

Montigny.

1917–1919: The Martin Gold Mining Company Ltd. was incorporated to develop the new gold

showing, and in 1917, underground work commenced as well as the erection of a small mill which

was completed the following year. A two-compartment vertical shaft was sunk 125 ft (38.1 m) on

the shore of Lac De Montigny, and a small amount of drifting and cross-cutting was done. About

600 pounds of rock were milled during the summer of 1918. Work was suspended in 1919.

1921–1923: Two years later, the property was sold to John Dalton. In 1922, an option was

secured by J.J. Godfrey and underground work resumed. The shaft was deepened to 325 ft

(99.1 m), and about 700 ft (213.4 m) of drifts and cross-cuts were excavated. A total of 4,500 ft

(1,371.6 m) of diamond drilling was also carried out. The option lapsed in 1923.

1934–1938: The mine was purchased by Shawkey Gold Mines Ltd. who became operator. During

1935, drifting and raising on the 125’, 225’ and 325’ levels revealed continuous lengths of ore. A

raise was completed to the surface, north of the shaft. Shrinkage stopes were prepared above the

first and second levels.

The shaft was deepened to 725 ft (221.0 m) and new levels were established at the 450’, 575’

and 625’ horizons. A new mill was erected and started production in 1936. Several gold-bearing

veins were discovered over the course of the exploration, but all the production came from the

discovery vein. The vein was developed for a maximum length of 1,000 ft (304.8 m). By 1938, this

had been mined out, and operations were suspended.

A total of 3,915 ft (1,193.3 m) of surface drilling, 45,885 ft (13,985.7 m) of underground drilling,

and 1,125 ft (342.9 m) of drifting and crosscutting had been carried out. Total production was

25,414 ounces of gold from 137,978 short tons (st) (125,174 t) of processed ore, for a recovery of

0.184 oz/t Au (6.31 g/t Au).

1945–1951: In 1945, Shawkey Mines Ltd. took over the abandoned workings and began an

extensive diamond drill program. From 1945 to 1947, the company completed about 37,000 ft

(11,277.6 m) of surface drilling, 28,500 ft (8,686.8 m) of which was used to explore the No. 10

Vein. Another 9,000 ft (2,743.2 m) of drilling was used in cross-sectional exploration on the east

side of the Thompson River and south of the No. 10 Zone. Underground drilling from old workings

amounted to 20,000 ft, used in lateral and depth tests from the 4th and 6th levels. Total

underground lateral development in the form of drives, crosscuts and drifts amounted to 2,337 ft

(712.3 m). Over 7,000 ft3 (198.2 m3) of rock was slashed. One drive on the 4th level was extended

926 ft southeast, under the Thompson River, to reach a body of diorite containing auriferous

quartz veins. The No. 10 Zone was explored for 2,300 ft (701.0 m) by drilling. Underground

drilling and drifting led to the discovery of new auriferous zones (No. 9, No. 11 and No. 12 zones).




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The No. 2 shaft was collared in July 1950, adjacent to the north side of the No. 10 Zone,

approximately 900 m south of the No. 1 shaft. Shaft-sinking to a depth of 743 ft (226.5 m) was

completed in April 1951. Four level stations were established at 250 ft (76.2 m), 400 ft (121.9 m),

550 ft (167.6 m) and 700 ft (213.4 m). After carrying out 927 ft (282.5 m) of drifting and

crosscutting on the 700’ level and 605 ft (184.4 m) on the 550’ level, as well as 2,265 ft (690.4 m)

of underground drilling, the mine was closed again in September 1951. A total of 1,735 st

(1,574 t) of ore grading 0.09 oz/t (3.09 g/t Au) was mined from various drifts and slashes.

1962–1964: The project came under the control of Con-Shawkey Gold Mines Ltd. The No. 2 shaft

(No. 10 Zone) was reopened, and more exploration was completed. A bulk sample of 51.3 st

(46.5 t) was sent to the Ministry of Natural Resources pilot plant, and an average grade of 0.23

oz/t (7.89 g/t Au) was obtained from 53 samples. Another bulk sample of 1,039 st (942.6 t) was

sent to the Malartic Goldfields Mill where an average grade of 0.053 oz/t Au (1.82 g/t Au) was

obtained. In addition, 14,000 ft (4,267.2 m) of surface drilling, 1,600 ft (487.7 m) of underground

drilling, 1,400 ft (426.7 m) of underground development, and Mag and EM surveys were also

carried out on the project.

1964–1966: Noranda carried out 3,710 ft (1,130.8 m) of diamond drilling that concentrated on the

No. 10 Zone. Mag and EM surveys were also performed on the project.

1972–1976: Umex acquired an interest in the project and drilled six drillholes for a total of 2,240 ft

(682.8 m). Mag and EM surveys were conducted on the project.

1979–1989: Les Mines Sigma (Québec) Ltée started work to acquire a 65% interest in the

Shawkey property from Valmag Inc. During three years, 45 km of Mag surveys and more than

11,500 m of diamond drilling were completed on the 22 Zone. From June to March 1984, another

24 km of Mag surveying was completed over Lac De Montigny. Twenty-three holes were drilled

for a total of more than 6,000 m. From June to August 1988, nine more holes were drilled for an

additional 2,224 m. The Shawkey South property and the 35% interest in the Shawkey property

were acquired by Western Quebec Mines Inc. in 1988 and 1989 from Valmag Inc.

In 1988, Placer Dome Inc. acquired Les Mines Sigma (Québec) and carried out a drilling program

comprising seven drillholes for a total of 1,897.5 m. Drilling concentrated on two target areas: the

West Zone and the 22 Zone. The Shawkey Property consists of four mining concessions in

Dubuisson Township.

1990–1997: The Shawkey property was under a joint venture with Placer Dome Inc. until

November 1997. In 1990, Placer Dome Inc. estimated the mineral inventory of the Shawkey

property to be 883,132 t in the “possible” category, with an average grade of 4.04 g/t Au (Lebel

and Lafleur, 1991). The mineral inventory was contained in six lenses designated by the letters A

to F. The mineral inventory was performed using the polygonal method and a specific gravity of

2.7 g/cm3.




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1997: Western Quebec Mines purchased, from Placer Dome, 70% of the outstanding shares of

Wesdome Resources Ltd. (which it would later own 100%), and a 65% interest in the Shawkey

property (which it would later own 100%).

2002–2003: In 2002, Western Quebec Mines drilled 11 holes totalling 1,248 m to test the No. 22

Zone. In 2003, the company added ten more holes for another 1,657 m. During the period

between 1936 and 1964, the Shawkey mine produced a total of 25,637 ounces of gold from

127,737 t of ore with an average grade of 6.24 g/t Au (Table 6-3).

Table 6-3: Shawkey mine production from 1936 to 1964





(oz)

1936-1938 1945-1951 1962-1964

125,174 1,574 989

6.31 3.08 2.11

25,414 156 67

TOTAL 127,737 6.24 25,637

6.2.4 Elmac Shaft Area

The following description of historical work in the Elmac Shaft Area is mostly modified and

summarized from Mailhiot (1920), Cooke et al. (1931), Hawley (1931), Bell (1935; 1936),

Koulomzine and Brossard (1946), Dresser and Denis (1949), Chevalier (1989a), Beauregard and

Gaudreault (2005), and the annual information forms of Western Quebec Mines Inc., Wesdome

Gold Mines Inc. and Wesdome Gold Mines Ltd. (1997–2013).

1919–1935: The property was originally known as Fosie-Kengrow, these being the names of the

men who staked the claims in 1919. From 1919 to 1935, the property was successively controlled

by Union Mining Corporation, Unison Gold Mines Ltd., Lorette Mines Ltd., Minorand Co-operative

Company Ltd., Minrand Gold Ltd., and finally by Crossroads Gold Mines Ltd. The deposit was

discovered in 1919. The discovery vein, which was exposed over a distance of roughly 200 ft

(61.0 m), displayed a contorted pattern with numerous offshoots and stringers. Some diamond

drilling was done in 1922, establishing sufficient ore for mining purposes, as long as ore grades

were high enough. Mining machinery was installed, and in 1925, a shaft was sunk to a depth of

100 ft (30.5 m) where some underground work was carried out. The mine was again operated in

1932–33, and there has been sporadic development work, including a small amount of

underground work and diamond drilling, as well as the assembling of a small mill. In 1935, the

mine workings were flooded.




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1940–1946: Elmac Malartic Mines Ltd. carried out work around the shaft and in underground

workings. A new mineralized zone was discovered (the “Carbonated Zone”). All known zones

were resampled. A drilling program was also conducted in the northern part of the property

(Claims Potter-Kee). A magnetic survey was carried out over the entire property in 1945.

1963–1965: Elmac Malartic Mines conducted an 18-hole drilling program for a total of 8,800 ft

(2,682.2 m). Seven holes were drilled in the northern part of the property (Potter-Kee Claims) on

Lac De Montigny. Nine holes were drilled in the centre of the property, in the shaft and

Carbonated Zone areas. The drilling program cut the Carbonated Zone, which assayed 23.78 g/t

Au over 2.74 m. Two holes were drilled near the Piché River in the southern part of the property.

In 1965, another drilling program (1531.0 m) was conducted on Lac De Montigny, resulting to the

discovery of a new mineralized zone associated with feldspar porphyries. The new zone assayed

4.14 g/t Au over 6.24 m.

1978–1983: Les Mines Sigma (Québec) Ltée optioned the property. A total of seven holes was

drilled on the Carbonated Zone totalling 1,905 m. Between 1981 and 1983, the company carried

out 22.3 km of magnetic surveying and 3.6 km of induced-polarization surveying. In addition, a

total of 53,823 ft was also drilled on the property.

1989: Les Mines Sigma (Québec) conducted an 8-hole drilling program in the northern half of the

property during the winter of 1989 for a total of 2,117.75 m.

1997–2002: Placer Dome Canada Ltd. sold the Elmac property to McWatters Mining Inc., who

officially became the new owner and operator of the property until 2003.

2003: In December of 2003, Western Quebec Mines purchased the Kiena Complex, including the

Elmac property.

6.2.5 Joubi Mine Area

The following description of historical work in the Joubi Mine Area is mostly modified and

summarized from Mailhiot (1920), Bell (1935), Ingham (1944b), Hinse (1975), Lavery (1983),

Laforest (1987) Castonguay (1995), and the annual information forms of Western Quebec Mines

Inc. (1996–2006).

1919: Before 1919, prospecting work on the Clowse claim led to the discovery of a narrow gold-

bearing vein. The vein was traced at surface over a length of 100 ft (30.5 m). A 27-ft-deep (8.2-m)

exploration shaft was sunk on the vein.

1934: During spring and summer 1934, Amity Gold Mines Ltd. carried out surface trenching,

surface sampling and diamond drilling, chiefly in the vicinity of the old shaft sunk by Clowse

Claim. Four holes were reportedly drilled to test this old occurrence.




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1941–1942: A diamond drilling program was started in July 1941 by Seventh Malartic Mines Ltd.,

and it continued until April 1942, when 28 holes had been completed for a total of 13,550 ft. Only

four holes were located on the Joubi property, one of which constituted the first indication of the

presence of the Range Line Zone.

1943–1944: In 1943, Perron Gold Mines optioned the property, and diamond drilling was

resumed in the spring of 1944. Eighteen holes were drilled during the program.

1960: Iso-Newlund Mines covered part of the property with a Mag survey, followed by five

drillholes.

1964: One hole was drilled by Amerel Mining Company, followed by a Mag survey.

1973: The Joubi Mining Corporation acquired the property in 1973, and drilled 24 holes on the

Range Line Zone.

1979: Massey-Gauthier Ltd. drilled six holes on the Range Line Zone.

1981: A joint venture was made up of U. F. Venture Associates (70% participating interest) and

Messeguay Mines Inc. (30% participating interest). The Joubi JV continued the assessment of the

Range Line Zone with 1,341.4 m of core drilling in 11 holes during 1981, bringing the total drilling

in the zone to 6,838 m in 46 holes.

1982: The participation in the JV was modified in 1982. The result was a 51.11% participating

interest for U. F. Venture Associates, a 30% participating interest for Messeguay Mines, and a

18.89% non-participating interest for V. Audet. In 1982, based on all drilling results to date,

W.N. Ingham estimated that the Range Line Zone contained 167,545 t grading 7.34 g/t Au in

three separate shoots (Lavery, 1983).




Between 1982 and 1983, 12 holes, totalling 2,402.7 m, were added on the Range Line Zone. In

late 1982, detailed total field magnetic, vertical gradient magnetic and VLF surveys had been

completed by the Joubi JV.

1984–1986: Western Quebec Mines carried out an exploration program including different

geophysical methods, followed by some 3,178 m of diamond drilling distributed in 11 holes.

1987–1989: Western Quebec Mines agreed to grant Messeguay Mines and Oasis Resources Inc.

working rights and an option to acquire a 49% interest in the Joubi property. Western Quebec

Mines had Minexpert Inc. complete an evaluation review, which recommended an underground

exploration program to define the essential parameters required for a better understanding of the

nature of the gold mineralization. Surface installations were constructed, a three-compartment




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shaft was sunk to a depth of 253 m with four shaft stations, 2,780 m of crosscuts and drifts were

developed, 1,094 m of raises were completed, and a diamond drilling program totalling 47,812 m

was carried out. During this period, Western Quebec Mines developed the zones defined by the

previous underground exploration program, conducted detailed drilling of the zones (3,645 m) and

began mining operations.

1990: In 1990, Western Quebec Mines completed development work and mining of the zones

defined in 1988–89, conducted a drilling program of 15,084 m (zone definition, surface

exploration, and exploration below the last level). In 1991, exploration work discovered the

F Zone at the border of the Dubuisson West property. Western Quebec Mines drilled 12,614 m of

core on the Joubi property. Since the mineable reserves on the Joubi property were almost

exhausted, Western Quebec Mines decided to deepen the shaft by 71 m to reach the lower

A Zone.

1992: In early 1992, Western Quebec Mines was granted an option from Republic Goldfields Inc.

on the Dubuisson East property. Western Quebec Mines also purchased the Dubuisson West

property from Republic Goldfields Inc. Both properties were adjacent to the Joubi property.

In spite of the shaft deepening, reserves were insufficient to maintain the operations. One

kilometre (1 km) of drifts was developed at the 2nd level to access the Dubuisson East zone.

Western Quebec Mines drilled 268 m of core on the Dubuisson West property, 3,610 m on the

Dubuisson East property (exploration and definition), and 2,409 m on the Joubi property. By the

end of the year, the reserves from the known economic zones on the Joubi property had been

exhausted.

1993: The Dubuisson East zones were developed and mined from the 2nd level to the surface.

Since reserves were rapidly running out, it was decided that the drift on the 5th level would be

extended under the Dubuisson East workings to access additional reserves. Definition drilling

(3,633 m of core) was completed on the Dubuisson East property. Mining of the Dubuisson East

zones was slowed by wall instabilities, which resulted in an excessive dilution problem.

1994: Western Quebec Mines continued to mine the Dubuisson East upper zones and to develop

the lower zone. The company also extracted a few pockets of ore on the Joubi property. Also

carried out were definition drilling and exploration on the Dubuisson East property (3,171 m), and

drilling of the A Zone below the 5th level (659 m) on the Joubi property.

1995: Western Quebec Mines commenced a two-year underground exploration and development

program on the Joubi property. The first phase of this program involved 200 m of shaft sinking to

a depth of 524 m, upgrading of production infrastructure, development of a loading system, and

establishment of levels 360 and 440. Due to the shaft sinking program, production was

interrupted for more than six months. Almost all production came from the Dubuisson East

property where the bulk of the reserves were located.




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1996: The emphasis continued to be on exploration following the shaft-deepening program and

the establishment of new drifts at a depth of 440 m. Western Quebec Mines completed a total of

15,000 m in 110 drillholes along the 1.1-km strike length of the Joubi shear zone. This drilling

identified a small, high-grade zone (DE-F) above the 440 m level. In addition, preliminary drilling

on the 440 m level identified the down-dip extension of zones DE-B/C over a strike length of

80 m, which remained open to the east and at depth.

1997–1998: In 1997, 90 holes were drilled for a total of 8,600 m, resulting in the discovery of the

DE-F zone below the 440 m level. In 1998, the company conducted an extensive underground

exploration effort involving 480 m of drifting and 14,200 m of drilling in 81 holes. The goal of the

program was to test the depth and strike potential of the Joubi shear zone in an effort to identify

significant new reserve blocks. Results were insufficiently encouraging to justify further

development.

1999: The Joubi mine was closed in 1999 after a 10-year production history. Total production

amounted to 62,283 ounces from 327,561 t of mined ore. Ore was custom-milled at facilities in

the Val-d'Or area when excess capacity was available, with gold recoveries of 98%. The

headframe was dismantled and the shaft sealed. The three-compartment shaft reached a depth

of 524 m with levels developed at 70 m, 120 m, 220 m, 280 m, 360 m and 440 m. Ore chutes

were installed at levels 120 m, 220 m, 280 m and 440 m.

Table 6-4: Joubi mine production from 1990 to 1999





(oz)

1990 1991 1992 1993 1994 1995 1996 1997 1998 1999

30,146 36,998 36,572 28,419 57,115 20,166 35,270 39,191 27,470 16,214

10.94 6.99 7.64 4.49 4.55 5.37 4.73 4.72 4.17 6.62

10,601 8,319 8,983 4,104 8,351 3,484 5,364 5,947 3,679 3,451

TOTAL 327,561 5.91 62,283

2016: Agnico Eagle Mines purchased some claims from Wesdome to acquire a portion of the

property. Agnico Eagle Mine, granted a 3% NSR on the Joubi Property.




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6.2.6 Dorval-Siscoe/Wesdome Deposit Area

The following description of historical work in the Dorval-Siscoe Area is mostly modified and

summarized from Beckman et al. (1933), Bell (1937), Ross et al. (1938), Ross and Asbury (1939),

Koulomzine (1941; 1942; 1943), Ingham (1944a; 1947), D’Aragon (1947), Auger (1947), Dresser

and Denis (1949), Young (1955), Salamis (1970), Prud’homme (1970), Audet (1975a; 1975b),

Matheson (1976), Audet (1979), Gardiner (1987; 1988), Sauvé et al. (1993), Beauregard and

Gaudreault (1999; 2005), Turcotte and Pelletier (2009), and the annual information forms of

Western Quebec Mines Inc., Wesdome Gold Mines Inc. and Wesdome Gold Mines Ltd.

(1996- 2013).

1933–1937: Dorval-Siscoe Gold Mines drilled 37 DDH for a total of 7,050 m, and established the

presence of a wide and intensely sheared zone (Dorval-Siscoe Main Break). Snowshoe Mines

Ltd. drilled five holes totalling 844 m on the Snowshoe intrusion. A diamond drillhole cut the

Dorval Siscoe Main Break. The material in the zone consisted of schistose and talcose

granodiorite with laminated quartz-carbonate stringers up to 10 in (25 cm) wide, and a feldspar

porphyry dyke about 10 ft (3 m) in section width that was injected and replaced by vein quartz.

The mineralization consists of pyrite, chalcopyrite, galena and tourmaline.

Good results obtained by Dorval-Siscoe Mines prompted the company to sink a three-

compartment shaft on Island No. 6 in 1937 and 1938, to a depth of 343 ft (104.5 m). Other work

included the development of about 850 m of drifts and crosscuts on the 300’ level, and a drilling

program of 14 underground DDH totalling 686 m. A strong vein was exposed for a length of 780 ft

(237.7 m) in a drift on the 300’ level.

Particular emphasis was placed on extending the K Zone in the Siscoe gold mine onto the Dorval-

Siscoe property. At the time, the K Zone was considered spatially related to most of the

mineralization on the Dorval-Siscoe property. The primary focus of the work was to outline a

broad zone of weak to moderate mineralization, as described in the report by Koulomzine (1941).

1941–1943: Camp Bird Gold Mines Ltd. took an option on the project held by Dorval-Siscoe Gold

Mines. A dip needle magnetic survey was carried out over 590 ha on the project, outlining many

magnetic anomalies. Twenty-four holes totalling 5,400 m were drilled from the surface, and 16

holes totalling 1,467 m were drilled underground from the earlier Dorval-Siscoe workings.

1945–1948: Another dip needle magnetic survey was conducted by Snowshoe Gold Mines Ltd.

over an additional 191 ha. In addition, 14 surface DDH, totalling 2,671 m, were drilled on the

periphery of a large circular magnetic depression (granodiorite plug) located in the centre of the

property. In 1946 and 1947, Western Quebec Mines (incorporated in 1945) began developing the

property held by Camp Bird Mines Ltd. Twelve surface DDH, totalling 3,394 m, were drilled on

magnetic anomalies, providing encouraging results. 1955: Snowshoe Gold Mines Ltd. carried out

an EM survey over the western part of the Snowshoe intrusion.

1963–1965: Western Quebec Mines drilled four holes (holes 87 to 90), totalling 1,559 m, on the

A Zone, and in 1964, completed a Mag survey over the western and southwestern part of the

property.




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1965–1970: Kerr Addison Mines and Western Quebec Mines conducted a geophysical survey

over the project, including 209 km of Mag and 12.9 km of EM surveys, and drilled four holes

totalling 1,613 m (65-1 to 65-4). In 1970, Western Quebec Mines initiated other geophysical

surveys on the western part of the project, including 132.9 km of Mag, and 182 km of Max-Min

(EM) and induced polarization (IP). The goal of these surveys was to delineate massive sulphides

associated with peridotites. In addition, nine holes (W70-1 to W70-9) were drilled, totalling

1,373 m.

1975–1983: Wesdome Resources Ltd., a company owned by Dome Exploration Co. Ltd. and

Western Quebec Mines, drilled 136 holes totalling 37,999 m. This phase can be split into four

major periods:

▪ 1975–1976: A 78-km VLF survey was carried out on the project, and 38 holes (82-1 to 82-

38) were drilled for a total of 10,584 m. The goal of this program was to evaluate the general

potential of the project;

▪ Winter 1979: Four holes (82-39 to 82-42) were drilled, for a total of 1,273 m. The goal was to

complete a transverse section north of the A and B zones (Audet, 1979);

▪ 1980–1981: A drilling program was carried out to delineate the mineralization and evaluate

gold reserves in the A and B zones and associated “flat” quartz veins. In total, 19,740 m

(67 holes) were drilled;

▪ Winter 1983: Drilling was conducted to extend the A Zone mineralization eastwards. A first

reconnaissance field program was carried out to identify mineralization in an altered

(albitized) monzodiorite on the western periphery of the property, adjacent to a project held

by Falconbridge Nickel Ltd. A total of 26 holes (82-105 to 82-130) were drilled for a total of

6,348 m (Duhaime, 1983).

1984: Wesdome Resources established a grid of 116.5 km on the project, and a total field and

gradient Mag survey was undertaken. During the same year, 51 holes, totalling 18,656 m, were

drilled on the project, mainly in two sectors: 21 holes in the intrusive dyke complex (the

Falconbridge Zone) for a total of 10,935 m; and 30 holes in the A and E zone extensions for an

additional 7,721 m. During this program, a new zone, the F Zone, returned encouraging results.

1987: Wesdome Resources completed a geophysical survey (seismic refraction) followed by a

30-hole drilling program (holes 82-177 to 206) for a total of 12,180 m.

1988: Another geophysical program consisting of seismic refraction, IP, Mag, EM and bathymetry

was carried out by Sigma Mines Ltd. After that, Wesdome Resources completed a 13-hole winter

drilling program (holes 82-207 to 219) for a total of 5,318 m, followed by another 13-hole drilling

program in the fall of 1988 (82-220 to 82-232), for an additional 4,524 m.

During the 1980s, geologists of Placer Dome Inc. and subsidiary Les Mines Sigma (Québec) Ltée

separately estimated the total contained mineralization for the A, B, C, D, E and E3 zones as

2.7 Mt grading 4.6 g/t Au (using various cut-off grades; Beauregard and Gaudreault, 2005).




JUNE 2020 6-18




1998–1999: Wesdome Gold Mines built a 5.5-km access road on the northern peninsula, and

completed rock mechanic tests and seismic surveys for overburden and rock quality on the

former site of the decline portal. Results revealed the presence of a deep trough (over 60 m) over

an east-west striking, multi-metre wide, sheared and altered ultramafic unit. These results

indicated a ramp project would not be feasible. One hundred (100) holes (37,278 m) were drilled,

including six holes on the Yankee Clipper property to the north. Three distinct zones of the

Wesdome property were tested: the sector north of the K Zone, the E3 and E zones, and the A

and B zones. Wesdome Gold Mines was planning to sink an exploration shaft on Island No. 3. A

600-m pilot hole was completed on the island for rock mechanic purposes prior to shaft sinking,

with positive results. Preparation work, which consisted of levelling the site (Island No. 3), was

conducted during the winter of 1998–99. In 1999, permitting and engineering work continued, and

limited mining equipment was purchased.

2000–2001: During 2000, work concentrated on the construction of the surface infrastructure

required to commence shaft sinking and underground development. The shaft was collared at a

depth of 23 m, the hoist and surface buildings were installed, the wharf-barge access system was

made fully functional, and was made using a specialty submarine electrical cable. In addition, two

exploration holes were drilled to test the western limits of the A zone.

6.2.7 Siscoe Mine Area

The following description of historical work in the Siscoe Mine Area is mostly modified and

summarized from Mailhiot (1920), Cooke et al. (1931), Hawley (1931), Bell (1937), Auger (1940),

Auger (1947), James (1949), Dresser and Denis (1949), Gill (1981), Timmins and Wing, (1981),

Trudel (1985), Allard (1988) and Sauvé et al. (1993).

1912–1919: Siscoe Island was staked in 1912 by Siscoe Mining Syndicate. Gold was discovered

on the northernmost part of the property during the initial prospecting in 1911 and 1912, but gold

showings on Siscoe Island were not reported until three or four years later. The first gold

discovery was made in 1913.

Between 1913 and 1919, exploration work was conducted on four principal veins: A, B, C, and D.

A 45-ft exploration vertical shaft was sunk on the A Vein. The vein was exposed over a length of

90 ft (27.4 m). The vertical exploration shaft of the B Vein was 32 ft (9.8 m) deep. On the C Vein,

an exploration shaft was sunk to a depth of 100 ft (30.5 m), inclined at an angle of 35°, following

the dip of the vein. The zone of the D Vein comprised a number of scattered veins and quartz

lenses measuring several feet in length. Several quartz veins contained nests or pockets rich in

native gold, in which wonderful specimens were sometimes found. An exploration shaft 88 ft

(26.8 m) deep was sunk on the vein.




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1923–1927: The property was bought by British Mineral Corporation, and in 1923, Siscoe Gold

Mines Ltd. was created. In June 1926, the underground workings included a shaft sunk to a

vertical depth of 105 ft (32 m) on the D Vein, and about 1,500 ft (457.2 m) of drifting and

crosscutting. Most of this work has been of an exploratory nature, directed to finding single veins

large enough to mine. The long crosscut that runs southwest from the shaft to the granodiorite-

basalt boundary cut a vein, known as the K Vein, which seemed to follow that boundary. In 1927,

active development on the C shaft and C Vein was recommenced. Since renewing operations,

the shaft was continued to a depth of 500 ft (152.4 m), and much lateral work was completed on

the 170’, 300’, 400’ and 500’ levels.

1929–1930: Milling equipment was installed, and in January 1929, the mine came into production.

Later in 1929, following the discovery of a vein nearly parallel to the cross-cut driven south of the

D-shaft, an extensive diamond drilling program was carried out that proved the continuity of the

vein, which became known as the New Vein. A new three-compartment shaft (the Central shaft)

was completed in spring 1930 to a depth of 472 ft (143.9 m). From this, cross-cuts were driven

northwest on the 300’ and 450’ levels at distances of 50 (15.2 m) and 100 ft (30.5 m) beyond the

New Vein. From the Central shaft, the workings on the 300’ level were driven east to connect with

the 5th level of the inclined C shaft by an 85-ft raise from the latter. At the 450’ level, a cross-cut

was driven east to intersect the C Vein on this level.

1930–1946: By the end of 1936, the mine had been developed on levels to a vertical depth of

1,350 ft (411.5 m). For the first 600 ft (182.9 m), levels are at intervals of 150 ft (45.7 m); below

this they were spaced at 125-ft intervals. The Central shaft serviced the entire mine. It was

deepened to 1,900 ft (579.1 m) to permit the opening of the newly opened 11th and 12th levels. By

the end of 1939, the Central shaft had reached the 19th level at a depth of 2,475 ft (754.4). At this

time, the mine had been continuously in production since 1929.

1946–1951: Starting in 1940, and despite an intense exploration program, no new ore sources

were found. In 1949, the mine was closed and all reserves were mined out. In 1951, the mill and

all equipment were sold. Exploration activities were suspended until 1981.

1981: Canzona Mineral Inc. carried out an EM (VLF) survey covering all of Siscoe Island. In the

spring, a 19-hole diamond drilling program was completed on Siscoe Island. Due to the significant

mineralized zones intersected, Phase 2 of the drilling program was completed in June, consisting

of four DDH, drilled northward into a VLF EM conductor associated with the K Zone.

1984–1990: From 1984 to 1990, Maufort Resources Inc. completed exploration and underground

work in joint venture partnership with Teck Corporation (1984–1987) and Cambior Inc. (1987–

1989). A total of 20,693 m of diamond drilling was carried out and the mine pumped dry.

1993–1997: In 1993, Maufort Resources Inc. changed its name to Dynacor Mines Inc. Dynacor

Mines completed mapping, outcrop stripping, diamond drilling and sampling. In 1997, nine DDH

were sunk on the property for a total of 3,170 m.




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1999: As part of its Reorganization Agreement, Dynacor Mines transferred all its interests in the

Siscoe property (100%) to Wesdome Gold Mines on November 9, 1999.

During the period between 1929 and 1949, the Siscoe mine had produced a total of 882,303

ounces of gold and 306,070 ounces of silver from 2,975,785 t of ore grading an average of

9.22 g/t Au and 3.20 g/t Ag (Table 6-5).

Table 6-5: Siscoe mine yearly production from 1929 to 1949

Year Metric tonnes

(t) milled

Recovered grade Au g/t

Gold production (oz)

Recovered grade Ag g/t

Silver production

(oz)

1929 1930 1931 1932 1933 1934 1935 1936 1937 1938 1939 1940 1941 1942 1943 1944 1945 1946 1947 1948 1949

27,067 30,613 50,508 58,059 87,407

112,630 135,236 164,364 181,895 170,342 171,965 176,251 208,710 288,668 290,407 294,366 241,466 143,148 64,649 60,984 17,049

17.07 18.02 22.10 26.06 19.48 16.93 14.82 13.08 12.89 12.19 9.76 8.15 6.63 5.13 4.35 4.16 4.33 4.95

10.37 8.27

12.69

14,853 17,740 35,883 48,651 54,729 61,291 64,446 69,138 75,383 66,783 53,982 46,159 44,461 47,630 40,656 39,384 33,610 22,799 21,556 16,212 6,957

1.38 1.44 0.99 2.04 3.14 2.83 3.70 3.23 3.61 3.63 3.76 3.17 3.79 4.20 3.51 3.30 2.30 2.24 2.33 2.15 0.72

1,200 1,420 1,614 3,810 8,826

10,247 16,089 17,090 21,095 19,906 20,765 17,973 25,457 38,961 32,808 31,196 17,864 10,304 4,847 4,206 392

TOTAL 2,975,785 9.22 882,303 3.2 306,070




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6.2.8 Siscoe Extension Area

The following description of historical work in the Siscoe Extension Area is mostly modified and

summarized from Hawley (1931), Bell (1937), Denis (1937), Ross et al. (1938), Auger (1940),

Auger (1947), Dresser and Denis (1949), Sullivan Consolidated Mines Ltd. (1957) and Dussault

(1988).

1929–1937: During the winter of 1929–1930, some 10,000 ft (3,048 m) of diamond drilling was

undertaken, by Siscoe Extension Gold Mines Ltd., on the property. Drilling was undertaken in the

hope of picking up the eastern extension of the Siscoe granodiorite and possible veins. Following

the drilling program, a three-compartment shaft was sunk on the lake’s shore, and underground

work was carried out on the 350’ level. In 1936, diamond drilling was carried out from both

underground and surface (from ice on Lac De Montigny). The latter was chiefly concerned with

exploration for the presumed continuation onto the property of the northeastern branch or split of

the K Zone of the Siscoe mine. But the drilling program was unsuccessful in locating the Siscoe K

Zone beneath the lake. The overburden was penetrated to depths of up to 202 ft (61.6 m), but

bedrock was not definitively reached. In 1937, the shaft reached a depth of 750 ft and the 750’

level was established. Drifting was carried out on the 750’ level.

1938: Operations were suspended in the fall of 1938, and the mine was flooded.

1946: Siscoe Gold Mines optioned the property and dewatered the underground workings.

Geological surveying and re-sampling were carried out underground. Following this work, the

option lapsed and the underground workings were flooded again.

1957: Sullivan Consolidated Mines Ltd. optioned the property and carried out an EM survey on

the property. Surface diamond drilling was done from the ice south of the main underground

workings.

1986: Maufort Resources Inc. acquired the property from Extension Holdings Inc. A total of 12

holes was drilled on the property for a total of 10,400 ft (3,169 m).

1987: Maufort Resources carried out a 56-hole drilling program on the property for a total of

81,361 ft (24,798.8 m). IP, Mag and seismic surveys were also carried out. The underground

workings were dewatered.

1993: Maufort Resources Inc. changed its name to Dynacor Mines Inc.

1999: As part of its Reorganization Agreement, Dynacor Mines transferred all its interests in the

Siscoe-Extension property (75%) to Wesdome Gold Mines Inc. on November 9, 1999.




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GEOLOGICAL SETTING AND MINERALIZATION

The Kiena Mine Complex area is located in northwestern Quebec (Figure 7-1), straddling the

southern part of the Abitibi Greenstone Belt (AGB) and the northern part of the Pontiac

Subprovince.

7.1 Regional Geology

7.1.1 Abitibi Greenstone Belt

The following description of the AGB is mostly modified and summarized from Monecke et al.

(2017) and references therein.

The Neoarchean AGB forms the northeastern portion of the Abitibi-Wawa Subprovince in the

southeastern portion of the Superior province (Figure 7-1). The southern Superior province

consists of a collage of E-trending Mesoarchean to Neoarchean terranes that underwent a

complex history of aggregation between 2720 and 2680 Ma (Percival, 2007). To the north, the

AGB is bounded by the Opatica Subprovince (Figure 7-1), a high-grade metamorphic terrain that

consists of tonalite, granodiorite, and granite intrusions, with minor outcrop areas of volcanic and

sedimentary rocks (Benn et al., 1992; Sawyer and Benn, 1993; Davis et al., 1994). Geophysical

constraints indicate that rocks of the Opatica Subprovince structurally underlie the supracrustal

rocks of the AGB (Benn and Moyen, 2008).




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Figure 7-1: Location of the Abitibi Greenstone Belt within the Superior province

Subdivision of Superior province from Thurston et al. (2008) and Stott et al. (2010);

modified from Monecke et al. (2017)

To the east and southeast, the AGB is truncated by the Mesoproterozoic Grenville front tectonic

zone, which is a southeasterly dipping zone of thrusts that juxtapose granulite facies metamorphic

rocks with low-grade of the AGB (Indares and Martignole, 1989; Daigneault et al., 1990; Culshaw

et al., 1997; Ludden and Hynes, 2000). To the southeast, the AGB is bounded by the Pontiac

Subprovince (Figure 7-2). Structural studies along the Abitibi-Pontiac contact indicate that the

AGB was thrust over the Pontiac Subprovince from the north (Camiré and Burg, 1993; Benn

et al., 1994; Daigneault et al., 2002; Bedeaux et al., 2017). To the west, the AGB is interrupted by

the 500-km-long NNE-trending Kapuskasing structural zone that exposes granulite facies

metamorphic rocks (Percival and West, 1994). The Kapuskasing structural zone is a W-dipping

thrust of Paleoproterozoic age along which Archean lower continental crust was upthrust (Percival

et al., 1989). The uppermost part of the stratigraphy of the Wawa Greenstone Belt (Williams et al.,

1991) to the west of the Kapuskasing structural zone is correlative with the AGB to the east

(Percival and Card, 1983; Ayer et al., 2010).




JUNE 2020 7-3

The AGB comprises E-trending successions of folded and faulted volcanic and sedimentary rocks

and intervening domes of intrusive rocks (Daigneault et al., 2004; Goutier and Melançon, 2007;

Thurston et al., 2008; Ayer and Chartrand, 2011). The volcanic successions in the AGB typically

have a steep dip and commonly young away from major intervening domes of intrusive rock

(Thurston et al., 2008). Submarine mafic volcanic rocks prevail, forming approximately 90% of the

outcrop area. Felsic volcanic rocks account for most of the remainder (Goodwin and Ridler, 1970;

Goodwin, 1977; Hannington et al., 1999), with komatiites forming a small but important part of

many of the volcanic successions (Imreh, 1984; Sproule et al., 2002; Houlé and Lesher, 2011;

Dostal and Mueller, 2013).

An important geologic feature of the AGB is the occurrence of major, E-trending ductile-brittle

fault zones. These zones cut across the entire belt from the Kapuskasing structural zone in the

west to the Grenville front in the east, dividing the supracrustal rocks and intervening domes into

distinct lozenge-shaped domains. The most two important fault zones in the southern AGB are

Destor-Porcupine fault zone (DPFZ) in the north and Larder Lake-Cadillac fault zone (LLCFZ) in

the south (Figure 7-2). These faults are subvertical (70°-90°) and dip either to the north or the

south. They have highly variable widths, ranging from tens to hundreds of metres (Poulsen,

2017), and are generally marked by intense ductile-brittle deformation and penetrative fabric

development. Most geologists agree that the fault zones are long-lived structures that controlled

sedimentation and volcanism in the AGB since at least 2679 Ma. (Dimroth et al., 1982; Mueller et

al., 1991, 1994; Cameron 1993; Mueller and Corcoran, 1998; Daigneault et al. 2002; Bleeker,

2012).

Based on recent geochronological information (Ayer et al., 2002b, 2005; Thurston et al., 2008),

six volcanic assemblages are distinguished that formed by submarine volcanic activity between

ca. 2750 and 2695 Ma (Figure 7-2). These assemblages spanned over 50 Ma years and are

listed below from oldest to youngest:

▪ Pacaud Assemblage (2750-2735 Ma);

▪ Deloro Assemblage (2734-2724 Ma);

▪ Stoughton-Roquemaure Assemblage (2723-2720 Ma);

▪ Kidd-Munro Assemblage (2719-2711 Ma);

▪ Tisdale Assemblage (2710-2704 Ma);

▪ Blake River Assemblage (2704-2695 Ma).




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Figure 7-2: Geologic map of the southern Abitibi Greenstone Belt and location of the Kiena Complex

LLCFZ = Larder Lake-Cadillac Fault Zone; DPFZ = Destor-Porcupine Fault Zone (modified from Thurston et al. (2008) and Poulsen (2017))




JUNE 2020 7-5

Volcanic rocks older than 2750 Ma locally are found in the AGB, as indicated by recent studies

southwest of Chibougamau, where 2795 to 2759 Ma volcanic rocks were mapped (Mortensen,

1993; Bandyayera et al., 2004; Davis and Dion, 2010; Leclerc et al., 2011, 2012)

In the southern Abitibi Greenstone Belt, the ≤2690 to ≤2685 Ma Porcupine assemblage consists

of flysch-like deposits that records sedimentation in a deep submarine environment (Rocheleau,

1980; Lajoie and Ludden, 1984; Stone, 1990; Born, 1995; Frieman et al., 2017). The Porcupine

assemblage locally includes minor calc-alkaline volcanic rocks (i.e., the Krist Formation in the

Timmins area; Ayer et al., 2002b). The ≤2679 to ≤2669 Ma Timiskaming assemblage in the

southern AGB is characterized by molasse-like clastic rocks deposited in a terrestrial setting

(Hewitt, 1963; Dimroth and Rocheleau, 1979; Hyde, 1980; Rocheleau, 1980; Mueller et al., 1991,

1994; Legault and Hattori, 1994; Born, 1995). The clastic deposits of this assemblage are locally

intercalated with predominately alkaline volcanic rocks (Cooke and Moorhouse, 1969; Mueller et

al., 1994; Wilkinson et al., 1999; Ispolatov et al., 2008).

The supracrustal rocks of the AGB are intruded by plutons of variable compositions and sizes.

Depending on emplacement age, several groups of plutons can be distinguished (Rive et al.,

1990; Feng and Kerrich, 1992a, 1992b; Sutcliffe et al., 1993; Mueller et al., 1995; Chown et al.,

2002; Beakhouse, 2011). Pre-2695 Ma intrusions in the southern AGB are commonly tonalitic to

granodioritic composition and their ages overlap with those of supracrustal rocks. The intrusions

are weakly to well foliated and complexly deformed, suggesting that they were folded together

with volcanic host-rock successions (Beakhouse, 2011). Pre-2695 Ma mafic to ultramafic

intrusions are found throughout the southern AGB. Many of these intrusions form sills or lenticular

units that crosscut stratigraphy at a low angle. Compositionally, the intrusions range from

peridotite to gabbro and diorite. A large number of intrusions in the southern AGB range in age

2695 to 2660 Ma, which broadly corresponds to the timing of sedimentation in the Porcupine and

Timiskaming successor basins. Intrusions of this age range are typically of grandioritic to granitic

and dioritic to quartz monzodioritic composition. Post 2660 Ma intrusions of granitic or

granodioritic compositions are rare in the southern AGB and mostly form part of large, multiphase

batholithic complexes (Beakhouse, 2011).

The greenstone belt is affected by a widespread greenschist facies metamorphism (Jolly, 1978;

Powell et al., 1993; Dimroth et al., 1983; Benn et al., 1994). The grade of metamorphism

increases to amphibolite at the fringes of some plutons and approaching the Pontiac and Opatica

Subprovinces or the Grenville front tectonic zone.




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7.1.2 Pontiac Subprovince

The Pontiac Subprovince (PS) (Figure 7-2 and Figure 7-3) consists principally of a turbiditic

succession composed of graywacke and mudstone with minor intercalated conglomerate and

basalt (Gunning and Ambrose, 1940; Imreh, 1976; Sansfaçon and Hubert, 1990; Fallara et al.,

2000; Pilote, 2013; De Souza et al., 2017). The metamorphic grade increases rapidly southward

in the Pontiac Group (Imreh, 1976; Card, 1990; Camiré and Burg, 1993; Benn et al., 1994), from

the biotite-chlorite zone along the southern contact of the Larder Lake-Cadillac fault zone, to

garnet and staurolite zones within about 2 km to the south (Sansfaçon and Hubert, 1990; Fallara

et al., 2005; Grant et al., 2005; Piette-Lauzière et al., 2014).

The minimum depositional age of the Pontiac Group is constrained by the crosscutting 2682

±1 Ma Lac Fournière Pluton (Fallara et al., 2000), whereas the maximum age is given by the

youngest detrital zircons from greywacke dated at about 2685 Ma (Davis, 2002; De Souza et al.,

2017).

7.2 Geology of the Kiena Mine Complex Area

The southern Abitibi Greenstone Belt in the Kiena Complex area consists of 2714–2700 Ma

volcano-plutonic assemblages, including the Malartic and Louvicourt groups, intruded by calc-

alkaline plutonic rocks (Figure 7-3). The Malartic Group comprises mainly komatiitic and tholeiitic

basalt flows and sills, with minor sedimentary rocks, which are interpreted as an oceanic floor in

an extensional environment related tomantle plumes, whereas the Louvicourt Group is mainly

composed of mafic to felsic volcanic rocks that formed in a subduction-related arc setting

(Desrochers et al. 1993; Daigneault et al. 2002; Scott et al. 2002). From south to north, the Kiena

Complex area is underlain by the lithologies of the Pontiac Group (PO), the Piché Group (PG),

the Cadillac Group (CG), and formations belonging to the Louvicourt Group and the Malartic

Group.

The following description of the Kiena Complex area geology is mostly modified and summarized

from Champagne et al. (2002), Champagne (2004), Scott et al. (2002), Olivo and Williams-Jones,

(2002), Scott (2005), Pilote et al. (1998a, 1998b, 1999, 2014a, 2014b, 2015a, 2015b, 2015c),

Pilote (2015a, 2015b), Monecke et al., 2017, and references therein.




JUNE 2020 7-7

Figure 7-3: Kiena Mine Complex property geology with historic and active mines and mineralized zones

(Adapted and modified from Pilote (2013, 2015a, 2015b))




JUNE 2020 7-8

7.2.1 Stratigraphy

Stratigraphic concepts were used to subdivide the E-Trending successions of folded and faulted

volcanic rocks from the AGB into formally named rock units (Figure 7-4). Stratigraphic criteria

have been successfully applied in mapping individual mining camp from the AGB. However,

stratigraphic correlation at the larger scale across and between mining camps is hampered by the

fact that boundaries between lithostratigraphic units of higher rank are often structural in nature.

In addition, the glacial cover is extensive in some areas, further complicating stratigraphic

correlation.

The AGB in the Val-d’Or–Malartic region has been divided into two stratigraphic groups based on

regional tectonics and volcano-sedimentary stratigraphy: the basal Malartic Group comprising the

La Motte-Vassan (LVF), Dubuisson (DF) and Jacola formations (JF), and the upper Louvicourt

Group comprising the Val-d’Or (VDF) and Héva formations (HF) (Figure 7-3).

Originally, the volcanic rocks in the Val-d’Or–Malartic region were assigned to the Malartic Group

(Gunning and Ambrose, 1940). Latulippe (1976) revised the stratigraphic nomenclature and

distinguished a Lower and Upper Malartic Group. According this stratigraphic nomenclature, the

La Motte-Vassan and Dubuisson formations form the Lower Malartic Group, whereas the Jacola,

Héva, and Val-d’Or formations represent the Upper Malartic Group. Based on a subsequent

revision of the stratigraphic nomenclature by Scott et al. (2002), the Malartic Group today

encompasses the La Motte-Vassan, Dubuisson, and Jacola formations. Scott et al. (2002) added

the Louvicourt Group in the recent stratigraphic nomenclature, which was further divided into the

Héva and Val-d’Or formations.

7.2.1.1 Pontiac Group (PO)

In the Kiena Complex area, the Pontiac Group (PO) covers the sector to the south of the

LLCFZ. The PO, located only in the Pontiac Subprovince, consists of turbiditic units (mostly

greywacke and siltstone) with rare monomict conglomerate intercalations intruded by thin

ultramafic units (Dimroth et al., 1982; Goulet, 1978, Ludden and Hubert, 1986). It is interpreted

as an accretionary prism created by the subduction of the Pontiac Subprovince under the

Abitibi Subprovince (Camiré and Burg, 1993; Card, 1990; Davis, 2002). It is contemporaneous

with the Cadillac Group with a maximum age between 2682 Ma and 2685 Ma (Davis, 2002;

Mortensen, 1993). The sedimentary strata of the Pontiac and Cadillac groups are interpreted

as synorogenic flysch-type assemblages (Mueller et al., 1996; Daigneault et al., 2002),

whereas Dimroth et al. (1982) considered the Pontiac Group to have been deposited in a

foreland belt.




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7.2.1.2 Piché Group (PG)

The Piché Group (PG), at the contact between the Cadillac and Pontiac sedimentary

successions, forms a relatively narrow (<2 km) band that marks the trace of the LLCFZ

(Latulippe 1976; Dimroth et al. 1982; Doucet and Lafrance 2005). The LLCFZ was then

delimited (Ambrose, 1944; Gunning, 1937; Gunning and Ambrose, 1940) based on the

presence of talc-schist and chlorite-schist correlated to the PG. The PG consists mainly of

highly strained and metamorphosed volcanic rocks (tholeiitic basalts, porphyritic andesites,

calc-alkaline tuffs, and komatiites) crosscut by felsic to mafic dykes and sills (Latulippe 1976;

Dimroth et al. 1982; Imreh 1984; Landry 1991; Beaudoin and Trudel 1989). However, it has

since been shown that the rocks of the LLCFZ were not uniformly affected due to

heterogeneities in the distribution and intensity of deformation. Primary textures, such as

spinifex and cumulates, have been preserved in areas where deformation is less intense.

These less deformed rocks are typically discontinuous and encompassed by bands of schists.

Latulippe (1976) proposed that the Piché Group be considered as a lithostratigraphic unit,

whereas Imreh (1984) proposed that the Piché Group be considered as a discordant unit.

The minimal age for the PG is constrained by a U-Pb age of 2708.8 ±1.0 Ma obtained from a

tonalite dyke that cuts the ultramafic units of the Buckshot pit near the Canadian Malartic

deposit (David et al., 2018).

7.2.1.3 Cadillac Group (CG)

The Cadillac Group (CG) (Figure 7-3) crops out extensively immediately to the north of the

LLCFZ, stretching from Rouyn Noranda in the west (Dimroth et al., 1982; Rocheleau, 1980) to

Val-d’Or (Figure 7-4) in the east (Pilote, 2015a). The CG (≤2687 Ma; Davis, 2002) consists of

sedimentary rocks including greywacke, pelitic schists, polymictic conglomerates, and iron

formations (Trudel et al. 1992). Daigneault et al. (2002) proposed that the sedimentary rocks

of the Cadillac Group represent an extensive volcano-sedimentary apron sequence that

straddles the LLCFZ.

The CG is a flysch-type sedimentary basin that rests unconformably over volcanic

assemblages (Mueller and Donaldson, 1992). The CG is a 150 km by 5 km basin located

along the LLCFZ to the north. The group is interpreted to be a lateral equivalent of the

Porcupine Group in Ontario (Ayer et al., 2002a; Thurston et al., 2008).

The sedimentary basin pinches out to a few hundred metres thick southeast of the Kiena

Complex (Figure 7-4). It is mostly composed of turbiditic sedimentary rocks with rare local

interlayering of polymictic conglomerates. The CG is identified based on its distinctive banded

iron formations (Dimroth et al., 1982). Deposition ages for sediments are estimated at

2687 Ma ±3 (Davis, 2002).




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7.2.1.4 Louvicourt Group (LG)

The Louvicourt Group (LG) is divided into the Héva (HF) and Val-d’Or (VDF) formations (Scott

et al., 2002).

Héva Formation (HF)

The Héva Formation (HF) can be as thick as 2.5 km and is composed of laterally extensive

massive to pillowed basalt flows, gabbroic sills and dykes, and minor felsic to mafic

volcaniclastic deposits (Scott et al., 2002). Volcaniclastic units are characterized by coarse or

fine tuff horizons with millimetre-scale laminations, intruded by gabbro. The volcanic rocks of

the HF have a U-Pb age of 2702 ±2 Ma (Scott et al., 2002). The HF belongs to the Blake River

assemblage.

Val-d’Or Formation (VDF)

The Val-d’Or Formation (VDF) has a stratigraphic thickness of 3 km to 5 km and is dominated

by massive to pillowed andesitic to rhyolitic lavas and associated volcaniclastic deposits.

Diorite intrusions represent a minor component. Andesites form up to 1-km-thick and 9-km-

long amalgamed flow units. Individual flow can be as thick as 80 m. The flows are intercalated

with amalgamated volcaniclastic beds 5 m to 40 m thick. The pillows exhibit a variety of forms,

from strongly amoeboid to lobed. Felsic lavas are laterally more restricted, but can be traced

for 1 km to 3 km along strike. Volcaniclastic units reach up to 40 m in thickness. The

volcaniclastic beds are composed of lapilli tuff, lapilli and blocks tuffs, and, to a lesser extent,

fine to coarse tuffs. The VDF yielded a U-Pb age of 2704 ±2 Ma (Scott et al., 2002). The VDF

belongs to the Blake River assemblage.

7.2.1.5 Malartic Group (MG)

The Malartic Group (MG) encompasses the La Motte-Vassan (LVF), Dubuisson (DF), and

Jacola (JF) formations (Scott et al.,2002).

Jacola Formation (JF)

The Jacola Formation (JF) has an apparent stratigraphic thickness of 1 km to 2 km and

consists of basalts and komatiites. Basaltic flows are typically non-vesicular, ranging from

massive to pillowed, and sometimes are in the form of flow breccias. Mafic volcaniclastic

deposits, primarily formed by quench fragmentation, are locally abundant. Individual basalt

units can be traced for distances of up to 5 km along strike. Individual komatiite units reach

thicknesses of 100 to 200 m and range from massive to pillowed. Occasionally, spinifex-

textured komatiite flows can be observed. The top of the JF has a U-Pb age of 2703.8

±1.3 Ma (Scott et al., 2002). The JF belongs to the Tisdale assemblage.




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Dubuisson Formation (DF)

The Dubuisson Formation (DF) (2708 ±2 Ma) consists mainly of pillowed and massive basalt

with various interbedded komatiitic flows (Imreh, 1980; Pilote et al. 1999). Ultramafic and

mafic flows are similar to those described in the LVF but in different proportions. The DF

belongs to the Tisdale assemblage.

La Motte-Vassan Formation (LVF)

The La Motte-Vassan Formation (LVF) crops out on the north side of Lac De Montigny.

Komatiites occur as sheet and tube-shaped flows that are intercaled with pillowed or massive

basalt flows (Latulippe, 1976; Imreh, 1984, Pilote et al., 2009). Komatiites are more abundant

than basalts (Imreh, 1980). A well-exposed example of spinifex-textured komatiite flow of the

Kidd-Munro assemblage is located at Spinifex Ridge north of Rivière-Héva (Champagne et al.,

2002; Champagne, 2004; Houlé et al., 2017). The komatiite succession at Spinifex Ridge is

interpreted to be age-equivalent to the 2714 ±2 Ma komatiite succession of the LVF at

Marbridge Ni deposit to the southwest (Pilote et al., 2009; Houlé et al., 2017).

The age of the LVF (2714 ±2 Ma) suggests it may be contemporaneous with the upper part of

the Kidd-Munro Assemblage (Figure 7-1). The LVF consists of komatiites, tholeiitic basalts

and magnesian basalts metamorphosed to amphibolite facies. The base of the sequence is

mostly represented by komatiites with some minor intercalated basalt. However, a decrease in

the proportion of komatiites is observed towards the top of the sequence (Imreh, 1984).

Komatiites are mainly found as two morphofacies: classic sheet flows with spinifex textures or

tube-shaped flows, or mega-pillows. The basalt flows are usually massive or pillowed (Imreh

1980).




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Figure 7-4: Stratigraphic chart of the Kiena Complex region with relevant isotopic ages

Stratigraphic relationships based on Dimroth et al., (1982); Mueller et al., (1996); Scott et al., (2002);

Tourigny et al., (1988). Figure from Bedeaux et al. (2017)




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7.2.2 Intrusive Rocks

Pre-2695 Ma mafic to ultramafic intrusions are found throughout the southern AGB. Many of

these intrusions form sills or lenticular units that crosscut stratigraphy at a low angle.

Compositionally, the intrusions range from peridotite to gabbro and diorite. In the Kiena Complex

area, the pre-2695 Ma Bourlamaque batholith (2699 ±1 Ma; Wong et al., 1991) is a good example

of synvolcanic intrusive rocks.

A large number of intrusions in the southern AGB range in age from 2695 to 2660 Ma, which

broadly corresponds to the timing of sedimentation in the Porcupine and Timiskaming successor

basins. Intrusions of this age range are typically of granodioritic to granitic and dioritic to quartz

monzodiorite composition.

In the Kiena Complex area, intrusions with ages between 2695–2685 Ma were emplaced

immediately prior to the deposition of the Temiscaming assemblage. A number of small

granodiorite, tonalite and monzonite range in age 2694 to 2685 Ma. This includes Camflo stock

(Jemielita et al, 1990), Kiena porphyry dykes (Morasse et al., 1995), Lamaque stock (Jemielita et

al., 1989), Norlartic dykes (Pilote et al., 1993), Sigma-Lamaque feldspar porphyry dykes (Wong et

al., 1991), and Snowshoe stock (Morasse et al., 1995).

Intrusions emplaced at 2679 to 2660 Ma formed during and immediately after the deposition of

the Timiskaming assemblage. Several small monzonite intrusions from the Malartic area yielded

U-Pb zircon and titanite ages of ~2678 Ma (De Souza et al., 2017).

Post-2660 Ma intrusions of granitic or granodioritic compositions are rare in the southern AGB

and mostly form part of large, multiphase batholithic complexes (Beakhouse, 2011). The La

Corne batholith is a good example and represents a large intrusion post-2660 Ma (Feng and

Kerrich, 1991; Powell et al., 1995; Ducharme et al., 1997).

The youngest igneous activity in the Kiena Complex area corresponds to the emplacement of

Proterozoic diabase dykes that cut across the Superior Province along a NE trend.

7.2.3 Structural Elements from Kiena Complex

Although the structural geology of individual mining camps is well established, there is no widely

agreed upon model for the structural evolution and generation of structures in the southern Abitibi

Greenstone Belt (Monecke et al., 2017). This is in part due to the highly variable quality of

exposures, strain heterogeneity, the lack of temporal constraints and, most importantly, variable

preservation of early structures in the different mining camps. Actually, the structural and tectonic

evolution of the southern AGB is not yet widely accepted.

Pilote et al. (2015c) established the nomenclature for the various structural elements in the Kiena

Complex area, as described below.




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The oldest regional schistosity is S1. It is systematically subparallel to bedding, S0. Within the

formations of the Malartic Group, the overall S1 trend is NW-SE. Both fabrics, S0 and S1, are

coplanar and show a moderate to steep dip to the north. S1 contains the primary stretching

lineation L1. In the southern and central parts of the Property, S0 and S1 are jointly folded into

Z-folds, with an average axial plane of N095°/85° and generally E-W axially planar cleavage (S2).

The axes of F1 and F2 folds are parallel to the plunges of the L1 stretching lineation contained in

S1.

A late S3 cleavage is the product of kinking and chevron folds in highly altered units showing a

strong pre-existing anisotropy. Dykes, mainly tonalite and monzonite, are deformed and affected

by S2. They trend to the SE, subparallel to the trace of the La Pause Fault. In places, they exhibit

a stretching lineation with a shallow westward plunge.

7.2.4 Large-scale Fault Zones

The Kiena Complex area has a series of large-scale shear zones and related subsidiary faults

trending ESE-WNW to SE-NW, subparallel to stratigraphy and dipping steeply to the north

(Figure 7-3). They are, from south to north: the Larder Lake-Cadillac Fault Zone (LLCFZ), the

Parfouru Fault (PF), the Marbenite Fault (MF), the Norbenite Fault (NF), the Callahan Fault (CF),

the K Shear Zone (KSZ) and the Rivière Héva Fault (RHF). The Kiena Complex area is cut by all

of them. Most faults at surface have been interpreted by geophysics and traced on high-resolution

aeromagnetic maps. Some faults correspond to a break in metamorphic grade.

The shear zones contain dykes or stocks of monzonitic or tonalitic composition that vary widely in

age (pre-, syn- or post-tectonic) and are spatially associated with gold mines (Norlartic, Marban,

Kiena, Sullivan, Goldex, Siscoe, Joubi, Sigma and Lamaque). The observed diversity in the styles

and ages of gold mineralization related to these large-scale shear zones demonstrates that

several distinct episodes of mineralization occurred.

7.2.4.1 Larder Lake-Cadillac Fault Zone (LLCFZ)

The LLCFZ in the AGB is a first order gold-bearing structure on a province-wide scale,

accounting for half of the gold production and reserves in Abitibi and more than 25% in

Canada (based on data by Dubé and Gosselin, 2007). The LLCFZ is a 250-km-long,

moderately to steeply dipping structure with a curvilinear trace (Figure 7-2 and Figure 7-3). It

dips northward or southward depending on location along its strike. A remarkable

characteristic of the LLCFZ is that over much of its 250-km length it marks the contact of a

persistent band of volcanic rocks, which in places is only 1 km to 100 m thick. Composed

mainly of ultramafic komatiite and tholeiitic basalt, the volcanic rocks provide sharp lithologic

contrasts with adjacent sedimentary units.




JUNE 2020 7-15

The LLCFZ in the Kiena Complex area is sharply defined as the contact between the

southward-facing volcanic successions of the Malartic Group and the younger folded, but

dominantly northward-facing, graywacke-mudstone successions of the Pontiac Group to the

south (Poulsen, 2017). The LLCFZ is confined to a 200-m-wide high-strain zone containing

thin, strongly deformed, but distinctive lithologic units. From north to south, these are

porphyritic diorite that intrudes mafic rocks of the Malartic Group, graywacke and mudstone

with lenses of conglomerate (assigned by most workers to the Cadillac Group), and felsic to

ultramafic volcanic rocks of the Piché Group (Robert, 1989). The ultramafic talc-chlorite-

carbonate schist of the Piché Group is compositionally a komatiitic to komatiitic basalt.

Spinifex textures are locally preserved (Robert et al., 1990). The anomalous high strain is

manifested by shape fabrics in primary lithic clasts and phenocrysts, intense schistosity, and

common minor folds that result in strong transposition of layers (Robert, 1989).

Located south of the Kiena Complex, the LLCFZ is generally oriented N110° and dips steeply

to the NNE.

7.2.4.2 Parfouru Fault (PF)

The PF is an ESE-WNW shear zone that dips steeply (75°) to the north or northeast and is

interpreted as an early synvolcanic structure (Daigneault, 1996; Bedeaux et al., 2017). The

shear zone can reach 300 m wide and has been traced for tens of kilometres.

7.2.4.3 Marbenite Fault (MF)

This major second order gold-bearing fault trends ESE-WNW to SE-NW and dips steeply to

the northeast. The MF was first observed at the Marban mine (Trudel and Sauvé, 1992; Sauvé

et al., 1993; and Beaucamp, 2010). Similar to the LLCFZ, the MF is characterized by the

presence of ultramafic talc-chlorite-carbonate schist that came from strongly deformed and

metamorphosed komatiitic basalt and/or komatiite. The MF is represented by an anastomosed

schist that sometimes is folded in some places. Thin mylonitic fabrics are also present within

the schist in some places.

In the Kiena Complex area, the MF is more than 100 m thick. It hosts the South Zone from

level 27 to level 32 and the S50 Zone from level 57 to level 100. A subsidiary folded fault,

located 400 m north of the MF, merges with the MF master fault at a depth of 500 m. This

subsidiary fault hosts a large part of the S50 Zone between level 12 and level 54.




JUNE 2020 7-16

7.2.4.4 Norbenite Fault (NF)

The Norbenite Fault (NF) is a major second-order gold-bearing fault that strikes WNW and

dips 40-60° to the northeast in the Norlartic Mine area where the fault was first observed

(Trudel and Sauvé, 1992; Sauvé et al., 1993). This fault also hosts the gold-bearing Kierens

Zone that was accessible by underground working from Norlartic mine.

The NF is 15 m to 110 m wide and has been traced for more than 10 km. The VC Zone at

Kiena occurs within this shear zone. The fault affects mainly the komatiitic units and

occasionally the basaltic units of the JF. It can be divided into two or three branches in some

places. Similar to the LLCFZ and the MF, the NF is characterized by the presence of

ultramafic talc-chlorite-carbonate schist that came from strongly deformed and

metamorphosed komatiitic basalt and/or komatiite.

7.2.4.5 Callahan Shear Zone (CSZ)

The CSZ strikes N090° and dips 60-80° to the north. The CSZ can reach up to 200 m in width

(Beaudoin et al., 1987).

7.2.4.6 K Shear Zone (KSZ)

The KSZ is a shear zone 300 m to 600 m wide that has been traced for more than 3 km. It

strikes N295° and dips 80° northeast. It is composed of talc and chlorite schists, actinolite

schists and minor sericite schists, and bodies of pure talc and massive actinolite (Olivo and

Williams-Jones, 2002; Olivo et al., 2007). The shear hosts the K Zone at the former Siscoe

mine and bounds the Main Zone to the south. The volcanic domain north of the KSZ is at

upper greenschist to amphibolite facies.

7.2.4.7 Rivière-Héva Fault (RHF)

The RHF is an 18-km-long ESE-WNW shear zone that dips steeply (80°) to the north or

northeast (Daigneault, 1996). The shear zone can reach 300 m wide and has been traced

over many kilometres. This structure corresponds to a change in metamorphic grade, from

greenschist facies in the south to amphibolite in the north.

7.2.5 Mineralization Types

The information of this section was taken from Turcotte et al. (2015) and Beausoleil et al. (2019).

Gold mineralization in the Property occurs in all rock types except the Proterozoic dykes but is

more common in intrusive bodies and basalt as these acted as competent rock units that

promoted fracturing during deformation. Gold mineralization is concentrated where there is a

marked competency contrast between these competent units and the adjacent deformed

komatiite and/or chlorite-talc schists.




JUNE 2020 7-17

According to Couture et al. (1994), there are at least two main gold mineralizing events in the

region: young deposits in which the gold mineralization did not experience much deformation after

its emplacement; and early mineralization in which mineralized bodies are commonly affected by

D1 asymmetric folds, are highly strained and are locally dismembered. In a few deposits, both

generations are present. Precise U-Pb zircon dating of an inter-mineral granodiorite dyke assigns

a minimum age of 2686 ±2 Ma to the gold mineralization at the Kiena mine (Morasse et al., 1995).

This age reveals that gold mineralization postdates volcanism and the Snowshoe plutonism but

predates regional syn-metamorphic deformation (ca. 2677-2645 Ma).

Gold-bearing veins in the region exhibit a great variety of orientations, mineralogy and

crosscutting relationships. For the purpose of this report, they are classified into the following

three main types:

▪ Type 1: early quartz-carbonate veins cut by various dykes;

▪ Type 2: deformed veins within a shear zone;

▪ Type 3: relatively weak deformed late quartz±tourmaline veins cutting all intrusive types and

previous gold-bearing vein systems.

On occasion, all three types may occur together.

At least 63 mineralized zones have been observed on the Property and are described in Turcotte

et al. (2015). The veins of these zones have been categorized as Type 1, 2 or 3 based on their

principal characteristics.

In general, mineralized zones on the Property occur near a large-scale fault. They are often

associated with a subsidiary shear zone that may be proximal, adjacent or host to the

mineralization. Alteration minerals are dominantly albite, carbonates and pyrite with lesser chlorite

and silica.

The gold occurrences found in shear zone settings are mainly restricted to competent units, and

thus the size and shape of the mineralized zones often depend on the size, shape and

concentration of the competent intrusive or basalt. Generally, it is possible to observe a flexure in

the shear zone.

In zones of structural dislocation, three settings for gold mineralization have been recognized:

▪ Shattered intrusive bodies, such as diorite or feldspar porphyry dykes, enclosed in talc-

chlorite schist;

▪ Zones of fracturing and brecciation in large bodies, such as basalt;

▪ Visible gold in quartz veins localized in schist and komatiite or in basalt and flow breccia near

the contact with the schist and komatiite.




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In large bodies of basalt, fracturing was generally restricted to narrow zones, and subsequent

mineralization resulted in narrow and often closely spaced mineralized zones. In narrower dykes,

the whole body is affected by fracturing, and subsequent mineralization was able to spread

throughout the dyke, forming large mineralized zones. Three factors control the size and shape of

mineralized zones associated with dykes in shear zone settings: the size of individual dykes, the

density of the swarms, and the shear zone folding intensity.

Quartz veins with visible gold are usually found in a folded shear zone (schist). Generally, the

veins seem to fill the available space created during the shear zone fracturing. The quartz veins

density clearly increased going toward the nose of the fold. In some areas, two or three zones are

stacked in the nose of the fold.

7.3 Mineralized Zones

Past gold production on the Property came from the Kiena, Siscoe and Shawkey mines. A

summary of the geological setting and mineralization is presented for each of these past

producers. The following sections are taken from Turcotte et al. (2015), and Beausoleil et al.

(2019).

7.3.1 Kiena Mine

A total of six zones were mined at Kiena. Five are aligned in an N-S corridor (from south to north):

South, S50, VC, North and 388. Those zones are situated in a sector were the Marbenite and

Norbenite faults are changing direction. Regionally, those faults are N290º but change direction to

N320º near those five zones. The sixth zone, Martin, is about 1.2 km east of the Kiena shaft.

7.3.1.1 S50 Zone

The S50 Zone was the main orebody at Kiena (Figure 7-5). It consists of a multistage

carbonate-albite-pyrite stockwork, breccia and replacement vein system (Morasse et al., 1995;

Morasse, 1998). It is located on the N-S limb of a local fold adjacent to the MF and is more or

less concordant with the upper contact of a moderately west-dipping tholeiitic flow with a

variably altered and schistose basaltic komatiite assigned to the JF. The mineralized zone

measures between 225 m and 600 m long, between 10 m and 50 m wide, and has a vertical

extent of 250 m to 1,000 m.




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Figure 7-5: Geological cross-section of S50 Zone

(Morasse et al., 1995)

The S50 Zone is associated with an intermediate to felsic dyke complex composed of a pre-

mineralization albitized diorite (albitite) dyke swarm and post-mineralization granodiorite and

feldspar porphyry dykes. It comprises Type 1 veins consisting of two mineralized types (from

oldest to youngest):

Carbonate (ankerite)-pyrite-Au replacement veins, also known as the “Breccia 1”

mineralization type.

Albite stockwork veins and breccias with disseminated pyrite, chalcopyrite, scheelite and

gold, also known as the “Breccia 2” mineralization type.




JUNE 2020 7-20

These Type 1 veins are cut by various dykes, including porphyry dykes that are weakly

mineralized by Type 3 calcite-quartz-pyrite-Au stockwork veins. The S50 zone was divided

into seven individual mineralized zones (A, B, C, D, J, K, L) based on the attitude and

composition of the mineralization and the presence or absence of a granodiorite dyke. These

mineralized zones can be regrouped into a high-grade core consisting of the A, B and D zones

and the lower part of the C, and a lower-grade mineralized shell comprised of the J, K and L

zones and the upper part of C (Figure 7-4). The C Zone is transposed within the MF.

Almost all zones in the MF are thin and elongated lenses and have a pitch angle of 60° to the

east, making an angle of around 30° with the down-plug axis of the fault plane. This probably

represents the stretching lineation in the fault zone.

7.3.1.2 S50 Deep Extension

Five mineralized lenses have been defined in the deep extension of the S50 Zone

(Figure 7-6). These lenses correspond to the extension at depth of the S50 Zone below level

96. They are named (from south to north) S50_104, S50_100, S50_101, S50_102 and

S50_103. The lenses are sheeted and parallel to, or contained within, the MF and are all part

of the S50 mineralized system. The five lenses are separated from each other by 5 m to 15 m

of less mineralized material. S50_102 and S50_103 are associated with a mylonite that

defines the hanging wall of the MF or its splay, near the contact between basalt and talc-

chlorite schist. S50_102 is in direct continuity at depth with the C Zone and is similar in

composition. The upper part of the S50_102 was clipped on the mined-out area starting below

the 94th level (940 m vertical) and has lateral and vertical extents of 250 m and 110 m,

respectively. The S50_103 has lateral and vertical extents of 110 m and 150 m, respectively.

S50_100, S50_101 and S50_104 occur at the eastern end of S50_102 in an area where the

MF widens and branches out. They are also spatially associated with the mylonite zone. They

have vertical extents between 430 m and 230 m and a horizontal extent of less than 150 m.

Their thicknesses vary from 3 m to 15 m. Host rocks are talc-chlorite-carbonate schists with

mineralized breccias typical of the S50 Zone. These breccias contains about 1-2%

disseminated pyrite and pyrrhotite. Albite and ankerite alteration are present in the S50_100

and S50_102 lenses. Sometimes visible gold in quartz veins is observed inside mineralized

breccia. Quartz veins with visible gold crosscut the mineralized and albitized breccia. Those

veins are similar to those observed in the Kiena Deep A Zone.




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Figure 7-6: S50 deep extension mineralized zones

(left: looking northwest; right: looking southwest)

7.3.1.3 VC Zone

The VC Zone is approximately 700 m north of the S50 Zone and is bounded to the north by

the Norbenite Fault. The following description of the VC Zone is mostly modified and

summarized from Beauregard and Gaudreault (2005). The VC Zone was mined by Wesdome

between 2006 and 2013. A total of 841,625 t of material with an average grade of 4.16 g/t Au

was mined from this zone.

The mineralized zone occurs as four vertical E-W pinch-and-swell lenses (VC1, VC2, VC3 and

VC4) in locally brecciated albitized basalt. The four lenses were mined over a vertical extent of

475 m. In cross-section, the lenses occur in a pseudo-en echelon pattern within a steep south-

plunging corridor 100 m across. In plan view, the individual lenses are 100 m to 180 m long

and average 8 m to 10 m wide, with a maximum width of 30 m. Gold mineralization is

associated with quartz veins and carbonate-quartz-pyrite veins (Turcotte et al., 2015),

presenting a similar mode of occurrence as seen in the S50 Zone.

In 2017, VC6 Zone was discovered when DDH 6187 intercepted a mineralized basalt and a

non-mineralized granodiorite dyke. DDH 6187 was drilled down-dip of the zone and returned a

value of 5.24 g/t on 48.7m.




JUNE 2020 7-22

7.3.1.4 VC1 and VC6

Four mineralized lenses were defined at depth in the extension of the VC1. The VC1_114 lens

is the continuity at depth of the VC1 lens and is interpreted to a depth of 315 m below the last

level mined (i.e., level 67). In plan view, the VC1_114 lens is 65 m long and averages 4 m to

6 m across. The VC1_114 lens is generally associated with a diorite dyke located within the

talc-chlorite schist. The dyke contains 1-5% pyrite and quartz-chlorite-pyrite veinlets ranging

from 2 mm to 5 cm. In places, quartz veins are present within the dyke or at its contact. Visible

gold is usually observed within the quartz veins and veinlet. The three other lenses are located

about 10 m north of the VC1_114 lens. VC1_111, VC1_112 and VC1_113 are sub-parallel to

the VC1_114 but shows a low dipping eastern mineralized shoot instead of a vertical

extension, based on preliminary understanding. The best results from VC1_114 were obtained

in DDH 6205 where a diorite dyke hosting quartz veins with visible gold returned a 7.5 m

interval grading 262.13 g/t Au or 17.7 g/t Au cut at 34.28 g/t (1 oz/t).

VC6_123 is located 150 m southeast of VC1_114. The mineralization is typical of the S50

Zone. The lens is elongated in shape, plunging to the south-west and measuring 185 m high

by 80 m wide with thicknesses ranging from 5 m to 12 m. The lens is oriented NE, parallel to a

20 m thick chlorite-carbonate shear zone that bounds the mineralization to the northwest. It is

hosted in a sheared basalt and in minor porphyritic or dioritic dykes. The main alteration

minerals are chlorite and carbonates, with minor patches of albite, silica and/or amphibole

alteration. Pyrite and pyrrhotite account for less than 3% of the rock. Locally, visible gold was

observed in quartz-carbonate veinlets. This zone is crosscut by a massive, slightly sheared

and barren granodiorite dyke. The dyke was possibly deformed at the same time as the

volcanic host rocks. Figure 7-7 presents the VC1 and de VC6 mineralized lenses.

Figure 7-7: VC1 and VC6 mineralized zones

(left: looking west-southwest; right: looking south-southwest)




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7.3.1.5 North Zone

The North Zone (formerly the No. 1 Zone) was first encountered by surface drilling in 1938.

The zone is 900 m north of the S50 Zone and 200 m north of the NF (Figure 7-3). It consists of

two parallel lenses shaped like flattened tubes, each measuring 10 m by 60 m to 90 m in plan

view and plunging 70° north, which can be traced over a length of at least 520 m. From 2007

to 2013, Wesdome extracted 458,860 t of material from the North Zone with an average grade

of 2.44 g/t Au.

The main structural control for the gold mineralization is a sodium-rich diorite intrusion in

basalt host rock. Mineralization occurs as a quartz-carbonate-sulphide matrix-filling in

brecciated and albite-altered diorite, and as quartz-carbonate-pyrite veins and veinlets in

basalt wall rock (Turcotte et al., 2015). These veins and veinlets are predominantly developed

in a steeply dipping orientation that is suggestive of a sheeted or ribbon-vein system. There

are subsidiary veinlets at oblique and sub-random orientations, which may be described as a

pseudo-stockwork. Well-developed albitized envelopes with disseminated pyrite encompass

the quartz-carbonate veins, and it is common to observe 5-10% disseminated pyrite where

intense and closely spaced veining is developed. The North Zone is cut by unmineralized

granodiorite and feldspar porphyry dykes.

7.3.1.6 388 Zone

The following description of the 388 Zone is mostly modified and summarized from Laplante

(2000). The 388 Zone has been defined as a small deposit parallel to the North Zone at a

distance of 300 m to the north (Figure 7-3). It was mined by Wesdome from 2007 to 2013. A

total of 146,268 t of material was mined from the zone at an average grade of 3.01 g/t Au.

The 388 Zone has a vertical extend of 100 m. Gold mineralization is associated primarily with

a single major quartz vein 2 m wide that developed in albitized basalt alongside a diorite dyke.

The vein is rather erratic and structurally meanders within an irregular zone of altered basalt

3 m to 6 m wide. Small scattered quartz-carbonate veinlets with minor pyrite occur in the wall

rock.

7.3.1.7 South Zone

The South Zone lies 100 m to the south of the S50 Zone in the MF deformation corridor. It

strikes northwest and dips 65° northeast. Discovered in 1981 (hole S-129), the South Zone

includes the hanging wall and footwall lenses, which are 18 m apart. They were mined on

three levels only (between levels 27 and 32) over a vertical extent of 55 m. They are parallel

and concordant to the enclosing volcanic host rocks, predominately basalts with minor diorite

dykes and komatiitic ultramafic flows. The basalts are strongly chloritized and carbonatized

with local silicification and brecciation (quartz-carbonate veining) accompanied by low

concentrations of pyrite (5%). The mineralization is similar to that observed in the S50 Zone.




JUNE 2020 7-24

The South Zone is generally strongly fractured by the fault zone. Wesdome tried to mine this

zone in 2012 and 2013, but after producing 37,076 t with an average grade of 1.74 g/t Au,

production was abandoned due to dilution problems.

7.3.1.8 ZS-130, ZS-131, ZS-132, ZS-133 and ZS-135

Five mineralized lenses have been defined in the South Zone area (Figure 7-8). Wesdome

extended the footwall and hanging wall lenses up-plunge to near-surface and 200 m down-

plunge, and these extensions are now named the ZS_131 and ZS_132 lenses, respectively.

ZS_130 is located less than 20 m south of ZS_131 and is sub-parallel to it. The maximum

lateral extension of the ZS_130 is 500 m, but both ends are poorly drilled up to date. ZS_135

is a small lens intercalated between ZS_130 and ZS_131, with lateral extents of 150 m and

vertical extents of 175 m. The thickness of the alteration and mineralization envelope that

encloses the four lenses ranges from 25 m to 60 m. A fifth lens (ZS_133) was defined near the

contact with the S50 Zone, in a distinct band of basalt. The ZS_133 is located 25 m north of

ZS_132 and has a vertical extent of 285 m with thicknesses ranging from 3 m to 6.5 m.

Figure 7-8: Zone South mineralized zones

(left: looking northwest; right: looking west-southwest)




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7.3.1.9 Kiena Deep A Zone

The first hole that intersected the Kiena Deep A Zone was 4320 collared from level 91 in

December 2007. This hole cut a quartz vein containing about 10 specks of visible gold within a

talc-chlorite schist. A value of 28.23 g/t Au over 1 m was obtained. In January 2010, two more

holes (4928 and 4929) were drilled in the Kiena Deep A Zone from level 91. Hole 4928

returned an average grade of 12.31 g/t Au (8.42 g/t Au cut) over 12.8 m. Hole 4929 cut an

average grade of 25.42 g/t Au (14.73 g/t Au cut) over 10.5 m. In 2012, four more holes were

drilled from the same level with the following best results: hole 5965B with 66.77 g/t Au (11.04

g/t Au cut) over 3 m, hole 5966 with 51.86 g/t Au (21.49 g/t Au cut) over 7.5 m, hole 5967 with

10.17 g/t Au (8.20 g/t Au cut) over 5.4 m, and hole 5974 with 42.49 g/t Au (8.80 Au g/t cut)

over 14.9 m. The six subsequent holes also intersected the Kiena Deep A Zone and cut gold-

bearing quartz veins and veinlets hosted by a talc-chlorite schist, like in hole 4320. In June

2016, Wesdome tried again to explore the area of the Kiena Deep A Zone.

The Kiena Deep A Zone is localized within the Marbenite Fault deformation corridor. Contrary

to previous beliefs, the Kiena Deep Zone is not the extension of the S50 Zone. The structures,

mineralization type, host rocks and grades are completely different. The MF plane at that

depth (1,350 m below surface) strikes NNW and dips shallower (55°) to the east compared to

the WNW direction near the surface and the dip of 70°. The geology and structures were

interpreted in 3D by Ravenelle (2018) based on drillhole data and reviewed by Pierre-Luc

Richard in 2019.

The Kiena Deep A Zone is divided into three main lenses: ZA, ZA1 and ZA2. A fourth smaller

lens (H1ZA) is located in the hanging wall of the ZA at the contact between the basalt and the

flow breccia units (Figure 7-9). The lenses occur along an isoclinal fold associated with the MF

and a subsidiary fault. All lenses in the Kiena Deep A Zone are variably altered to chlorite,

carbonate and amphibole. Pyrite content is less than 2% with traces of pyrrhotite and

chalcopyrite. According to Ravenelle (2018), two styles of gold mineralization occur in the

zone; the first style largely dominates Kiena Deep and the second style is seen occasionally:

Laminated veins (shear veins) hosted within sheared ultramafic rocks (grade is up to

several ounces of gold per ton). Folding is observed in shear zones. Veins are composed

of vitreous to milky quartz with carbonates. There is more than one gold bearing vein

generation. Sulphides range from traces to 1% locally (pyrite, pyrrhotite, chalcopyrite,

sphalerite and galena).

Stockworks of quartz-carbonate veins and veinlets hosted in basalt (like the S50 style of

gold mineralization) typically yielding less than 10 g/t Au.




JUNE 2020 7-26

The main lenses ZA, ZA1 and ZA2 are coplanar to the MF to the SW, conforms to the fold

hinge in the middle part and is sub-vertical to the NE. The axial plan of the isoclinal fold

undulates and plunges abruptly to the ESE (Figure 7-10). Lens ZA is located at the contact of

the basalt domain and the chlorite-carbonite schist. We generally observe a feldspar porphyry

dyke in the hanging wall of the ZA, which is a guide for the interpretation. This dyke seems to

have undergone the same deformation as the host schist. The ZA has a horizontal footprint of

150 m and a vertical extent of 460 m with thicknesses ranging from 3 m to 18 m.

ZA1 and ZA2 lenses are sub-parallel to the ZA and are entirely in the chlorite-carbonate schist

of the fault along with minor intermediate dykes. ZA1 shows a vertical extent of 550 m and a

horizontal footprint of 120 m with thicknesses ranging from 3 m to 8 m. ZA2 is less drilled than

ZA and ZA1 and is actually separated in two domains by a low-grade area. The lower part of

the ZA2 has a much higher grade than the upper part and is also thicker with thicknesses of

up to 15 m in the lower part. The overall vertical extent of the ZA2 is 590 m. The lenses ZA1

and ZA2 still have to be defined on the sub-vertical contact. Definition drilling is ongoing in this

area.

Figure 7-9: Kiena Deep A zones folded within the schist

(left: looking northwest; right: looking south-southwest)




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Figure 7-10: Typical cross-section of the Kiena Deep A Zones (1615NE)

7.3.1.10 Kiena Deep B Zone

The Kiena Deep B Zone consists of the ZB_140 lens (Figure 7-11). It is the only mineralization

at Kiena hosted in the footwall of the MF. The shape, elongation and plunge of the zone are

similar to that of S50 zones at depth in the MF. The lens is stacked 25 m below the S50_100

lens in basalt. Alteration consists of chlorite, amphibole and carbonate, with local biotite and

silica. The pyrite content is less than 2%. High-grade intervals are associated with quartz-

carbonate-albite-tourmaline extension veins and veinlets.




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Figure 7-11: Kiena Deep B Zone

(left: looking north; right: looking southwest)

7.3.2 Siscoe Mine

The geology of the Siscoe mine is predominantly felsic intrusive rock of the Siscoe Stock and

tholeiitic basalts with minor intercalations of ultramafic lavas belonging to the DF. The contact

between the Siscoe Stock and the DF is marked by the K Shear Zone. Most of the producing

veins are located in the stock (85-90% of total production), although minor amounts of gold were

also extracted from the shear zone and from small veins hosted in volcanic rocks of the DF. Type

1, 2 and 3 veins were documented in the most productive gold vein systems of the mine. Type 1

veins correspond to the Siscoe Main Zone that produced almost half of the total production from

the mine, and to the Siscoe, Hope, F and G veins. Type 2 veins correspond to the dismembered

quartz veins of the K Zone. Type 3 veins correspond to the Siscoe C Vein, the first producer at

the mine that yielded some 40,000 oz of gold, and to the Siscoe No. 27 Vein.

7.3.3 Shawkey Mine

The rocks that underlie the Shawkey mine area belong to the Jacola Formation. The

characteristic lithologies found on the Property consist of ultramafic to mafic flows at the base of

the sequence, followed by massive and minor pillowed basalts, basaltic flow and pillow breccia,

tuff breccia and basaltic tuffs. The Shawkey No. 1 Vein is located about 250 m northeast of the

NF in a secondary shear called the Martin Shear Zone. This shear zone is filled by a quartz vein

(the Main Vein) accompanied by quartz stringers (Type 2 veins) in the walls. The Main Vein is

hosted by altered basalts cut by dioritic dykes.




JUNE 2020 8-1

DEPOSIT TYPES

Gold mineralization on the Kiena Mine Complex shares many geological attributes with other

vein-type gold deposits of the Val-d’Or district and with orogenic gold deposits (also known as

lode gold, greenstone-hosted quartz-carbonate vein, or mesothermal deposits) in terms of host

rock composition, mineralogy and hydrothermal alteration.

The degree of fracturing was the primary control on the formation of mineralized zones on the

Property. Thus, gold mineralization is mainly hosted in fractured competent units that acted as

fluid conduits and precipitation sinks, both during and after deformation. Alteration, notably

albitization, likely played a key role in host unit competency. The competency contrast between

intrusive bodies, or basalts, and the talc-chlorite schists may be responsible for strain localization

at the rheological boundary, and it induced a secondary permeability that provided greater access

to hydrothermal gold-bearing fluids during episodic shear zone movements.

The presence in the same area of more than three types of gold-bearing veins exhibiting a wide

range of orientations, mineralogy and crosscutting relationships, and the fact that several

generations of dykes and veins are involved, suggests that gold mineralization was the product of

multiple mineralizing phases (Beausoleil et al., 2019).

8.1 Archean Greenstone-Hosted Orogenic Gold Deposits

The Kiena Mine Complex mineralization presents characteristics of typical Archean greenstone-

hosted orogenic gold deposits. The following description is taken from Simard et al. (2013) unless

specified otherwise.

Greenstone-hosted quartz carbonate vein deposits occur in deformed greenstone belts of all ages

elsewhere in the world; especially those with variolitic tholeiitic basalts and ultramafic flows

intruded by intermediate to felsic porphyry intrusions, and sometimes with swarms of albitite or

lamprophyre dykes (Dubé and Gosselin, 2007).

Archean greenstone-hosted orogenic gold deposits are typically distributed along first-order

compressional to transpressional crustal-scale fault zones (Figure 8-1), characterized by several

strain increments (e.g., Cadillac–Larder Lake Fault Zone) that mark the convergent margins

between major lithological boundaries, such as volcano-plutonic and sedimentary domains.

Large-scale carbonate alteration is also commonly distributed along those major fault zones and

associated subsidiary structures (Dubé and Gosselin, 2007). This gold deposit type is, however,

seldom located within these first-order structures. Major, or first-order faults are interpreted as

primary hydrothermal pathways to higher crustal levels (Eisenlohr et al., 1989; Colvine, 1989;

McCuaig and Kerrich, 1998; Kerrich et al., 2000; Neumayr and Hagemann, 2002; Kolb et al.,

2004; Dubé and Gosselin, 2007); however, only a few significant gold deposits are hosted in

major faults such as the Ajjanahalli mine, Dharwar Craton, South India (Kolb et al., 2004), and the

McWatters mine and the Orenada deposit, Abitibi Subprovince, Canada (Robert, 1989; Morin

et al., 1993; Neumayr et al., 2000; 2007). Significant mineralized quartz veins are commonly




JUNE 2020 8-2

hosted in second- and third-order shear zones (Eisenlohr et al., 1989). Structurally, these shear

zones vary from brittle–ductile to ductile, depending on their depth of formation (Hodgson 1993;

Robert and Poulsen, 2001). They are formed at intermediate depths ranging from 5 km to 10 km

(Dubé and Gosselin, 2007). At depths greater than 10 km, quartz veins are seldom located within

shear zones whereas gold mineralization is mostly associated with disseminated sulfides (Witt

and Vanderhor, 1998).

At the deposit scale, the nature, distribution and intensity of the wall-rock alteration is largely

controlled by the composition and competence of the host rocks and their metamorphic grade.

Typically, the alteration haloes are zoned and characterized at greenschist facies by iron-

carbonatization and sericitization, with sulphidation of the immediate vein selvages (mainly pyrite,

less commonly arsenopyrite) (Dubé and Gosselin, 2007).

The main gangue minerals are quartz and carbonate with variable amounts of white micas,

chlorite, scheelite and tourmaline. The sulphide minerals typically constitute less than 10% of the

mineralized material. The main minerals are native gold with pyrite, pyrrhotite and chalcopyrite

without significant vertical zoning. The mineralization is syn- to late-deformation and typically

post-peak greenschist-facies or syn-peak amphibolite-facies metamorphism (Dubé and Gosselin,

2007).

There is a general consensus that the greenstone-hosted quartz-carbonate vein deposits are

related to metamorphic fluids from accretionary processes and generated by prograde

metamorphism and thermal re-equilibration of subducted volcano-sedimentary terranes. The

deep-seated gold transporting metamorphic fluid has been channelled to higher crustal levels

through major crustal faults or deformation zones. Along its pathway, the fluid has dissolved

various components, notably gold, from volcano-sedimentary packages, including a potential

gold-rich precursor. The fluid is then precipitated as vein material or wall rock replacement in

second and third order structures at higher crustal levels through fluid pressure cycling processes

and temperature, pH and other physico-chemical variations (Dubé and Gosselin, 2007).




JUNE 2020 8-3

Figure 8-1: Inferred crustal levels of gold deposition showing the different types of gold deposits and the inferred deposit clan

(from Dubé et al., 2001; modified from Poulsen et al., 2000)

8.2 Gold Mineralization in the Val-d'Or District

The following is taken from Couture et al., (1994), who published a detailed description and

chronology of the Archean greenstone-hosted quartz carbonate vein of the Val-d’Or district.

Gold mineralization occurs in all rock types but is more commonly located within intrusive bodies

that acted as competent rock units promoting fracture during deformation. In the Val-d'Or district,

there are two main generations of gold quartz veins: young deposits in which the gold

mineralization did not experience much deformation after its emplacement; and early

mineralization in which mineralized bodies are commonly affected by D2 asymmetric folds, are

highly strained and locally dismembered. In a few deposits both generations are present.




JUNE 2020 8-4

At least two major gold mineralizing events have been recognized in the Val-d’Or district on the

basis of morphological and structural features, mineralization and alteration mineral assemblages,

and crosscutting relationships with dated intrusive rocks (Robert, 1990, 1994; Sauvé et al., 1993;

Couture et al., 1994). The older mineralizing event is manifested by veins and breccias (e.g.,

Norlartic, Marban, Kiena mines, and Main mineralized zone at Siscoe mine) that are mainly

associated with second-order shear zones and commonly folded or boudinaged by D1

deformation. These veins and breccias are cut by diorite and tonalite dykes, which have U-Pb

zircon ages of 2692 ±2 (Pilote et al., 1993) and 2686 ±2 Ma (Morasse et al., 1995). The younger

gold event, which produced the Sigma, Lamaque, Perron-Beaufor, Shawkey, Wesdome and

Camflo deposits, as well as the C Vein (quartz-tourmaline) at the Siscoe mine, is represented by

veins commonly associated with third-order shear zones. These veins clearly crosscut plutonic

rocks intruded between 2694 ±2 Ma (Wong et al., 1991) and 2680 ±6 Ma (Jemielita et al., 1990),

and they may have formed during the last stages of D1 deformation.

Young gold mineralization is characterized by networks of shear-hosted quartz-

carbonate±tourmaline±scheelite veins and associated subhorizontal extension veins. This is well

documented at the Sigma mine and also occurs in other deposits east of Val-d'Or, namely

Lamaque, Perron, and L.C. Béliveau. Mineralized veins and associated structures crosscut all

rock types except Proterozoic dykes. In the Sigma deposit (Robert and Brown, 1986) the gold-

bearing quartz-tourmaline vein system is hosted by andesite of the Val-d'Or domain (2705 ±1 Ma,

Wong et al., 1991; 2706 ±3 Ma, Machado et al., 1991), porphyritic diorite, and feldspar porphyries

(2704 ±3 Ma and 2694 ±2 Ma, respectively, Wong et al., 1991) metamorphosed to greenschist

facies. The porphyritic diorite is deformed, but feldspar porphyry dykes cut D2 folds and thus

postdate regional D2 folding (Robert and Brown, 1986). The vein network consists of coeval and

cogenetic steeply dipping shear-hosted veins and sub-horizontal extensional veins (Robert and

Brown, 1986; Figure 8-2). Preserved delicate vein-filling textures and crosscutting relationship

indicate that gold mineralization postdates the youngest intrusion as well as metamorphism and

much of the deformation (Robert and Brown, 1986). Rutile associated with the mineralization has

been dated by U-Pb at 2599 ±9 Ma (Wong et al., 1991). Similar vein geometry and morphology

were also described in the Lamaque mine (Daigneault, 1983), where most of the mineralization is

hosted by small circular tonalite plugs crosscutting porphyry intrusion similar to that of Sigma.

Jemielita et al. (1990) reported U -Pb ages of 2685 ±3 and 2682 ±2 Ma for the Lamaque Main

tonalite plug and 2593 ±5 Ma for rutile associated with gold mineralization. Similar age

relationships can be inferred from structural studies at the Perron and Béliveau mines (Tessier,

1990; Gaumond, 1986, respectively). West of Val-d'Or, significant gold was extracted from the

post-D2 Camflo quartz monzonite dated at 2680 ±6 Ma (Jemielita et al., 1990) and 2685 ±10 Ma

(Zweng et al., 1993), whereas titanite and rutile, associated with the gold mineralization, yield

ages of 2625 ±7 Ma (Jemielita et al., 1990) and 2621 Ma (Zweng et al., 1993).




JUNE 2020 8-5

Figure 8-2: Schematic diagram of the geometric relationships between the structural elements of veins and shear zones and the deposit-scale strain axes (Robert, 1990; Modified after Dubé and Gosselin, 2007)




JUNE 2020 9-1

EXPLORATION

This chapter presents the exploration work performed by Wesdome on the Project since

November 2018.

9.1 Surface Exploration

No surface exploration work has been carried out.

9.2 Underground Exploration

In August 2017, Wesdome developed an exploration ramp to provide additional underground

drilling platforms. The exploration ramp allowed shorter holes to be drilled with better angles and

accelerated access to the Kiena Deep zones.

Ramp development started at the 100 level. The CMAC-THYSSEN Mining Group Inc. was

contracted to perform the underground development. The first drilling bay was completed in

October 2017 and diamond drilling began on the Kiena Deep Zone.

From August 2017 to November 2018, a total development of 2.2 linear kilometres allowed

diamond drilling to be done from a more optimal direction (i.e., to the northwest) to intersect the

steeply plunging zone to the southeast and provide drilling platforms that allowed definition drilling

in the central area of the Kiena Deep A Zone. The enhanced drill platforms also made it easier to

drill several step-out exploration holes.

In 2019, development has continued with a new exploration ramp on level 79. Totalling 575.5 m

of development, the 79 Level Ramp has been completed in early 2020. It provides optimal drill

platforms for testing the possible up-plunge extension of the Kiena Deep A Zone and extensions

of VC Zones between the 670-m level and the 1,070-m level.




JUNE 2020 10-1

DRILLING

This chapter presents the drilling program carried out by Wesdome between October 12, 2018

and January 15, 2020 (the “2018-2020 Program”) on the Kiena Mine Complex.

10.1 Drilling Methodology

The underground drilling program was performed by Forage Orbit Garant Inc. (Orbit-Garant).

Drilling was conducted with NQ caliber (47.6 mm core diameter) and telescoped with BQ caliber

when rock quality was poor (i.e., faults, shears, schist). Electric diamond drill rigs were used,

starting with four rigs in October 2018 and increasing to a total of seven rigs in January 2020. The

drill types and models were as follows:

▪ One YU1800 (125 hp): max 1,200 m (NQ); max 1,600 m (BQ);

▪ Two YU1200 (100 hp): max 750 m (NQ); max 1,000 m (BQ);

▪ Two YU615 (100 hp): max 400 m (NQ); max 600 m (BQ);

▪ Two B15 rigs (100 hp): max 450 m (NQ); max 650 m (BQ).

Every hole was drilled with maximum stabilization using two hexagonal core barrels and a 36"

shell until reaching the schistose unit.

10.1.1 Drillhole Location/Set-up

Diamond drillholes (DDH) for the 2018-2020 Program were planned using vertical cross-sections

and plan views with the aim of improving the accuracy of the interception angle in mineralized

zones.

The coordinate system in use was a Local Mine Grid.

The software used were GeoticGraph, AutoCad and Promine. Hole collars were implanted by a

surveyor who drew a line between the front side spad and a backside spad. The drillers aligned

themselves according to the line and started the hole at the most suitable place. After drilling,

collars were initially surveyed by contract surveyors, and later by a Wesdome technician. Collar

azimuth and dip were measured when possible.

10.1.2 Drillhole Orientation during Operation

The drillhole orientation is checked and monitored using a downhole surveying device as follows:

▪ REFLEX EZ-Trac instrument is used to conduct deviation surveys. Single-shot

measurements are taken every 50 m until reaching the schistose unit, at which point the

stabilization tools are removed;




JUNE 2020 10-2

▪ Multi-shot measurements are taken at every 3 m while the rods are pulled out;

▪ If ground conditions in the schistose area are deemed acceptable, single-shot

measurements are resumed;

▪ At the end of the hole, a final multi-shot survey is performed with measurements every 3 m

while the rods are pulled out until reaching the end of the first multi-shot survey or the

beginning of the schist.

The REFLEX instrument is handled by the drilling contractor who transfers the multi-shot data to

the Wesdome geology department. The single-shot data is transcribed and provided in paper

format to Wesdome geologists.

10.1.3 Drilling

Recovered drill cores by the drilling contractor are in NQ and BQ size. Drillholes are generally

started in NQ caliber in mafic rocks. When the drillhole reaches the komatiites, schists or a major

fault, BQ caliber is used, otherwise NQ rods get stuck at the bottom of the hole. The core is

collected in a standard drilling tube and the driller’s helper carefully places the core into wooden

core boxes at the drill rig. He/she also marks the depth (m) after each 3 m runs with wooden

blocks and wraps the boxes with tape once it is full. Core trays are numbered with a permanent

marker indicating the drillhole number and the sequential box number.

Generally, the drillhole is abandoned in the komatiites or schists. Occasionally, it is possible to

pierce and stop the drillhole in the mafic units; if this occurs, the drillhole is terminated by the

geologist.

Once the drillhole is terminated and the final downhole survey reading is collected, the drill crew

pulls the rods for mobilization to the next drill site. Every hole is systematically grouted with

cement. The collars of underground holes are identified by a conical plug with a metal tag

displaying the hole number. Sometimes it is not possible to complete the multi-shot survey due to

the rock and structures encountered (faults, high fracturing, etc.).

10.1.4 Core Logging and Measurement

In the core shack, Wesdome employees remove the tape and place the boxes on the logging

tables. The geologists rotate the core so that all pieces are fitted together, showing a cross-

sectional view. They verify that distances are correctly indicated on the wooden blocks placed

every 3 m. The core is then measured and the boxes are labelled.

Wesdome geologists log and record the data using GeoticLog software. Lithologies (principal and

secondary), alteration, mineralization, veins, structures, magnetism, samples and assay results

are compiled in the database.




JUNE 2020 10-3

10.1.4.1 Core Recovery

The core recovery is calculated by measurement in centimetres of core in the core tray divided

by the centimetres claimed to be drilled on the meterage blocks. This number multiplied by

100 is recorded as percent recovery. Core recovery is recorded for each drill run. Specific

areas of loss are noted, if possible, and marked by placement of a wooden marker and the

estimated loss. The ideal core recovery is 100%; however, it is not always possible due to

ground conditions or sometimes loss of drill core during the coring process, e.g., grinding, etc.

For the 2018-2019 DDH included in the MRE, the average core recovery was 98%.

10.1.4.2 Rock Quality Designation

The rock quality designation (RQD) is designed to give qualitative and quantitative information

on the stability of rock surrounding and included in mineralized material. This information is

used to determine the mineability and rock control procedures that will be required to extract

the mineralized material.

RQD is a quantitative index of rock quality based on a core recovery procedure in which the

core recovery is determined by incorporating only those pieces of hard, solid core longer than

twice the diameter of the core. For NQ core, the nominal diameter is 5 cm, so the length index

is 10 cm; shorter lengths of core are ignored. RQD is determined for each core run as these

are the only definitively known distance markers. RQD is determined using the following

formula for each core run:

RQD (%) = 100 x the sum of the length of the core pieces equal to or longer

than 10 cm / Core run length

It is important to distinguish between mechanical breaks and natural breaks identified in the

core.

RQD is valid for solid core only and should not be used for very poorly disaggregated

materials such as highly weathered rock, clays or un-cemented aggregates.

The average RQD is 81% based on 13,401 measurements for the 2018-2019 DDH included in

the MRE.

10.1.4.3 Core Photography

Once logged by the geologist, all drill core is photographed wet, four boxes at a time. The

objective of core photos is to have a digital image recorded with sufficient details to clearly see

core features prior to destructive sampling procedures. This record can be used later to qualify

rock quality features and to examine core images against geological logging if the core is

unavailable for examination. The photos are also used, as required, during the construction of

geological sections.

Once the core is photographed, it is assigned to the core saw operator for splitting and

sampling.




JUNE 2020 10-4

10.1.5 Core Storage

After the sampling process, the core boxes are stored under roofed racks in the outdoor core

storage area, which is enclosed with secure fencing. Every box is labelled with an aluminum tag

displaying the hole number, the box number and the From-To meterage. An Excel spreadsheet

serves as an inventory of box location in the core storage area.

10.2 Recent Diamond Drilling

As of August 6, 2019 (closeout date of the MRE database), Wesdome had completed 140 new

DDH during the 2018-2019 campaign on the Property for a total of 36,050 m (Table 10-1). Of

these 140 new drillholes (Appendix B), 66 are in Kiena Deep A Zones, resulting in the inclusion of

an additional 18,365 m of drill data (Figure 10-1).

As of March 24, 2020, Wesdome had completed an additional 164 DDH for 47,861 m that are not

included in the herein MRE (Appendix B).

Since October 2018, the close out date for the December 2018 MRE, five underground drill rigs

have continued to operate in order to verify the up-plunge extension (VC6 and VC1 zones), test

the down dip extension, and perform infill drilling in the Kiena Deep A Zones. Underground drill

rigs were active at level 67 (one drill rig) and in the new exploration ramp (four drill rigs).

In December 2019, a sixth drill rig was added at the 79 level. In January 2020, a seventh drill rig

was added at the 79 Level. At the effective date of the report, seven underground drill rigs have

continued to operate in order to verify the up-plunge extension (VC6 and VC1 zones), test the

down dip extension, and perform infill drilling in the Kiena Deep A Zones. Underground drill rigs

were active at 79 Level ramp (two drill rigs) and in the exploration ramp (five drill rigs).

Table 10-1: Summary of the drilling completed on the Property during the 2018-2019 Program

(included in the MRE)

Year Drillholes count Total length (m)

2018-2019 140 36,050




JUNE 2020 10-5

Figure 10-1: Location of drillholes throughout the Property with their status




JUNE 2020 10-6

Figure 10-2: Kiena Mine Complex diamond drillhole locations, close-up view of the 2019 MRE zone




JUNE 2020 11-1

SAMPLE PREPARATION, ANALYSES, AND SECURITY

11.1 Wesdome Data

The origin of Wesdome Gold Mines on the property can be traced to Western Quebec Mines Inc.,

incorporated in 1945 when developing the Dorval-Siscoe property. Commercial production started

in 2006 and temporary shutdown occurred on June 30, 2013.

Since February 2019, Wesdome has ceased sending samples to Actlabs Laboratory and has only

used ALS in Val-d'Or.

11.1.1 Core Handling, Sampling and Security

The drill core is boxed and sealed at the drill rigs and transported by the drillers to the

underground station and brought to surface via the service cage at the end of each shift. A

technician from Wesdome takes over the core handling and brings it to the core shack. After

being logged and sampled, individual sample bags are placed in rice bags along with the list of

samples. QA/QC samples are prepared and bagged ahead of time by Wesdome personnel and

are batched at the core shack following the geologist’s instructions. Batches have been shipped

daily to the ALS Global labs facility in Val-d’Or, Quebec since December 7, 2018. Batches and

shipments contain variable numbers of samples. During the period of October 2018 to January

2019 samples were also shipped to Actlabs in Val-d’Or.

11.1.1.1 Gold Assays Samples

To create representative and homogenous samples, sampling honours lithological contacts,

i.e., no sample crossed a major lithological boundary, alteration boundary or mineralization

boundary.

Sampling intervals are determined by the geologist during logging and marked on the core

boxes or on the core itself using coloured lumber pencils with a line drawn at right angles to

the core axis. Sample lengths typically range from 0.5 m to 1.50 m with a preferred length of

1.0 m for the mineralized zones. The sampled core is considered representative. Two

shoulder samples, each having a sample length of approximately 1.0 m to 1.5 m, are collected

from the non-mineralized core above and below the mineralized intervals.

Samples are numbered in consecutive order utilizing sample tag books containing numerical

sequences of 50 pre-labeled triplicate water durable sample tags (three tags per sheet). The

first of the tags remained with the sample tag book as an archival record of the samples ’

parameters. The second tag is used to indicate the position of the sample in the core box. This

is a permanent sample reference that will remain on the wooden core box. The third and last

tag is inserted inside the sample bag. From each sample sheet, the last two tags are

separated from the page and tucked under the core at the beginning of each sample by the

geologist.




JUNE 2020 11-2

The sample sequence includes blank samples, duplicate samples and Certified Reference

Materials (CRMs) that are inserted into the sample stream using sample numbers that are in

sequence with the core samples. A CRMs sample, consisting of material of known metal

content and internationally recognized and verified, is included in the sample sequence by the

trained core sampler. A “blank” sample is material technically devoid of any metals. Blanks

and CRMs are stored in a designated secure area in the core shack. There is never any

written reference to the location of any control samples on sample bags, sample tags or

dispatch documentation for the assay lab.

Once logged and labelled, the core of each selected interval is sawed in half using a typical

table-feed circular rock saw. The core saw operator, trained in core cutting procedures,

executes the core cutting at the Wesdome core shack. The logging geologist has already

clearly marked out all pertinent cores for cutting and sampling. The core is sawn in half, along

its length, with a diamond saw. One half (consistently from the same half of the split core) is

put into the plastic sample bag and the other half is retained and kept in the core box for later

reference. The paired sample tags are then torn with one tag stapled to the core box at the

start of its sample interval and the other tag placed into the sample bag with the core sample.

When cutting, the core saw operator looks for visible gold inside the veins and reports it to the

geologist when positively observed.

The sample tag number is also written on the outside of the sample bag using a permanent

marker. The bag will then be closed using a zip tie and stored in sequence prior to sample

dispatch preparation. Sample bags are packed in large ‘rice’ bags sealed with a zip tie that is

‘broken’ or opened at the assay laboratories only.

The range of sample numbers inside the bag is written on the rice bag. The sealed rice bags

are stored in the core shack in Val-d’Or until shipping to the laboratories. For the 2018 to 2019

drilling campaigns, the samples were transported by a Wesdome employee to Actlabs and

ALS Global labs in Val-d’Or, where the samples were prepared and processed.

11.1.1.2 Core Density Samples

New specific gravity (SG) tests were conducted on 37 samples. SG was measured by water

displacement method at the ALS laboratory using the OA-GRA08 ALS method (see method

below).

About 0.10 m to 0.20 m of core was selected for each density measurement. The dry mass

was measured followed by the submerged mass. Both measurements were recorded and the

density was measured using the following formula:

𝑆𝑝𝑒𝑐𝑖𝑓𝑖𝑐 𝐺𝑟𝑎𝑣𝑖𝑡𝑦 =𝑀𝑎𝑠𝑠𝐷𝑟𝑦 𝑖𝑛 𝑎𝑖𝑟

(𝑀𝑎𝑠𝑠𝐷𝑟𝑦 𝑖𝑛 𝑎𝑖𝑟 − 𝑀𝑎𝑠𝑠𝑆𝑢𝑏𝑚𝑒𝑟𝑔𝑒𝑑 𝑖𝑛 𝑤𝑎𝑡𝑒𝑟)⁄




JUNE 2020 11-3

11.1.2 Methods of Preparation, Processing and Analysis

11.1.2.1 Lab Accreditation and Certification

ALS and Techni-Lab (now Actlabs) have the ISO/IEC 17025:2005 accreditation through the

ALA (Canadian Association for Laboratory Accreditation Inc.) They are both independent

commercial laboratories.

11.1.2.2 ALS Sample Analysis Procedure

At ALS laboratories, samples are sorted, bar-coded and logged into the ALS Webtrieve

program. Damaged samples are documented and Wesdome personnel are informed.

Samples are dried to constant weight and weighted (WEI-21). The sample is then crushed to

P70 2,000 µm (CRU-31). A split is collected using a riffle splitter (SPL-21) and a reject

duplicate split is prepared from that original sample (SPL-21d). A pulverization split of 250 g is

then prepared for both the original and duplicate split (PUL-31; PUL-31d) at P85 75 µm. A pulp

duplicate is also prepared from the original sample (SPL-34). When a metallic sieve analysis

is conducted (Au-SCR21), a pulverization of 1,000 g P95 106 µm is done (PUL-35a).

Samples are then analyzed by fire assay (FA) with atomic absorption (AA) spectroscopy from

30 g pulps (Au-AA23). The lower detection limit is 0.005 g/t. When assay results are higher

than 3 g/t, the sample is re-assayed with a gravimetric finish (Au-GRAV21) on a 30 g pulp. If

results are higher than 10 g/t, gravimetric finish is done and the Metallic sieve method

(Au-SCR21) is also conducted. In this case, 1,000 g is pulverized and screened to 100 µm.

Duplicate assay is done on screen undersize and the entire oversize fraction is assayed.

Results are provided through a secure server and downloaded, by the geologist in charge of

the project, in Excel format and the official certificate (sealed and signed) in PDF format.

As part of ALS internal quality control program, four QA/QC samples are inserted by ALS per

batch of 24 samples (one blank, two standards and one pulp duplicate). A method blank and

certified reference material is applied and reported for each furnace load to monitor the fire

assay process. A duplicate crushed sample is drawn at random and assayed for each work

order to monitor precision.




JUNE 2020 11-4

11.1.2.3 Actlabs Sample Analysis Procedure

Once the samples are received at the Actlabs facility, they are sorted, bar-coded and logged

into the Actlabs LIMS program. Damaged samples are documented and Wesdome personnel

are informed with photographs. Samples are dried at 60°C, crushed to P80 passing 10 mesh,

and split into 250 g to 300 g using a Jones riffle splitter. The sub-sample is pulverized to P80

passing 200 mesh. Samples are analyzed by FS with AA spectroscopy from 30 g pulps. The

lower detection limit is 0.01 g/t. When assay results are higher than 3 g/t Au but lower than 10

g/t Au, core sample pulps are re-assayed by FA with gravimetric finish, while sample results

higher than 10 g/t Au are rerun with the metallic sieve (MS) method. In this case, 350 g is

pulverized and assayed.

Results are provided through a secure server and downloaded, by the geologist in charge of

the project, in Excel format and the official certificate (sealed and signed) in PDF format. As

part of Actlabs internal quality control program, four QA/QC samples are inserted by Actlabs

per batch of 24 samples (one blank, two standards and one pulp duplicate).

11.1.3 Sample Shipping and Security

The following procedures are applied to ensure a safe and secure management of materials and

data as it pertains to core samples at the Kiena Mine Complex:

▪ All core samples submitted for preparation and analysis to the laboratories are secured in

rice bags with zip ties and collected directly at the core shack by the laboratory under the

supervision of a member of Kiena’s team;

▪ The lab is notified by email that the samples are en route and is instructed to notify Kiena’s

geologists when the samples arrive at the prep lab;

▪ The sample shipment contains a sample submittal form as well as a sample dispatch list

detailing the security tag number, rice bag number and the number of samples contained in

each rice bag;

▪ The sample submittal form and sample dispatch list are electronically transmitted to the

laboratories once the shipment has left the core shack;

▪ Samples are sent to:

Actlabs

1960, 3e avenue

Val-d'Or, QC J9P 7B2, Canada

ALS

1324, rue Turcotte

Val-d’Or, Qc, J9P 3X6

▪ Results are downloaded by Bruno Turcotte, Senior Project Geologist for Wesdome, via a

secure server, as Excel files;

▪ QA/QC data is evaluated before the samples are integrated into a master database;




JUNE 2020 11-5

▪ The core boxes are stored under roofed racks in the outdoor core storage area enclosed by

secure fencing. The exact location of each hole in the outdoor core library is recorded in an

Excel spreadsheet for future reference;

▪ The sample pulps and rejects are stored at the Kiena Mine Complex.

11.2 Quality Assurance and Quality Control (QA/QC)

Canadian National Instrument 43-101 (NI 43-101) Standards of Disclosure for Mineral Projects

requires mining companies reporting results in Canada to comply with the CIM Best Practice

Guidelines. The guidelines describe the elements required in the reports, but do not provide

guidance for Quality Assurance and Quality Control (QA/QC) programs.

QA/QC programs have two components: Quality Assurance (QA) deals with the prevention of

problems using established procedures, while Quality Control (QC) aims to detect problems,

assess them and take corrective actions. QA/QC programs are implemented, overseen and

reported on by a Qualified Person (QP) as defined by NI 43-101.

QA programs should be rigorous, applied to all types and stages of data acquisition and include

written protocols for: sample location, logging and core handling; sampling procedures;

laboratories and analysis; data management; and reporting.

QC programs are designed to assess the quality of analytical results for accuracy, precision and

bias.

The materials conventionally used in mineral exploration QC programs include standards, blanks

and duplicates. Definitions of these materials are presented hereunder:

▪ Standards are samples of known composition that are inserted into sample batches to

independently test the accuracy of an analytical procedure. They are acquired from a known

and trusted commercial source. Standards are selected to fit the grade distribution identified

in the Kiena mineralization;

▪ Blanks consist of material that is predetermined to be free of elements of economic interest

to monitor for potential sample contamination during analytical procedures at the laboratory;

▪ Duplicates are samples submitted to assess both assay precision (repeatability) and to

assess the homogeneity of mineralization. Duplicates can be submitted from all stages of

sample preparation with the expectation that better precision is demonstrated by duplicates

further along in the preparation process;

As per NI 43-101, quality control samples were inserted into the sample batches sent to the

laboratory. Inserts included pulp duplicate samples, blank samples and standards.

Table 11-1 summarizes the QA/QC samples submitted to the laboratories along with routine drill

core samples.




JUNE 2020 11-6

Table 11-1: Samples submitted to the laboratories for analysis during the 2018 to 2019 drilling campaigns

Type of sample Quantity %

Primary drill core samples 22,614 77.9

Field blanks 1,823 6.3

CRMs 1,757 6.0

Coarse duplicates 1,007 3.5

Pulp duplicates 1,830 6.3

TOTAL 29,031 100%

11.2.1 Duplicates

Duplicate samples are submitted to assess both assay precision (repeatability) and to assess the

homogeneity of mineralization.

Coarse duplicates consist of second splits of crushed material. This material will then need to be

pulverized.

Pulp duplicates consist of second splits of prepared samples ready to be analyzed and are

indicators of analytical precision, which may also be affected by the quality of pulverization and

homogenization.

As part of the Wesdome QA/QC program, the laboratory assayed one coarse duplicate for every

20 samples. Coarse duplicate started within the first batch of samples sent to ALS. The QA/QC

program also included one pulp duplicate for every 20 samples. Figure 11-1, Figure 11-2 and

Figure 11-3 show the scatterplots of the various duplicate for each laboratory. The correlation

coefficient varies from 90.5% to 94.0%. The results show relatively good reproducibility,




JUNE 2020 11-7

Figure 11-1: Zoomed in scatterplot with linear trend of the coarse duplicates and original samples’ results from ALS for the 2018-2019 drilling program (n=1,611)

y = 0.9661x + 0.0097R² = 0.8948

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Original Sample

Duplicate

N = 1,611




JUNE 2020 11-8

Figure 11-2: Zoomed in scatterplot with linear trend of the pulp duplicates and original samples results from ALS for the 2018-2019 drilling program (n=1,613)

y = 0.8395x + 0.0273R² = 0.9135

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N = 1,613




JUNE 2020 11-9

Figure 11-3: Zoomed in scatterplot with linear trend of the pulp duplicates and original samples’ results from Actlabs for the 2018-2019 drilling program (n=822)

(Three higher grade samples are not shown on the scatterplot)

11.2.2 Blanks

Blanks are used to monitor for potential sample contamination that may take place during sample

preparation and/or assaying procedures at the laboratory. Sample of barren crushed white marble

(blank) were used by Wesdome.

Note that only the results from the drillholes used for the MRE are presented herein. Additional

QA/QC measures from post-MRE drillholes were reviewed by the QP and deemed acceptable. In

some cases, investigations and/or re-assays may be necessary before using them for a future

MRE.




JUNE 2020 11-10

One blank sample was inserted for every 20 samples. According to Wesdome’s QA/QC protocol,

if any blank yields a gold value above 0.1 g/t Au, all samples from the 20 samples batch should

be reanalyzed. Of the 1,823 blanks analyzed, four blank samples failed the protocol, which

represents 0.2%. Figure 11-4 shows the results of the blank material used during the 2018-2019

Program on the Project.

Figure 11-4: Results for blanks used by Wesdome during the 2018-2019 drilling program

Detection limit for Actlabs was 0.01 g/t and for ALS 0.005 (g/t)

(802 samples assayed by Actlabs and 1,021 samples assayed by ALS, both by fire assay with atomic absorption finish)

Generally, the blank can indicate contamination at the laboratories. The re-assay of one batch

was requested following a failure for sample Y703875. ALS laboratory was called and

explanations were asked. Following the re-assay, the conclusion showed that the inversion of a

sample tag was probably the cause of the failure.




JUNE 2020 11-11

11.2.3 Certified Reference Materials (Standards)

Accuracy and precision are monitored by the insertion of Certified Reference Materials (CRMs). A

suite of commercially available CRMs is used on the Project. One sample was inserted for every

20 samples. Table 11-2 shows the CRMs used for the 2018-2019 Program.

Note that only the results from the drillholes used for the MRE are presented herein. Additional

QA/QC measures from post-MRE drillholes were reviewed by the QP and deemed acceptable. In

some cases, investigations and/or re-assays may be necessary before using them for a future

MRE.




JUNE 2020 11-12

Table 11-2: Standard Certified Reference Materials used at the Kiena Mine Complex during the 2018-2019 drilling campaign

Standard (CRMs)

Method Laboratory Certified gold

value (g/t) Quantity inserted

Mean grade (Au g/t) w/o

gross outliers

Lower process limit

(mean - 3SD)

Upper process limit

(mean + 3SD)

Failed (outliers)

Gross outliers

(%) passing QC

SJ80 AA ALS 2.656 284 2.580 2.485 2.827 30 5 89.4

SJ80 AA Techni-Lab 2.656 267 2.603 2.485 2.827 22 0 91.8

SK94 AA ALS 3.899 243 3.793 3.647 4.151 22 2 90.9

SK94 AA Techni-Lab 3.899 254 3.895 3.647 4.151 13 0 94.9

SL76 AA ALS 5.960 353 5.790 5.384 6.536 11 5 96.9

SL76 AA Techni-Lab 5.960 31 5.911 5.384 6.536 0 0 100.0

SN75 AA ALS 8.671 43 8.105 8.074 9.268 8 0 81.4

SN75 AA Techni-Lab 8.671 12 8.516 8.074 9.268 1 0 91.7

SJ95 AA ALS 2.789 50 2.756 2.627 2.951 5 2 90.0

SK109 AA ALS 4.102 19 4.083 3.850 4.354 0 0 100.0

SN91 AA ALS 8.679 13 8.365 8.097 9.261 1 0 92.3

SK94 Grav ALS 3.899 71 3.939 3.647 4.151 22 0 69.0

SL76 Grav ALS 5.960 93 5.933 5.384 6.536 7 0 92.5

SN75 Grav ALS 8.671 13 8.450 8.074 9.268 3 0 76.9

SN91 Grav ALS 8.679 11 8.476 8.097 9.261 1 0 90.9

Total 1,757 146 14 91.7




JUNE 2020 11-13

The selection of the CRMs was based on anticipated cut-off grade and the 10 g/t re-analysis

metallic sieve grade.

CRMs were considered failed by Wesdome when a gold result exceeded three standard

deviations (±3 SD) beyond the expected value. If the analytical value exceeded the ±3 SD control

limits, systematic re-assaying was not always requested. For each failed standard, Wesdome QC

protocol stipulates that only samples included with mineralized zone material should be

reanalyzed. If there is no significant gold result within the batch of 20 samples, no re-assay is

ordered. During the 2018-2019 drilling programs, 146 CRMs representing 8.3% failed.

Considering the low failure rate, the location of these failed CRMs, and the actions taken when

such failures occurred, the QP is of the opinion that the failed CRMs are not material for the

purpose of this MRE and show the natural statistical spread in the data.

11.2.4 Check Assays

As part of the Wesdome QA/QC program, rejects from a selection of samples were submitted to a

second laboratory to assess both the assay precision (repeatability) and the homogeneity of

mineralization. Rejects Check Assays consists of the resulting crushed material from the original

assay procedure being sent to a different laboratory. This material will then need to be pulverized

and assayed by the second laboratory.

In this case, ALS (Val-d’Or) was the original laboratory and Actlabs (Val-d’Or) was chosen as the

secondary laboratory.

Figure 11-5, Figure 11-6 and Figure 11-7 show the scatterplots of the various assay methodology

for each laboratory. The overall correlation coefficient is low, but overall it is typical of high nugget

effect mineralization. It is noteworthy to observe that metallic sieve shows a significantly better

reproducibility than gravimetric finish and that the latter shows a significantly better reproducibility

than atomic absorption finish. Based on these observations, metallic sieve should continue to be

used and perhaps even used for lower grades; currently samples above 10 g/t Au.




JUNE 2020 11-14

Figure 11-5: Scatterplot showing results from the check assay program for the metallic sieve methodology

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Au (g/t) - Original Assay (ALS)

Scatterplot of the check assays duplicatesMetallic Sieve (MS)

Duplicate

N =214




JUNE 2020 11-15

Figure 11-6: Scatterplot showing results from the check assay program for the gravimetric finish methodology

Note that Wesdome’s protocol ensures that samples returning grades above the 10 g/t Au threshold using the atomic gravimetric finish methodology are re-assayed using the metallic sieve methodology.

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Scatterplot of the check assays duplicatesGravimetric Finish

Duplicate

N =190




JUNE 2020 11-16

Figure 11-7: Scatterplot showing results from the check assay program for the atomic absorption methodology Note that Wesdome’s protocol ensures that samples returning grades above the 3 g/t Au threshold using the

atomic absorption methodology are re-assayed using the gravimetric finish methodology.

11.3 Conclusion

The QP, Pierre-Luc Richard, P. Geo., reviewed the sample preparation, analytical and security

procedures, as well as insertion rates and performance of blanks, standards, duplicates, and

check assays for the 2018-2019 drilling programs and concluded that the observed failure rates

are within expected ranges and that no significant assay biases are present. According to the

QP’s opinion, the procedure and quality of the data are adequate to industry standards and

support the Mineral Resource Estimate.

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Au (g/t) - Original Assay(ALS)

Scatterplot of the check assays duplicatesAtomic Absorbtion (AA) Finnish

Duplicate

N =1,634




JUNE 2020 12-1

DATA VERIFICATION

The Mineral Resource Estimate (MRE) in this Report is based on drill data from several eras of

drilling on the Kiena Mine Complex that includes holes completed in the 1940s.

For the purpose of this MRE, Pierre-Luc Richard, P. Geo., performed a basic verification on the

entire Project database. All data was provided by Wesdome in a local mine coordinate system.

The database close-out date for the resource estimate is August 6, 2019; data from 6,616 DDH

(976,170.3 m) was incorporated in the resource estimate block model area. The last hole included

in the database was #6533.

The overall Kiena Mine Complex project database contains 8,413 diamond drillholes (DDH) at the

MRE Report effective date.

12.1 Site Visit

Pierre-Luc Richard, P. Geo. and Qualified Person, and Charlotte Athurion, P. Geo., both of BBA,

visited the Kiena Mine Complex from August 6-8, 2019. The site visit included a visual inspection

of historical core and core drilling in progress, a field tour, an underground visit and discussions of

the current geological interpretations with Wesdome geologists.

The site visit also included a review of sampling and assays procedures (Figure 12-2), the QA/QC

program, downhole survey methodologies, and the descriptions of lithologies, alteration and

structures (Figure 12-1).

Pierre-Luc Richard, P. Geo. also visited the site on a few other occasions during the course of the

mandate to exchange ideas with the on-site geologists.

Charlotte Athurion, P. Geo., visited the Wesdome office several times during the mandate in order

to support Wesdome’s geologist with the geological interpretation and modelling.

12.2 Sample Preparation, Analytical, QA/QC and Security Procedures

Wesdome procedures are described in Chapters 10 and 11 of the current Report. Discussions

held with on-site geologists confirmed that the procedures were adequately applied.

Pierre-Luc Richard reviewed sections of mineralized core while visiting the Project (Figure 12-1).

All core boxes were labelled and properly stored (Figure 12-3). Sample tags were present in the

boxes and it was possible to validate sample numbers and confirm the presence of mineralization

in witness half-core samples from the mineralized zones.

The historical data used in this MRE was taken before the implementation of the National

Instrument NI 43-101. Little information is available about sample preparation, analytical, QA/QC

or security procedures. However, it is assumed that exploration activities conducted in the past

were in accordance with prevailing industry standards at the time.




JUNE 2020 12-2

Figure 12-1: Core review in the core logging facility during the August 6-8, 2019 site visit

A) Core logging facility with Wesdome’s employees describing core; B) Core box close-up showing a mineralized interval with associated sample tag

Figure 12-2: Sampling procedures review during the August 6-8, 2019 site visit

A) Core saw used to sample the core; B) Sample bags after being sawn; C) Standards used for QA/QC; D) Sample batches ready to be sent to the Laboratory.




JUNE 2020 12-3

Figure 12-3: Storage review during the August 6-8, 2019 site visit

A) Pulps safely stored on the property; B) Additional pulps and rejects; C) Outdoor core storage facility on the property.

12.2.1 Drillhole Location

For drilling conducted between 2018 and 2019, drill collars have been surveyed using a Leica

TS16. The Kiena Mine Complex surveyor collar data (Excel spreadsheet) was compared with the

collar from the Geotic database for discrepancies. For the historical data, approximately 10% of

the drillholes intersecting mineralized zones of this MRE were checked against the historical

paper logs.

12.2.2 Downhole Survey

Downhole survey data for the 2018 to 2019 drilling programs were checked for discrepancies.

Spurious measurements are tagged by the Wesdome geologist as “false” in the database and are

not considered by the software for the modelling. For the historical data, surveys of approximately

10% of the drillholes intersecting mineralized zones of the MRE were checked against the

historical paper logs at the Kiena Mine Complex.




JUNE 2020 12-4

12.2.3 Assays

Clovis Auger, P. Geo., from BBA was granted access to the original assay certificates directly

from ALS for all holes drilled by Wesdome (2018-2019 Program) on the Project. The original logs

and the assay certificates for the historical DDH were also available at the Kiena Mine Complex

office. All the results from the assays from ALS were verified for the 2018-2019 Program. Assays

for approximately 10% of the historical DDH intersecting the current MRE mineralized zones were

also verified. The assays recorded in the database were compared to the original certificates from

the different laboratories. Minor discrepancies were noted and modified for the historical DDH.

In the assay table, the gravimetric finish result always replaces a value obtained by AA finish and

when a sample was assayed using the metallic screen procedure, the value recorded as “Au (g/t

AVG)” always corresponds to the gold value obtained by metallic sieve method. Values lower

than the detection limits were set to zero (0). With AA and gravimetric finish, when a sample has

a reject duplicate, the average of the assay results is used in the Wesdome database.

12.3 Underground Voids

Underground workings were imported from the Wesdome working files. Previous MRE voids were

validated in the 2018 MRE and the robustness of the 3D shapes were also validated. The

underground voids include shafts, drifts, raises, stopes and the exploration ramp developed since

2017.

Pierre-Luc Richard considers that the precision and details from the voids to be acceptable and

reliable for the current estimate. Uncertainties remain, especially concerning the stopes. In order

to address the uncertainty, all block grades were initialized to zero within a 5-m buffer around

existing stopes, using a clipping boundary. Uncertainties do not have a significant effect on the

MRE.

12.4 Conclusion

Pierre-Luc Richard, P. Geo., is of the opinion that the drilling, sampling and assaying protocols in

place are adequate. The database for the Kiena Mine Complex is of good overall quality. In the

QP’s opinion, the Project database is appropriate to be used for the estimation of Mineral

Resources.




JUNE 2020 13-1

MINERAL PROCESSING AND METALLURGICAL TESTING

This chapter summarizes metallurgical testwork carried out for the Project in 2018 and 2019.

Section 13.1 presents the series of laboratory testwork carried out in 2018 by the Centre

Technologique des Résidus Industriels (CTRI).

In 2019, a new metallurgical testwork program, based on the CTRI testwork program observations

and the former operation, was designed and managed by BBA (Section 13.2). The testwork was

performed by SGS Lakefield (SGS) between August and November 2019.

13.1 CTRI 2018 Metallurgical Tests

13.1.1 Testwork Program

In 2018, Wesdome mandated CTRI (Noël, 2019) to perform some preliminary gold leaching tests

on mineralized material from the Kiena Deep A and S50 Zones. The objective of the testwork was

to benchmark the leaching performance of the Kiena Deep A Zone to the previously mined S50

Zone. Wesdome selected and prepared the samples used for this testwork campaign. It was not

possible for CTRI to confirm the samples’ representativeness of the deposit. Fifteen 48-hour

cyanidation tests in 4 L bottles on gold mineralized material were performed.

13.1.2 Sample Preparation

The mineralized material received included 57 sub-samples for a total of 92 kg. Following a

discussion with Wesdome representatives, the 57 samples were divided to produce five

composites: four for the Kiena Deep A Zone and one for the S50 Zone (Table 13-1). It is important

to mention, to avoid confusion with the 2019 Metallurgical testwork, that the Kiena Deep A Zone

composites used by CTRI are currently classified as Kiena Deep A and A1 Zones.

Table 13-1: Composite splits and weights

Composite Weight (kg)

Kiena Deep A Zone – 1 16.2




S50 Zone 23.7




JUNE 2020 13-2

The 57 sub-samples were assayed for gold individually. Following discussions between Wesdome

and CTRI representatives, a blending sample recipe was prepared. Table 13-2 presents the

calculated gold head grade for each composite. Final composites were assayed for gold head grade

validation (Table 13-2). All assays were performed at the ALS geochemistry laboratory located in

Rouyn-Noranda (Quebec).

In the recipe, each composite listed in Table 13-1 was prepared selecting individual samples from

the 57 sub-samples. The samples included, in a given composite, were blended, homogenized

twice (with a splitter) and then bagged into 1 kg charges. The resulting composite charges were

assayed for gold and used for direct cyanidation (in triplicate) for each composite listed in

Table 13-2.

Table 13-2: Composite feed assays

Composite Calculated (g/t) Assay (g/t)

Kiena Deep A Zone – 1 19.51 18.24




S50 Zone 3.11 3.05

13.1.3 Cyanidation Tests

For each composite, the cyanidation tests were conducted in 4 L bottles in triplicate, and the

leaching time was 48 hours. As discussed with Wesdome representatives, the cyanidation

parameters used were the same as the ones used historically at the Kiena Mill:

▪ Grind: 80% passing 75 μm;

▪ Cyanide: 500 g/t NaCN;

▪ Lead Nitrate: 75 g/t;

▪ pH: 11.3 (maintained with lime).

As tests were performed in bottles, no air or oxygen was added during the tests. In addition, as the

lead nitrate addition was usually done at the grinding stage in the mill, the lead nitrate was added

in the bottle one hour prior to the addition of cyanide to simulate the real circuit. Cyanide and pH

were monitored at regular intervals to maintain the proper conditions. At the end of each test, the

solid residue and the solution were collected and sent to the Laboratoire Expert Inc. facility located

in Rouyn-Noranda for gold assaying.




JUNE 2020 13-3

Test Results

The gold recoveries for all cyanidation tests are shown in Table 13-3. For each composite, the

results represent the average of the three cyanidation tests. The 48-hour recoveries for the Kiena

Deep A Zone cyanidation tests ranged from 98.4% to 99.7%.

The 48-hour recoveries for the S50 Zone gave a value of 95.7%, which is in the range of the

historical data from the Kiena Mill.

Table 13-3: Cyanidation test results

Composites NaCN

(kg/t)

CaO

(kg/t)

Au tail

(g/t)

Recovery

(%)

Re-calc Head

(g/t)

Kiena Deep A Zone – 1 0.15 1.22 0.19 98.8 16.40

Kiena Deep A Zone – 2 0.18 1.23 0.23 98.4 13.88

Kiena Deep A Zone – 3 0.11 1.14 0.02 99.7 7.12

Kiena Deep A Zone – 4 0.11 1.30 0.16 99.3 22.21

S50 Zone 0.21 1.56 0.12 95.7 2.83

Lime consumption for all cyanide tests ranged from 1.1 kg/t to 1.6 kg/t. The consumption of NaCN

varied between 0.11 kg/t and 0.21 kg/t. Even though the head grades were higher than the S50

Zone, the reagent consumption was higher for the S50 Zone compared to the Kiena Deep A Zone

for both cyanide and lime. This was probably due to the presence of sulphur (pyrite) in the S50

Zone.

13.2 BBA 2019 Metallurgical Testwork

In July 2019, BBA was selected by Wesdome to design and manage the 2019 testwork program

for the Kiena Mine Complex project. The testwork program started in August 2019 and was

completed in November 2019.

The objective of the testwork program was to obtain preliminary design information and validate

gold recovery for composites selected by Wesdome and BBA.

The high gold recoveries obtained for Kiena Deep A Zone composites (CTRI testwork program,

see Section 13.1) suggested variables that could be investigated in the current testwork program,

e.g.: the impact of pre-oxidation (addition of lead nitrate and pre-aeration), leach time (48 h), and

coarser grind on leach kinetics and gold recovery.




JUNE 2020 13-4

An additional observation from the 2018 CTRI program indicated a significant variance between

the gold assays and the recalculated gold grades for each composite (nugget effect), suggesting

the presence of free gold in the Kiena Deep A Zone. Considering that the gold head grades of the

Kiena Deep A Zone were high (see Table 13-3), it was recommended to conduct gravity testwork

as part of the 2019 testwork program.

The metallurgical testwork was performed at SGS Lakefield and included the following key

deliverables (the number of tests completed is indicated in brackets):

▪ Chemical analysis of composites prepared for testwork;

▪ Grindability testing:

- SMC (3), BWi (3), RWi (3), Ai (3), CWi (1);

▪ Metallurgical testwork:

- Whole ore leach (WOL) (6);

- Gravity recovery and leaching of gravity tails:

o Gravity recoverable gold assessment (3);

o Gravity tailings cyanidation gold recovery (24);

▪ Preliminary selection and estimate of flocculant consumptions by static settling tests (3).

The testwork program flowsheet is shown in Figure 13-1.




JUNE 2020 13-5

Figure 13-1: Testwork program flowsheet

13.2.1 Testwork Samples

The selected samples are shown in Table 13-4. The minimum required sample amount was 25 kg,

and the actual gold grades were within 90-95% accuracy. Sample selection and sample preparation

planning were performed by BBA.




JUNE 2020 13-6

Table 13-4: Testwork samples

Sample ID Zone Sample amount

(kg) Gold grade

(g/t)

A A 36 20.42 (Target)

21.30 (Actual)

A1 A1 33 11.92 (Target)

10.80 (Actual)

A2 A2 32 16.41 (Target)

17.80 (Actual)

The following Figure 13-2 shows the location of the samples used to prepare the master composite.

Figure 13-2: Samples collected to prepare composites A, A1 and A2 for 2019 testwork program

The colour attribute represents “A” Zone (blue), “A1” Zone and “A2” (green)”.




JUNE 2020 13-7

13.2.2 Comminution Tests

The comminution test results are summarized in Table 13-5. The SMC and BWi values show that

the three zone samples are moderately soft. Moreover, the BWi test suggests that the samples are

moderately soft even at finer sizes. CWi value indicates that the Zone A sample is medium hard.

However, this test was completed with a subpar particle size; therefore, the result is only indicative.

According to SGS, the Bond Abrasion Test results are in the range of slightly abrasive to moderately

abrasive.

Table 13-5: Comminution test results

Sample name

Relative density

SMC Parameters Work Indices (kWh/t) Ai (g) A x b ta SCSE CWi RWi BWi

A 2.82 54.0 0.50 8.8 9.5 13.2 12.5 0.181

A1 2.89 55.2 0.49 8.9 N/A 12.2 12.0 0.207

A2 2.84 52.3 0.48 9.0 N/A 13.0 12.7 0.213

13.2.3 Metallurgical Tests

Cyanidation Tests

Whole ore leaching

Whole ore leaching (WOL) tests were conducted on the three master composites: A, A1 and A2.

Common conditions for leaching testwork were:

▪ Grind: 80% passing 75 μm;

▪ Cyanide: 2.4 kg/t NaCN (1.5 g/L maintained);

▪ pH: 11.3 (maintained with lime);

▪ DO: 8-10 mg/L (using air);

▪ Leaching time: 48 hours;

▪ Pulp density: 40%.

Table 13-6 presents the matrix of testwork conditions for WOL. Note that the letter designation at

the end of the test ID represents tests with pre-conditioning (“A”) and without pre-conditioning (“B”).




JUNE 2020 13-8

For each composite, the cyanidation tests were conducted in 2 L bottles. Due to higher gold grade

and the presence of coarse gold in the composites, it was decided to use higher cyanide

concentration (2.4 kg/t NaCN) than the one used historically at the Kiena Mill (0.5 kg/t NaCN).

Additionally, due to the existence of coarse gold in the mineralized material, gold lock up in the

laboratory grinding mills were relatively high compared to typical testwork. In the following tables,

all gold recovery calculations include gold losses in the grinding mills. These losses do not exceed

1% but they still impact the overall gold recovery calculations.

Slurry was aerated with air at the beginning of the testwork including pre-conditioning. In addition,

as the lead nitrate addition was usually done at the grinding stage in the mill, the lead nitrate was

added at the beginning of the pre-conditioning to simulate the real circuit. Cyanide and pH were

monitored at regular intervals to maintain the proper conditions.

Table 13-6: Whole ore leaching test conditions (three master composites)

Test ID

Pre-conditioning Leaching

Pulp density (wt%)

Time (h)

pH Dissolved

oxygen

Addition of lead nitrate at the

beginning of pre-conditioning

(g/t)

pH Dissolved

oxygen NaCN (kg/t)

A-FWOL.A 40 1 11.3 8-10 1,000 11.3 8-10 2.4

A-FWOL.B 40 0 11.3 8-10 0 11.3 8-10 2.4

A1-FWOL.A 40 1 11.3 8-10 1,000 11.3 8-10 2.4

A1-FWOL.B 40 0 11.3 8-10 0 11.3 8-10 2.4

A2-FWOL.A 40 1 11.3 8-10 1,000 11.3 8-10 2.4

A2-FWOL.B 40 0 11.3 8-10 0 11.3 8-10 2.4

WOL test results are shown in Table 13-7. WOL recoveries for Composites A and A2 are 99% with

and without the pre-conditioning step. The WOL recovery for Composite A1 without pre-

conditioning is 95% and with pre-conditioning is 100% (approximately). All samples consumed 20-

25% of the added NaCN. Normalized percent gold extraction is the calculated gold recovery based

on head and residue grades. In WOL tests of samples A1 and A2, the pre-conditioning step slightly

increased the amount of consumed NaCN. On the other hand, the pre-conditioning step decreased

the amount of consumed CaO in the WOL tests of all three samples.




JUNE 2020 13-9

Table 13-7: WOL test results of all three master composites

Test ID

Reagent consumption Au Extraction, including mill clean out, % Head average

(g/t) (direct and calculated)

NaCN consumed

(kg/t)

CaO consumed

(kg/t) 2 h 6 h 24 h 48 h

Normalized % Au extraction

A-FWOL.A 0.49 0.93 48 89 96 99 99 19.6

A-FWOL.B 0.51 0.78 25 54 90 99 99

A1-FWOL.A 0.47 0.84 40 69 100 100 100 11.8

A1-FWOL.B 0.6 0.72 23 45 82 95 95

A2-FWOL.A 0.45 0.89 59 86 99 99 99 16.9

A2-FWOL.B 0.5 0.76 27 61 88 99 99

The kinetic curves for the WOL testing are shown in Figure 13-3. Gold recoveries for samples A

and A2 with pre-conditioning reached 85-90% in approximately 6 hours. More than 80% of the gold

was recovered in less than 25 hours. Figure 13-3 indicates that pre-conditioning increases the

leaching kinetics. Pre-conditioned samples reached recoveries above 95% in 25 hours while non-

pre-conditioned samples reached recoveries between 80% and 90% in 25 hours. Based on this

significant difference in kinetics between the pre-conditioned and non-pre-conditioned samples, it

is advised to implement a pre-conditioning step prior to leaching.

Figure 13-3: WOL test kinetic curve

0

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60

Gold

Extr

action,

%.

Retention Time, h

A-FWOL.A

A-FWOL.B

A1-FWOL.A

A1-FWOL.B

A2-FWOL.A

A2-FWOL.B




JUNE 2020 13-10

Gravity Tests

Knelson/Mozley gravity test results are shown in Table 13-8. The tests resulted in 58.9%-66.7%

gold recovery for the three samples. The implementation of a gravity circuit should be able to

effectively recover coarse and high gold grade particles.

Table 13-8: Knelson/Mozley gravity test results

Sample name

Grind size, P80,

µm

Gravity concentrate Tailings assay Au,

g/t

Head (calc) Au, g/t

Head (direct) Au, g/t Wt %

Assay Au, g/t

Au Rec'y %

A 251 0.1 16,044 60.7 7.8 19.9 21.3

A1 254 0.1 9,374 66.7 2.9 10.7 10.8

A2 246 0.1 11,717 58.9 6.7 16.4 17.8

Gravity Tails Leaching

Gravity tails leaching (GTL) were conducted on the three master composites: A, A1 and A2. The

leach testwork conditions were typically:

▪ Grind: 80% passing 75 μm and 100 μm;

▪ Cyanide: 1.5 kg/t and 0.8 kg/t NaCN (1.0 g/L and 0.5 g/L maintained);

▪ pH: 11.3 (maintained with lime);

▪ DO: 8-10 mg/L (using air);

▪ Leaching time: 48 hours;

▪ Pulp density: 40%.

Table 13-9 presents the matrix of testwork conditions for gravity tails leaching. Note that the letter

designation at the end of the test ID represents tests with pre-conditioning (“A”) and without pre-

conditioning (“B”).

For each composite, the cyanidation tests were conducted in 2 L bottles. As described before, due

to the higher gold grade and presence of coarse gold in the composites, it was decided to use

higher cyanide concentration (1.58 kg/t and 0.8 kg/t NaCN) than the one used historically at the

Kiena Mill (0.5 kg/t NaCN). Additionally, due to the existence of coarse gold in the mineralized

material, gold lock up in the laboratory grinding mills were relatively higher than other projects. In

the following tables, all gold recoveries' calculations include gold losses in the grinding mills. These

losses do not exceed 1% but they still affect the overall gold recovery calculations.




JUNE 2020 13-11

Slurry was aerated with air at the beginning of the testwork including pre-conditioning. In addition,

to simulate a real gold circuit, lead nitrate was added at the beginning of the pre-conditioning step

rather than during the grinding stage. Cyanide and pH were monitored at regular intervals to

maintain the proper conditions. The gravity tails leaching test identification (ID) codes are explained

in Table 13-9. Pre-conditioning includes pre-aeration and lead nitrate addition.

Table 13-9: Gravity tails leaching test ID codes

Test ID Particle size

(µm) Pre-conditioning

Lead nitrate addition

NaCN addition

FGT.101 75 (Fine) Yes High High

CGT.201 100 (Coarse) Yes High High

FGT.102 75 (Fine) No None High

CGT.202 100 (Coarse) No None High

FGT.103 75 (Fine) Yes Low High

CGT.203 100 (Coarse) Yes Low High

FGT.104 75 (Fine) Yes High Low

CGT.204 100 (Coarse) Yes High Low

Gravity tails leaching test conditions are shown in Table 13-10. As presented previously, the final

letter designation in the test ID represents tests with pre-conditioning (“A”) and without pre-

conditioning (“B”).

Table 13-10: Gravity tails leaching test conditions for each master composite

Test ID

Pre-conditioning Leaching

Pulp density (wt%)

Time (h)

pH Dissolved

oxygen

Addition of lead nitrate at the beginning of

pre-conditioning (g/t)

Particle size

(µm)

pH Dissolved

oxygen NaCN (ppm)

FGT.101.A 40 1 11.3 8-10 150 75 11.3 8-10 1,000

CGT.201.A 40 1 11.3 8-10 150 100 11.3 8-10 1,000

FGT.102.B 40 0 11.3 8-10 0 75 11.3 8-10 1,000

CGT.202.B 40 0 11.3 8-10 0 100 11.3 8-10 1,000

FGT.103.A 40 1 11.3 8-10 75 75 11.3 8-10 1,000

CGT.203.A 40 1 11.3 8-10 75 100 11.3 8-10 1,000

FGT.104.A 40 1 11.3 8-11 150 75 11.3 8-10 500

CGT.204.A 40 1 11.3 8-12 150 100 11.3 8-10 500

Gravity tails leaching test results are shown in Table 13-11.




JUNE 2020 13-12

Table 13-11: Gravity tails leaching test results

Zone Test ID

Reagent Consumption Au Extraction, including mill clean out, %

NaCN consumed

(kg/t)

CaO consumed

(kg/t) 2 h 6 h 24 h 48 h

Normalized % Au extraction

Residue, Au

g/t

Calculated head, Au

g/t

Head average (g/t) (direct and

calculated) (1)

A

FGT.101.A 0.20 0.89 57 90 ~99 98.2 98.3 0.13 8.19

7.82

CGT.201.A 0.22 0.84 55 87 98 96.8 97.1 0.23 7.78

FGT.102.B 0.21 0.85 38 77 ~99 97.9 98.1 0.15 7.89

CGT.202.B 0.18 0.71 35 75 98 96.5 96.7 0.26 8.01

FGT.103.A 0.31 0.88 63 96 ~99 97.5 97.8 0.17 7.55

CGT.203.A 0.33 0.91 56 91 ~99 96.6 96.9 0.24 7.53

FGT.104.A 0.23 0.93 53 89 ~99 97.8 98.1 0.15 7.49

CGT.204.A 0.18 0.95 44 79 98 96.0 96.2 0.30 7.89

A1

FGT.101.A 0.20 0.84 55 85 98 96.5 97.1 0.11 3.92

3.80

CGT.201.A 0.35 0.82 52 83 98 95.3 95.8 0.16 3.97

FGT.102.B 0.37 0.81 45 78 ~99 96.2 96.8 0.12 3.86

CGT.202.B 0.32 0.79 41 71 98 94.9 95.5 0.17 3.92

FGT.103.A 0.32 0.89 57 87 ~99 96.5 97.1 0.11 3.98

CGT.203.A 0.31 0.80 51 78 91 95.3 95.8 0.16 3.99

FGT.104.A 0.19 0.92 50 82 98 96.6 97.4 0.10 3.79

CGT.204.A 0.28 0.86 41 75 98 95.4 96.1 0.15 3.91

A2

FGT.101.A 0.44 0.92 76 94 ~99 97.6 97.6 0.16 6.73

6.63

CGT.201.A 0.33 0.88 72 88 95 95.7 95.5 0.30 7.02

FGT.102.B 0.45 0.87 39 67 97 97.8 97.7 0.15 6.85

CGT.202.B 0.39 0.83 37 67 95 95.5 95.2 0.32 7.06

FGT.103.A 0.35 0.79 84 96 99 97.4 97.4 0.17 6.62

CGT.203.A 0.36 0.88 77 92 ~99 96.7 96.5 0.23 6.93

FGT.104.A 0.24 1.03 35 66 96 95.1 96.2 0.25 5.95

CGT.204.A 0.24 0.96 62 88 97 96.3 96.2 0.25 6.82

(1) Average of direct gravity tailings heads and cyanidation calculated heads.




JUNE 2020 13-13

A Zone

Gravity tails leaching test kinetic curves of Kiena Deep A Zones samples are shown in Figure 13-4

and Figure 13-5. These figures show that the effect of particle size on the overall recovery is

insignificant. Pre-conditioning significantly increases the leaching kinetics of the A Zone sample.

Moreover, increasing the lead nitrate content from 75 g/t to 150 g/t decreased the leaching kinetics

for coarse and fine size samples. On the other hand, after 24 hours of leaching, gold recoveries

reached their maximum regardless of the sample particle size and whether or not pre-conditioning

was applied.

Figure 13-4: Zone A, effect of pre-leaching and lead nitrate at 75 microns




JUNE 2020 13-14

Figure 13-5: Zone A, effect of pre-leaching and lead nitrate at 100 microns

The effects of increasing the NaCN content and decreasing the particle size on the gold recovery

and leaching kinetics can be observed in Figure 13-6. Based on this figure, it can be said that

increasing the NaCN to 1 g/L increases the leaching kinetics and overall gold recovery. Coarser

(100 microns) samples have shown slower kinetics and lower gold recovery in comparison to their

finer counterparts. However, the gold recovery differences were relatively small and did not exceed

1-2%. Although, less cyanide was added and coarser samples had relatively slower kinetics, the

kinetics of these samples are still considered to be good.




JUNE 2020 13-15

Figure 13-6: Zone A, the effect of different CN addition on gold recovery

A1 Zone

Gravity tails leaching test kinetic curves of A1 Zone sample are shown in Figure 13-7 and

Figure 13-8. These figures show that reducing the grind size to 75 microns slightly increases the

overall gold recovery (1-2%). Moreover, pre-aeration and addition of lead nitrate significantly

increases the leaching kinetics of the A1 Zone sample during the first hours of leaching. On the

other hand, increasing the lead nitrate content from 75 g/t to 150 g/t slightly decreased the leaching

kinetics of fine size samples while it increased the leaching kinetics of coarse size samples. In

Figure 13-8, it is shown that the sample with 75 g/t lead nitrate with added pre-conditioning reached

maximum gold recovery in 48 hours. However, it is believed that this was caused by experimental

error and, this sample also reached its maximum gold recovery in 24 hours. After 24 hours of

leaching, gold recoveries of all samples reached their maximum regardless of the sample particle

size and whether or not pre-conditioning was applied.




JUNE 2020 13-16

Figure 13-7: A1 Zone, effect of pre-leaching and lead nitrate at 75 microns

Figure 13-8: A1 Zone, effect of pre-leaching and lead nitrate at 100 microns




JUNE 2020 13-17



increasing the NaCN to 1 g/L increases the leaching kinetics and overall gold recovery. Coarser

(100 microns) samples have shown slower kinetics and lower gold recovery in comparison to their

finer counterparts. However, the gold recovery differences were small and did not exceed 1-2%.

Although, less cyanide was added and coarser samples had relatively slower kinetics, the kinetics

of these samples are still considered to be good.

Figure 13-9: A1 Zone, the effect of different CN addition amounts

A2 Zone

Gravity tails leaching test kinetic curves of the A2 Zone sample are shown in Figure 13-10 and

Figure 13-11. These figures show that the effect of particle size on the overall recovery is

infinitesimal. Moreover, pre-aeration and addition of lead nitrate significantly increased the leaching

kinetics of A2 Zone samples. On the other hand, increasing the lead nitrate content from 75 g/t to

150 g/t slightly decreased the leaching kinetics for coarse and fine size samples. In these figures,

the effect of pre-conditioning on leaching kinetics is clearly observed. For the A2 sample, the best

pre-conditioning parameter was pre-aerating with 75 g/t lead nitrate addition. Based on

observations, it can be said that all A2 samples reached their maximum gold recoveries in 24 hours.

The misrepresentation was simply caused by experimental errors.




JUNE 2020 13-18

Figure 13-10: Zone A2, effect of pre-leaching and lead nitrate at 75 microns

Figure 13-11: Zone A2, effect of pre-leaching and lead nitrate at 100 microns




JUNE 2020 13-19



increasing the NaCN to 1 g/L increases the leaching kinetics and overall gold recovery. However,

the final recovery differences were small and did not exceed 1-2%.

Figure 13-12: Zone A2, the effect of different CN addition amounts

Summary (Zone A, A1 and A2)

In Table 13-12, the overall gravity and gravity tails leaching gold recoveries of samples Zone A, A1

and A2 are shown. Note that in the table, the letter designation at the end of the test ID represents

tests with pre-conditioning (“A”) and without pre-conditioning (“B”). The results indicate that gravity

concentration and leaching can yield extremely high gold recoveries. However, gold recoveries

higher than 95% can already be achieved with WOL as well. The addition of a gravity circuit will not

significantly increase the overall gold recovery, but it will increase the process kinetics and reduce

the feed variability in the leaching circuit




JUNE 2020 13-20

Table 13-12: Overall recoveries of gravity and gravity tails leaching

Zone Test ID

Pre-conditioning Leaching Overall

normalized % Au

extraction

Pulp density

(wt%)

Time (h)

pH Dissolved

oxygen

Lead nitrate

addition

(g/t)

Particle size

(µm)

pH Dissolved

oxygen NaCN (ppm)

A

FGT.101.A 40 1 11.3 8-10 150 75 11.3 8-10 1,000 99.3

CGT.201.A 40 1 11.3 8-10 150 100 11.3 8-10 1,000 98.8

FGT.102.B 40 0 11.3 8-10 0 75 11.3 8-10 1,000 99.2

CGT.202.B 40 0 11.3 8-10 0 100 11.3 8-10 1,000 98.7

FGT.103.A 40 1 11.3 8-10 75 75 11.3 8-10 1,000 99.1

CGT.203.A 40 1 11.3 8-10 75 100 11.3 8-10 1,000 98.8

FGT.104.A 40 1 11.3 8-11 150 75 11.3 8-10 500 99.2

CGT.204.A 40 1 11.3 8-12 150 100 11.3 8-10 500 98.5

A1

FGT.101.A 40 1 11.3 8-10 150 75 11.3 8-10 1,000 99.0

CGT.201.A 40 1 11.3 8-10 150 100 11.3 8-10 1,000 98.6

FGT.102.B 40 0 11.3 8-10 0 75 11.3 8-10 1,000 98.9

CGT.202.B 40 0 11.3 8-10 0 100 11.3 8-10 1,000 98.5

FGT.103.A 40 1 11.3 8-10 75 75 11.3 8-10 1,000 99.0

CGT.203.A 40 1 11.3 8-10 75 100 11.3 8-10 1,000 98.6

FGT.104.A 40 1 11.3 8-11 150 75 11.3 8-10 500 99.1

CGT.204.A 40 1 11.3 8-12 150 100 11.3 8-10 500 98.7

A2

FGT.101.A 40 1 11.3 8-10 150 75 11.3 8-10 1,000 99.1

CGT.201.A 40 1 11.3 8-10 150 100 11.3 8-10 1,000 98.2

FGT.102.B 40 0 11.3 8-10 0 75 11.3 8-10 1,000 99.1

CGT.202.B 40 0 11.3 8-10 0 100 11.3 8-10 1,000 98.1

FGT.103.A 40 1 11.3 8-10 75 75 11.3 8-10 1,000 99.0

CGT.203.A 40 1 11.3 8-10 75 100 11.3 8-10 1,000 98.6

FGT.104.A 40 1 11.3 8-11 150 75 11.3 8-10 500 98.5

CGT.204.A 40 1 11.3 8-12 150 100 11.3 8-10 500 98.5

Note: The letter designation at the end of the test ID represents: tests with pre-conditioning (“A”) and without pre-conditioning (“B”).

13.2.4 Reagent Consumptions

Reagent consumptions for GTL and WOL tests are shown in Table 13-13. WOL samples consumed

more NaCN than GTL samples. In GTL, sample A consumed 54%, sample A1 46% and sample A2

26% less cyanide than the amounts consumed in WOL. However, the amount of NaCN consumed

in all leaching tests were relatively low. Lastly, lime consumption was average for this type of

mineralized material.




JUNE 2020 13-21

Table 13-13: Reagent consumptions per test type

Zone

Gravity tails leaching Whole ore leaching

Test ID NaCN consumed

(kg/t) CaO consumed

(kg/t) Pb(NO3)2 (Lead nitrate),

g/t Test ID

NaCN consumed (kg/t)

CaO consumed (kg/t)

A

FGT.101 0.20 0.89 150

A-FWOL.A1 0.49 0.93 CGT.201 0.22 0.84 150

FGT.102 0.21 0.85 N/A

CGT.202 0.18 0.71 N/A

FGT.103 0.31 0.88 75

A-FWOL.A2 0.51 0.78 CGT.203 0.33 0.91 75

FGT.104 0.23 0.93 150

CGT.204 0.18 0.95 150

A1

FGT.101 0.20 0.84 150

A1-FWOL.A1 0.47 0.84 CGT.201 0.35 0.82 150

FGT.102 0.37 0.81 N/A

CGT.202 0.32 0.79 N/A

FGT.103 0.32 0.89 75

A1-FWOL.A2 0.60 0.72 CGT.203 0.31 0.80 75

FGT.104 0.19 0.92 150

CGT.204 0.28 0.86 150

A2

FGT.101 0.44 0.92 150

A2-FWOL.A1 0.45 0.89 CGT.201 0.33 0.88 150

FGT.102 0.45 0.87 N/A

CGT.202 0.39 0.83 N/A

FGT.103 0.35 0.79 75

A2-FWOL.A2 0.50 0.76 CGT.203 0.36 0.88 75

FGT.104 0.24 1.03 150

CGT.204 0.24 0.96 150




JUNE 2020 13-22

13.2.5 Solid and Liquid Separation Tests

In order to select the most efficient flocculant for thickening, flocculant scoping tests were

performed. Based on the flocculant selection test results, SNF Flomin 913 VHM was found to be

the most effective flocculant among the flocculants tested. The flocculant scoping test results are

shown in Table 13-14. Composite A tailings required a dosage of 9 grams per metric tonne of dry

mineralized material and settled very fast, while Composite A1 and A2 tails settled fast at the higher

dosage of 12 g/t. These dosages were found to be relatively low when they were benchmarked with

similar mineralized material types. High clarity of supernatants has also proven the high efficiency

of SNF Flomin 913 VHM.

Table 13-14: Flocculant scoping test results summary

Corresponding composite ID

Pulp pH Pulp % solids

(%w/w) Flocculant Dosage g/t

Supernatant clarity

Settling rate

A

10.5

10.0 913 VHM 9 Clear Very Fast

A1 10.0 913 VHM 12 Clear Fast

A2 10.0 913 VHM 12 Clear Fast

13.3 Summary

Two metallurgical testwork programs were conducted with mineralized material from the Kiena

deposit in 2018 (Kiena Deep A from zones A and A1, and S50 zones) and 2019 (Kiena Deep A

Zones A, A1 and A2). Table 13-15 presents a summary of the best overall gold recoveries at

comparable testwork conditions.




JUNE 2020 13-23

Table 13-15: Overall Au recovery for each testwork program

Composite ID Program Protocol P80,

microns

Au calc,

g/t

Au rec,

%

NaCN cons.,

kg/t

CaO cons.,

kg/t





S50 Zone CTRI WOL 75 3.11 95.7 0.25 1.56

A SGS WOL(1) 75 19.6 99.0 0.49 0.93

A1 SGS WOL(1) 75 11.8 100 0.47 0.84

A2 SGS WOL(1) 75 19.6 99.0 0.45 0.89

A1 SGS GTL(2) 73 19.9 99.3 0.2 0.89

A2 SGS GTL(2) 75 10.7 99.0 0.2 0.84

A2 SGS GTL(2) 70 16.4 99.1 0.44 0.92

(1) 1 kg/t lead nitrate

(2) 150 g/t lead nitrate (in leaching).

Overall Au recoveries were in a similar range, but the best results were found in GTL followed by

SGS’s WOL. The SGS’s GTL NaCN consumption was comparable to the CTRI’s WOL testwork.

The highest NaCN consumption was found in SGS’s WOL tests. The highest lime consumption

was found in the testwork carried out at CTRI. The lowest Au recovery was found in CTRI’s WOL

results.

13.4 Conclusions

The PEA metallurgical testwork was conducted for the Kiena Mine Complex at two laboratories:

the Centre Technologique des Résidus Industriels (CTRI) and SGS Lakefield. The PEA test

program was performed using composite samples representing different zones of the Kiena Mine

Complex. The comminution tests used Kiena composites from zones A, A1 and A2. Metallurgical

tests used composites from Kiena A zones 1 to 4 and S50 Zone (CTRI program); and Kiena A, A1

and A2 zones (SGS program).

▪ The comminution test results have shown that all samples are moderately soft. Furthermore,

the Bond Abrasion Test indicate that the samples are in the range of slightly abrasive to

moderately abrasive.

▪ The calculated gold feed grades of Kiena A samples from zones 1 to 4 and S50 Zone were

found to be 19.51, 12.22, 6.45, 22.38 and 3.11 g/t respectively. On the other hand, the

calculated gold feed grades of samples Kiena A, A1 and A2 were found to be 19.6 g/t,

11.8 g/t and 16.9 g/t (WOL) and 19.9 g/t, 10.7 g/t and 16.4 g/t (GTL) respectively.




JUNE 2020 13-24

▪ The WOL tests resulted in high gold recoveries independent of the laboratory. Gold

recoveries varied between 95.7% and 99.7%. The effects of pre-aeration and lead nitrate

addition (pre-conditioning) on overall gold recoveries and leach kinetics were analyzed. Pre-

conditioning was found to have minimal effect on the overall gold recoveries, but it improved

the leach kinetics (tests with pre-conditioning reached a gold recovery plateau after 24 hours

of leaching). The majority of WOL test samples reached their maximum gold recoveries in

approximately 48 hours.

- Pre-conditioning has significantly positive effect on leach kinetics; therefore, it is

suggested to validate this observation in the next phase of testwork.

- Pre-conditioning (lead nitrate) does not reduce the cyanide consumption. A slight

increase in lime consumption was observed in the tests with pre-conditioning.

▪ In Knelson/Mozley gravity tests, high gold recoveries of 61%, 67% and 59% were achieved

for samples A, A1 and A2 respectively. Installation of a gravity circuit to the existing process

has the potential to recover gold faster and reduce the amount of coarse gold particles

reporting to the leach circuit for improved process economics.

▪ The GTL tests resulted in extremely high gold recoveries. Like WOL test results, gold

recoveries varied between 95.7% and 99.7%. When gravity recoveries were added to the

GTL recoveries, the best overall gravity and gravity tails leaching recovery ranged from 99%

to 99.3%. Additionally, the impacts regarding the of grind size, NaCN content, pre-aeration

and lead nitrate addition (pre-conditioning) on overall gold recoveries and leach kinetics were

analyzed. Slightly increased gold recovery was achieved by finer grinding (75 microns) and

higher amount of NaCN addition. Pre-conditioning was found to have minimal effect on the

overall gold recoveries, but it improved the leach kinetics. The majority of the GTL test

samples reached their maximum gold recoveries in approximately 24 hours.

- Pre-conditioning has a positive effect on leach kinetics; therefore, it is recommended to

implement pre-aeration and lead nitrate addition to the process. However, the kinetic

characteristics of the mineralized material need to be confirmed in future testworks.

- Although, finer grind size increased the gold recovery, the effect of grind size should be

investigated with other grind sizes. For achieving the highest gold recoveries, it is

suggested to keep the final grind size at around 75 microns.

- The effect of increased NaCN addition on overall gold recovery was minimal. It is

suggested that the NaCN content be maintained at around 0.5 g/L.

▪ Slightly higher overall gold recoveries (~ 0.3% to 1.0%) were achieved by GTL in comparison

to WOL. An additional improvement was observed in the GTL tests: Due to the gold

extraction achieved by the gravity system, the gold feed to leach was lower than WOL and

as a result, the NaCN dosage and consumption were lower for GTL than WOL. A validation

testwork program should be conducted to confirm the magnitude of the NaCN savings. If this

observation is confirmed by the testwork, the installation of a gravity circuit may be

considered and validated with a trade-off study.




JUNE 2020 13-25

▪ NaCN consumptions in WOL and GTL tests were less than 0.6 kg/t, which is considered low.

However, in GTL, sample A consumed 54%, sample A1 46% and sample A2 26% less

cyanide than the amounts consumed in WOL. The amount of consumed CaO was found to

be slightly lower than 1 kg/t in each test for all samples. CaO consumptions of all samples

were found to be average.

▪ SNF Flomin 913 VHM was found to be the most effective and efficient flocculant among the

flocculants tested. The dosage range of 9-12 g/t was found to be relatively low.

13.5 Recommendations

High gold recoveries were achieved in the metallurgical testwork program regardless of the

flowsheet employed (WOL or GTL). However, additional tests need to be performed to confirm the

high gold recovery from mineralized material across the Kiena Mine Complex. The following actions

are recommended:

▪ Future testwork should be conducted using samples with gold head grades based on the

latest mine plan gold head grades:

- The effect of pre-conditioning without lead nitrate addition should be investigated in

future testwork;

- Additional testwork should be performed to optimize relatively shorter leach time;

- Cyanide destruction testwork and filtration testwork should be conducted to validate

process assumptions and future scenario.

▪ Variability samples should be prepared from more representative zones of the Kiena Mine

Complex and they should undergo WOL and GTL for further characterization. Additional

grindability testwork is recommended to complete the mineralized material hardness

characterization and model the circuit operation.

▪ GTL and WOL tests have given similar recovery values but NaCN consumptions in these two

different test types were significantly different. Although, the addition of a gravity circuit will

increase the CAPEX costs and equipment maintenance costs, it can potentially reduce

NaCN consumption by up to 50% in grams per tonne of mineralized material milled. An

additional trade-off study should be performed to evaluate the impact of installing a gravity

circuit on the CAPEX and OPEX.

▪ The preliminary analysis shows that the installation of gravity and gravity leach circuit does

slightly increase the final gold recovery. NaCN consumption averages from two different

leaching tests were calculated and compared. In GTL, approximately 42% less NaCN was

consumed compared to WOL. This reduction leads to annual savings of more than $150,000

on the cost of reagents. Therefore, the total annual saving can be between $150,000 and

$200,000 when gravity circuit is installed. However, this total should be considered as

indicative and a more detailed study should be performed.




JUNE 2020 13-26

13.6 Future Steps

The following future testwork is recommended for the Kiena Deep deposit:

▪ A comminution testwork program to study the mineralized material hardness variability (5 to

10 samples per ore zone);

▪ A metallurgical testwork program to study the Au recovery variability with Au head grade (5

to 10 samples per ore zone);

▪ Gravity recoverable gold (GRG) testwork to characterize the nature of the gravity gold in the

Kiena Deep A Zone. A cyanide leaching optimization program could be implemented

following a GRG testwork program;

▪ Optional: an optimization testwork program to study the optimization of leaching variables for

the option selected in the current testwork program (WOL or GTL):

- Stirred reactor tests could be conducted to validate or optimize process variables such

as cyanide addition, oxygen vs. air, lead nitrate addition, etc.

▪ A preliminary cyanide destruction testwork program based on the future tailings handling

system;

▪ A dynamic settling testwork program to optimize reagent addition;

▪ Filtration testwork.

It is recommended to build a preliminary grinding circuit model using information available to

estimate mill performance and final particle size. This information will be important for designing

the filtration and tailings systems.




JUNE 2020 14-1

MINERAL RESOURCE ESTIMATE

14.1 Introduction

In 2019, BBA was retained by Wesdome Gold Mines to review and audit the updated Mineral

Resource Estimate (MRE) for the Kiena Mine Complex project (the “Project”), which incorporates

historical drilling data and recent drilling programs. The MRE presented herein was initially

presented in a Technical Report entitled "Mineral Resource Estimate for the Kiena Mine Complex

Project" dated November 7, 2019 (Richard and Torrealba, 2019). It includes drill data as of

August 6, 2019.

Since the close out date of this MRE, five underground drill rigs have continued to operate in

order to verify the up-plunge extension (VC6 and VC1 zones), test the down dip extension and

perform infill drilling in the Kiena Deep A Zones. Underground drill rigs were active at level 67

(one drill rig) and in the new exploration ramp (four drill rigs). In December 2019, a sixth drill rig

was set-up at the new level 79 development in order to explore the VC6 and VC1 zones

extensions. A seventh drill rig was also mobilized at level 100 in early January 2020.

As of January 15, 2020, Wesdome had completed an additional 112 DDH for 32,413 m that are

not included in the herein MRE (Appendix B). The QP is of the opinion that while the addition of



It should be noted that the MRE presented herein has an effective date prior to the new CIM Best

Practice Guidelines (published November 29, 2019) and, therefore, was produced in accordance

to the previous Guidelines. That being said, the QP is of the opinion that, had this MRE been

prepared under the new Guidelines, it would not have had a material impact on the MRE

presented in this Report.

14.2 Methodology

The herein MRE combines two different approaches:

▪ A block model mineral resource estimate for the zones in the former Kiena Mine area (the

‘block model MRE”) which was prepared by Karine Brousseau, P. Eng., Senior Engineer of

Wesdome and has been reviewed and audited by Pierre-Luc Richard, P. Geo., Qualified

Person of BBA.;

▪ A polygonal mineral resource estimate for the zones outside of the Kiena Mine area (the

“polygonal MRE”) which was prepared by Turcotte et al. (2015) and reviewed and modified

by Pierre-Luc Richard, P. Geo., Qualified Person of BBA.




JUNE 2020 14-2

The methodology for the block model MRE was the following:

▪ Geological wireframes were constructed in Leapfrog Geo 4.5™. Geovia® GEMS 6.8.2.2 was

used for the compositing, 3D block modelling, interpolation and classification. Statistical

studies were conducted using Excel and Snowden Supervisor v. 8.11.

The methodology for the estimation of the mineral resources involved the following steps:

▪ Database verification and validation;

▪ Review of the 3D modelling;

▪ Drillhole intercept;

▪ Basic statistics and composite generation for each unit;

▪ Capping;

▪ Geostatistical analysis including variography;

▪ Block modelling and grade interpolation;

▪ Block model validation;

▪ Resource classification;

▪ Cut-off grade calculation;

▪ Preparation of the mineral resource statement.

The Polygonal MRE was prepared by Turcotte and Pelletier in 2009, validated in 2015 by

Turcotte et al. using AutoCAD, Microsoft Excel and Microsoft Access using the polygonal method

on longitudinal sections. The Polygonal MRE was herein validated using Microsoft Excel,

AutoCAD, and historical longitudinal sections. For the herein Polygonal MRE, the following steps

were carried-out:

▪ Confirmation of the lack of new material information;

▪ Review and validation of the resource with Excel;

▪ Review of the classification in long section;

▪ Review of the underground cut-off grade.




JUNE 2020 14-3

Figure 14-1: 2019 MRE block model and polygonal resources location




JUNE 2020 14-4

14.3 Resource Database

The block model resource database for the Kiena Mine area, as of August 6, 2019, consisted of

349 surface and 6,267 underground drillholes with a cumulative length of 893,318.52 m

(Figure 14-2). The average length of a drillhole is 135 m. Of these 6,616 drillholes, a subset of

714 holes cut across the mineralized zones. A total of 36,050 m in 140 drillholes was added to the

block model resource database since the last mineral resource estimate announced on

December 12, 2018.

The polygonal resource database consisted of 216 drillholes cutting across the mineralized

zones. No new information was added to the polygonal resource database since 2015.

The resource estimation for the Project relies on historical and recent drilling programs. BBA

included the historical drillhole information into the resource estimation for the following reasons:

1) historical information was validated as part of the mandate and no discrepancies were found;

and 2) recent drillholes were drilled in the vicinity of historical drillholes and the results show

comparable geology and mineralization outlines.

The resource database was validated before proceeding to the resource estimation. The

validation steps are detailed in Chapter 12 of this Report. Minor variations have been noted

during the validation process but have no material impact on the 2019 MRE.

The QP is of the opinion that the database is appropriate for the purposes of the mineral resource

estimation and that the sample density allows a reliable estimate to be made of the size, tonnage

and grade of the mineralization in accordance with the level of confidence established by the

mineral resource categories as set forth in the CIM Standards.




JUNE 2020 14-5

Figure 14-2: 3D view looking west of the 3D model and of the drillholes included in this resource estimate




JUNE 2020 14-6

14.4 Geological Interpretation and Modelling

A total of 48 zones were interpreted for the purpose of this MRE (Table 14-1).

Table 14-1: Mineralized zones of the 2019 MRE

Polygonal MRE Block model MRE

Domain Zone Domain Zone

Martin Martin

Zone S50

S50_100

Dubuisson

Dubuisson S50_101

Dubuisson North 1 S50_102

Dubuisson North 2 S50_103

NorthWest NorthWest S50_104

Presqu'ile Zone 1

Zone VC1

VC1_111

Zone 2 VC1_112

Wesdome

A VC1_113

A1 VC1_114

AF Zone VC6 VC6_123

AH

Zone South

ZS_130

AH1 ZS_131

AH2 ZS_132

AH3 ZS_133

B ZS_135

C Zone B ZB_140

D

Kiena Deep Zone A

ZA

E ZA1

E0 ZA2

E1 H1ZA

E3 E4 F

F1 F2 F4 F6 L




JUNE 2020 14-7

14.4.1 Geological Model

14.4.1.1 Polygonal MRE

For the polygonal MRE, 28 mineralized zones, grouped in 5 deposits, were interpreted outside

of the Kiena Mine area (Table 14-1). The Wesdome Deposit was interpreted in 2009 (Turcotte

and Pelletier, 2009) and validated in 2015 (Turcotte et al., 2015). The other deposits were

interpreted by the Wesdome exploration team and published for year-end 2014 and validated

in Turcotte et al. (2015).

The interpretations of these zones were made on cross and longitudinal sections and

validated in GEMS with a minimum true thickness of 1.5 m for the Wesdome deposit and

2.5 m for the seven other zones. As no new information has become available through recent

exploration or drilling programs, the QP is of opinion that the interpretation remains valid.

14.4.1.2 Block Model MRE

For the Block Model MRE, geological wireframes were constructed in Leapfrog Geo™ by

Charlotte Athurion of BBA and Karine Brousseau of Wesdome with the help of Bruno Turcotte

of Wesdome. The model comprises 20 mineralized zones which have a minimum thickness of

3 m (Figure 14-3 and Figure 14-4). They were modelled using geological knowledge of the

deposit, grade continuity and geological information provided in the DDH logs (i.e.: lithology,

alteration and structure) and in the historical underground mappings. Geological interpretation

of the basalt and schists units was also carried-out for the purpose of better constraining the

mineralized zones. These zones are further described in Section 7.3 of this report.

As the Kiena Deep A zones are folded, they were subdivided in three subdomains

representing the two limbs and the hinge for each mineralized zone.

The QP reviewed the geological model in 3D view, plan view and cross-section and is of the

opinion that the level of detail to which the geology model was constructed represents

adequately the Complexity of the deposit. In the QP’s opinion, the geological model is

appropriate for the size, grade distribution and geometry of the mineralized zones and is

suitable for the resource estimation of the Kiena Mine area.




JUNE 2020 14-8

Figure 14-3: 3D view showing the mineralized zones and undergrounds voids in the Kiena Mine area looking west




JUNE 2020 14-9

Figure 14-4: Plan view showing the mineralized zones and undergrounds drifts in the Kiena Mine area looking down




JUNE 2020 14-10

14.4.2 Voids Model

The void model for this MRE was provided by Wesdome. Figure 14-3 presents a 3D view of the

underground voids used for the 2019 MRE. Validations were made to make sure that any new

developments were considered for this new MRE.

14.4.3 Overburden and Topography

The topographic surface and the overburden-rock interface were created by Wesdome in

Leapfrog Geo™ and are based on the drillholes collar coordinates and elevation and the

lithological description.

14.5 Data Analysis - Block Model MRE

14.5.1 Raw Assay Statistics

All raw assay data that intersected the mineralized zones were assigned individual rock codes.

These coded intercepts were used to produce basic statistics on sample lengths and grades. A

total of 7,749 assays are included in the modelled wireframe.

Basic statistics on the raw assays are presented in Table 14-2.

Table 14-2: Basic statistics on raw assays for each domain

Raw Assays

Sector Count

sample

Uncut mean (g/t

Au)

Std. dev.

COV Max

(g/t Au) Min

(g/t Au) Uncut median

(g/t Au)

S50_100 478 1.99 2.84 1.43 42.17 0.005 1.27

S50_101 516 2.19 3.65 1.67 52.64 0.005 1.49

S50_102 830 3.64 11.11 3.06 256.67 0.005 1.66

S50_103 311 1.95 7.83 4.03 134.46 0.005 0.84

S50_104 237 2.25 7.25 3.23 72.37 0.005 0.32

VC1_111 209 4.08 12.54 3.07 147.57 0.010 1.23

VC1_112 73 8.32 49.65 5.97 424.01 0.005 0.14

VC1_113 96 34.97 262.85 7.52 2,578.56 0.005 0.07

VC1_114 358 2.83 6.74 2.38 57.10 0.005 0.59

VC6_123 206 3.28 6.72 2.05 54.93 0.002 1.14

ZS_130 547 1.57 4.63 2.94 96.62 0.005 0.69

ZS_131 460 2.36 4.57 1.94 43.50 0.005 0.69




JUNE 2020 14-11

Raw Assays

Sector Count

sample

Uncut mean (g/t

Au)

Std. dev.

COV Max

(g/t Au) Min

(g/t Au) Uncut median

(g/t Au)

ZS_132 554 1.60 2.65 1.65 28.80 0.005 0.60

ZS_133 161 1.99 8.57 4.30 105.60 0.005 0.61

ZS_135 153 1.39 4.01 2.88 36.50 0.005 0.04

ZB_140 175 2.40 7.23 3.01 58.96 0.005 0.09

ZA 1,270 25.02 118.76 4.75 2,769.58 0.002 0.28

ZA1 627 14.02 62.99 4.49 704.00 0.002 0.27

ZA2 412 26.52 94.77 3.57 886.84 0.002 0.63

H1ZA 76 5.05 23.44 4.64 191.00 0.005 0.04

14.5.2 Compositing

Compositing of drillhole samples was conducted in order to homogenize the database for the

statistical analysis and remove any bias associated to the sample length that may exist in the

original database. The composite length was determined using original sample length statistics

and the thickness of the mineralized zones.

Inside the mineralized zones, the average sample length is 1.01 m and the median is 1.00 m.

Figure 14-5 shows the sample length distribution within the mineralized zones.

As a result, 8,402 composites were generated with a length of 1 m but ranging from 0.60 m to

1.45 m when necessary after redistributing the tails.

Grades of 0.00g/t Au were assigned to all missing intervals during the compositing process.




JUNE 2020 14-12

Figure 14-5: Sample length distribution within the mineralized zones

14.5.3 Outlier Handling

An outlier is an observation that appears to be inconsistent with the majority of the data. It is

common practice to statistically examine the higher grades within a population and to trim the

outlier to a lower grade value based on the results of a statistical study. The capping is performed

on high-grade values considered to be outliers. High-grade capping was done on the composited

assay data and established on a per deposit or zone type basis.

In addition, a high grade limit or second capping value was used for the second and third pass

grade interpolation to restrict high grade impact at greater distance from the drillhole intersect for

some zones (Table 14-8). It should be noted that this restriction approach is not a capping

method per say, but rather a way to exclude higher grades to be used during the interpolation

process when estimating blocks outside this restricted search ellipsoid.




JUNE 2020 14-13

The capping values were defined by searching for abnormal breaks or change of slope on the

grade distribution probability plot while making sure that the coefficient of variation of the capped

data was ideally lower than, or around 2.00 and no more than 10% of the total contained metal

was enclosed within the first 1% of the highest-grade samples. The use of various statistical

methods allows for a selection of the capping threshold in a more objective and justified manner.

In any cases where the coefficient of variation was higher than 2.00, a restrictive search ellipsoid

was used at a value allowing to reach that coefficient of variation of 2.00.

Basic statistics for composited assays and capped composites are summarized in Table 14-3.

Figure 14-6 to Figure 14-15 show graphs supporting the capping threshold decisions.

Figure 14-6: Graphs supporting capping threshold decisions on composites for the S50 deposit

(blockcodes 100 to 104)




JUNE 2020 14-14

Figure 14-7: Graphs supporting capping threshold decisions on composites for the VC1 deposit

(blockcodes 111 to 114)




JUNE 2020 14-15

Figure 14-8: Graphs supporting capping threshold decisions on composites for the VC6 Zone

(blockcode 123)




JUNE 2020 14-16

Figure 14-9: Graphs supporting capping threshold decisions on composites for the South Zone

(blockcodes 130 to 133 and 135)




JUNE 2020 14-17

Figure 14-10: Graphs supporting capping threshold decisions on composites for the B Zone

(blockcode 140)




JUNE 2020 14-18

Figure 14-11: Graphs supporting capping threshold decisions on composites for the Kiena Deep A Zone A subzones

(blockcodes 202, 204, 206) Note that a second capping applied as a restricted search ellipsoid was set at 100g/t Au and that any grade above said threshold was discarded during the interpolation process when estimating blocks outside this restricted search

ellipsoid. The restricted search ellipsoids for these zones vary from 20x15x10 to 30x26x10m.




JUNE 2020 14-19

Figure 14-12: Graphs supporting capping threshold decisions on composites for the Kiena Deep A Zone A1

(blockcodes 212, 214, 218) Note that a second capping applied as a restricted search ellipsoid was set at 55g/t Au and that any grade above said

threshold was discarded during the interpolation process when estimating blocks outside this restricted search ellipsoid. The restricted search ellipsoids for these zones vary from 20x15x10, and 30x26x10m.




JUNE 2020 14-20

Figure 14-13: Graphs supporting capping threshold decisions on composites for the Kiena Deep A Zone A2 – upper domain

(blockcodes 222-224)




JUNE 2020 14-21

Figure 14-14: Graphs supporting capping threshold decisions on composites for the Kiena Deep A Zone A2 – lower domain

(blockcode 228) Note that a second capping applied as a restricted search ellipsoid was set at 50g/t Au and that any grade above said

threshold was discarded during the interpolation process when estimating blocks outside this restricted search ellipsoid. The restricted search ellipsoid for this zone was set at 28x20x10m.




JUNE 2020 14-22

Figure 14-15: Graphs supporting capping threshold decisions on composites for the H1ZA Zone

(blockcode 300)




JUNE 2020 14-23

Table 14-3: Basic statistics on composites and high-grade capping value for each deposit Composites (1m) - Au (g/t)

Sector Count sample

Uncut mean

Std. dev.

COV Max Min Uncut

median

2019 capping

value

Number capped

% capped

Metal loss (%)

Cut mean

Cut COV

Zone S50 2,487 2.41 7.10 2.946 256.67 0.00 1.24 25 17 0.68 9.83 2.18 1.51

Zone VC1 788 5.69 62.88 11.06 1,738.79 0.00 0.64 30 15 1.90 49.75 2.80 1.98

Zone VC6 231 3.33 6.01 1.81 43.10 0.00 1.31 15 9 3.90 16.21 2.79 1.30

Zone South 2,343 1.49 4.04 2.72 105.60 0.01 0.34 20 10 0.43 6.05 1.39 1.83

Zone B 192 1.74 4.11 2.36 30.83 0.01 0.09 20 2 1.04 6.23 1.63 2.10

Kiena Deep Zone A 1,251 19.92 83.91 4.21 1,827.92 0.00 0.37 165 37 2.96 31.77 13.59 2.62

Kiena Deep Zone A1 635 11.18 45.98 4.11 654.02 0.01 0.30 110 15 2.36 31.78 7.64 2.81

Kiena Deep Zone A2 W and H 203 8.41 40.21 4.78 440.65 0.00 0.40 50 5 2.46 50.24 4.26 2.42

Kiena Deep Zone A2 L 200 39.06 108.01 2.77 886.84 0.00 2.44 200 8 4.00 29.89 27.30 1.91

Hanging Wall 1 - Zone A 72 3.34 13.61 4.08 101.97 0.00 0.05 20 2 2.78 47.85 1.73 2.57




JUNE 2020 14-24

14.5.4 Density

Bulk density is an important parameter used to calculate tonnages for the estimated volumes

derived from the resource-grade block model.

A total of 82 density measurements were conducted on the Project by Wesdome. The samples

selected were from a variety of lithologies located across the Property and included a range of

associated gold grades. The specific gravity (SG) measurement was determined by the water

displacement method. A summary of the SG data is presented in Table 14-4.

Table 14-4: Summary of the density measurements

Lithology Count SG SG (mean)

Prophyry Dyke 4 2.72

Mafic Dyke 3 2.73

Gabbro 9 2.83

Amphibolite 5 2.91

Schist 27 2.81

Flow breccia 1 3.00

Basalt 23 2.83

Komatiite 3 2.94

Quartz vein 7 2.80

TOTAL 82 2.82

For this MRE, a fixed density value of 2.80 was used for all the mineralized zones of the Kiena

Mine Complex corresponding to the mean of the SG data of the schist unit that hosts the majority

of the mineralized zones and the density value used when the mine was operating.

A fixed density of 2.00 g/cm3 was assigned to the overburden and 0.00 g/cm3 was assigned to

underground workings.

14.5.5 Variogram Analysis

A semi-variogram is a common tool used to measure the spatial variability within a zone.

Typically, samples taken far apart will vary more than samples taken close to each other. A

variogram gives a measure of how much two samples taken from the same mineralized zone will

vary in grade depending on the distance between those samples, and therefore allowing building

search ellipsoids to be used during interpolation.




JUNE 2020 14-25

Three dimensional directional variography using the Snowden Supervisor v8.11 software was

carried out on the composites. Variograms were modelled in the three orthogonal directions to

define a 3D ellipsoid for each unit. The most representative zone was used for each deposit. The

three directions of ellipsoid axes were set by using the variogram fans and visually confirmed with

geological knowledge of the deposit. Lag distances were set according to drillhole grid spacing

specific to the structural domain analyzed.

Then, a mathematical model was interpreted in order to best-fit the shape of the calculated

variogram for each direction. Three components were defined for the mathematical model: the

nugget effect, the sill, and the range.

All variography tests were modelled with a nugget effect, as determined from the downhole semi-

variograms and two spherical structures.

Table 14-5 presents the chosen variogram model parameters for each zone and Figure 14-16

illustrates an example of the variography results.

The nugget effect values range from zero to 38% and are typical of gold deposits.

Table 14-5: Variogram model parameters for each mineralized zone

First structure Second structure

Sector Blockcode Nugget Sill Range X

(m) Range Y

(m) Range Z

(m) Sill

Range X (m)

Range Y (m)

Range Z (m)

Zone S50 101 0.23 0.23 20 10 5 0.54 28 20 10

Zone VC1 114 0.38 0.11 35 12 5 0.51 50 25 10

Zone VC6 123 0.00 0.46 15 11 5 0.54 30 20 10

Zone South 132 0.05 0.64 15 15 5 0.31 45 32 8

Zone B 140 0.26 0.12 16 14 5 0.62 30 29 10

ZA_W 202 0.10 0.12 26 30 5 0.78 43 40 10

ZA_H 204 0.10 0.45 20 15 5 0.45 36 18 10

ZA_V 206 0.00 0.12 25 10 5 0.88 45 21 10

ZA1_W 202 0.10 0.12 26 30 5 0.78 43 40 10

ZA1_H 204 0.10 0.45 20 15 5 0.45 36 18 10

ZA1_L 206 0.00 0.12 25 10 5 0.88 45 21 10

ZA2_W 202 0.10 0.12 26 30 5 0.78 43 40 10

ZA2_H 204 0.10 0.45 20 15 5 0.45 36 18 10

ZA2_L 228 0.10 0.31 28 20 5 0.59 49 39 10

H1ZA 206 0.00 0.12 25 10 5 0.88 45 21 10




JUNE 2020 14-26

Figure 14-16: Example of the variography study for the South Zone 132

14.6 Data Analysis - Polygonal MRE

14.6.1 Outlier Handling

According to Turcotte and Pelletier (2009) the Wesdome deposit contains 2,940 samples grading

more than 0.00 g/t Au within the boundaries of the 21 mineralized zones. The capping value used

for the polygonal MRE was 67 g/t Au and was determined using a histogram plot and applied on

raw assays. The QP believes that the 2009 high-grade capping value of 67 g/t Au remains valid

and can be used for the 2019 MRE.

The capping value for the 7 other zones was set by Wesdome in 2014 at 34.28 g/t Au applied on

raw assays. Wesdome had used this value in the past when operating the Kiena mine. The QP

believes that the 2014 high-grade capping value established at 34.28 g/t Au is valid and can be

used for the 2019 MRE.




JUNE 2020 14-27

14.6.2 Density

According to Turcotte and Pelletier (2009), a specific gravity of 2.80 t/m3 was used for the

polygonal MRE which is consistent with the SG results obtained by water displacement method.

14.7 Block Modelling

The block model for the Kiena Mine area was set in Geovia® GEMS 6.8.2.2.

14.7.1 Block Model Parameters

The parameters provided in Table 14-6 were used for the current mineral resource estimate.

Individual block cells have dimensions of 5 m long (X-axis) by 5 m wide (Y-axis) by 5 m vertical

(Z-axis).

The size of the blocks were chosen in order to best match the drilling pattern, thickness of the

zones, complexity of the geology model and underground mine planning.

Table 14-6: Kiena Mine area block model parameters

Properties X (column) Y (row) Z (level)

Origin coordinates 12,600 11,800 3,100

Number of blocks 400 340 400

Block extent (m) 2,000 1,700 2,000

Block size (m) 5 5 5

Rotation 0

The block model was coded using the percent model method typical of Geovia GEMS™,

reflecting the proportion of each solid inside every block. All blocks falling within a solid were

assigned the corresponding solid block code.

14.7.2 Search Ellipsoid Strategy

The ranges of the ellipsoids used for the interpolation were established using the variography

study and correspond to the range of the first structure for the first pass, to the second structure

for the second pass and to two times the second structure for the third pass. The third pass was

designed to adequately populate all the block of the mineralized zones (Figure 14-17).




JUNE 2020 14-28

It is noteworthy to mention at this point that the classification was mostly based on drillhole

spacing and, therefore, some interpolated blocks were not converted into the Inferred

classification. Refer to section Mineral Resource Classification (Section 14.10) for more details.

Table 14-7 presents the orientation and ranges of the search ellipsoids for each pass.

In addition, a high grade limit or second capping value was used for the second and third pass

grade interpolation to restrict high grade impact at greater distance from the drillhole intersect

(Table 14-8).




JUNE 2020 14-29

Figure 14-17: Example of search ellipsoids for the S50 zones for the three interpolation passes




JUNE 2020 14-30

Table 14-7: Search ellipsoid ranges by interpolation passes

Pass 1 Pass 2 Pass 3

Sector Blockcode GEMS orientation Search ellipsoid ranges Search ellipsoid ranges Search ellipsoid ranges

Azimut Dip Azimut X (m) Y (m) Z (m) X (m) Y (m) Z (m) X (m) Y (m) Z (m)

Zone S50 100 - 104 267 50 157 20 10 10 28 20 10 56 40 20

Zone VC1 111 - 114 105 83 239 35 12 10 50 25 10 100 50 20

Zone VC6 123 28 59 231 15 11 10 30 20 10 60 40 20

Zone South 130 - 135 256 69 124 15 15 10 45 32 10 90 64 20

Zone B 140 286 54 170 16 14 10 30 29 10 60 58 20

ZA_W 202 345 0 255 26 30 10 43 40 10 86 80 20

ZA_H 204 272.7 41.6 195.5 20 15 10 36 18 10 72 36 20

ZA_V 206 110 -40 110 25 10 10 45 21 10 90 42 20

ZA1_W 212 345 0 255 26 30 10 43 40 10 86 80 20

ZA1_H 214 272.7 41.6 195.5 20 15 10 36 18 10 72 36 20

ZA1_L 218 287.4 63.2 215.7 28 20 10 49 37 10 98 74 20

ZA2_W 222 345 0 255 26 30 10 43 40 10 86 80 20

ZA2_H 224 272.7 41.6 195.5 24 12 10 36 18 10 72 36 20

ZA2_L 228 287.4 63.2 215.7 28 20 10 49 39 10 98 78 20

H1ZA 300 110 -40 110 25 10 10 45 21 10 90 42 20




JUNE 2020 14-31

Table 14-8: Restricted search ellipsoid parameters

Restricted search ellipsoid parameters

Rockcode Blockcode Range 1 Range 2 Range 3 Threshold value (g/t)

ZA_W 202 26 30 10 100

ZA_H 204 20 15 10 100

ZA_V 206 25 10 10 100

ZA1_W 212 26 30 10 55

ZA1_H 214 20 15 10 55

ZA1_L 218 28 20 10 55

ZA2_L 228 28 20 10 50

14.7.3 Interpolation Parameters

Estimation and search parameters were evaluated through Kriging Neighbourhood Analysis

(KNA) and contact analysis.

KNA was conducted on each unit and on the most representative shear zones with the Snowden

Supervisor software. KNA provides a quantitative method of testing different estimation

parameters (i.e. block size, discretization and min/max of composites used for the interpolation)

by evaluating their impact on the quality of the results. The interpretation of these helps select the

optimal value for each parameter.

Following this study, the parameters provided in Table 14-9 were chosen for the interpolation of

the block model.

Table 14-9: Interpolation parameters

Interpolation parameters Pass 1 Pass 2 Pass 3

Minimum number of composites used 12 8 4

Maximum number of composites used 24 24 24

Maximum number of composites per drillhole used 4 4 4

Minimum number of drillholes used 3 2 1




JUNE 2020 14-32

14.7.4 Interpolation Methodology

The interpolation was run on a set of points extracted from the capped composited data. The

block model grades were estimated using ordinary kriging (OK) methods. Hard boundaries

between the mineralized zones were used in order to prevent grades from adjacent zones being

used during interpolation. As a block was estimated, it was tagged with the corresponding pass

number.

For comparison purposes, additional grade models were generated using: 1) inverse distance

squared (ID2); 2) nearest neighbour (NN); and 3) OK on uncapped composited data.

14.8 Block Model Validation

Every step of the block modelling process was revised to ensure fair representation of the

available data in the Block Model resource model.

More specific validations were completed on the block model including visual review of the

interpolated grades in relation to the raw and composited data, checks for global and local bias,

graphical validation (swath plots), statistical analysis of the model and comparison to other

estimation methods.

14.8.1 Visual Validation

Block model grades were visually compared against drillhole composite grades and raw assays in

cross-section, plan, longitudinal and 3D views (Figure 14-18). This visual validation process also

included confirming that the proper coding was done within the various domains and checks for

global and local bias.

The visual comparison shows that the block model is consistent and correlate well with the

primary data without excessive smoothing.

Visual comparisons were also conducted between ID2, OK and NN interpolation scenarios. The

OK scenario used for the resource estimate produced a grade distribution honouring drillhole data

and the style of mineralization observed on the Kiena Mine area.




JUNE 2020 14-33

Figure 14-18: Comparative example of the grade distribution between the blocks and the composites in section (A) and longitudinal (B) views

14.8.2 Statistical Validation

Grade averages for the OK, NN and the ID2 models were tabulated in Table 14-10. This

comparison did not identify significant issues. As expected, block grade averages are generally

lower than the composite grades and initial grades were well represented throughout the

estimation process.

The average grades generated by the ID2 interpolation method are very close to those reported

from the OK interpolation method. This information provides a general indication that the resource

model is reasonable.




JUNE 2020 14-34

Table 14-10: Comparison of the block and composite mean grades at a zero cut-off grade for Inferred and Indicated blocks

(blocks >50% inside a mineralized zone)

Sector Number of

composites Composite grade

(g/t Au) Number of blocks

OK grade model (g/t Au)

ID2 grade model (g/t Au)

NN grade model (g/t Au)

S50_100 522 1.77 543 0.71 1.86 1.89

S50_101 558 1.93 717 1.53 1.84 1.98

S50_102 844 2.97 686 2.92 2.84 3.05

S50_103 315 1.51 373 1.34 1.34 1.13

S50_104 248 1.75 1044 1.96 2.04 1.53

VC1_111 218 3.11 362 2.77 2.96 3.18

VC1_112 74 2.80 184 2.87 2.75 4.68

VC1_113 113 3.42 326 3.62 3.64 4.70

VC1_114 383 2.44 499 1.79 2.46 2.58

VC6_123 231 2.79 352 3.05 2.73 2.94

ZS_130 613 1.41 3271 1.22 1.29 1.47

ZS_131 600 1.70 2127 1.61 1.77 1.74

ZS_132 788 1.18 2389 1.43 1.50 1.61

ZS_133 170 1.33 988 1.44 1.46 1.26

ZS_135 172 1.28 500 1.32 1.41 1.19

ZB_140 192 1.63 1244 1.62 1.74 1.48

ZA 1251 13.59 2838 12.44 11.14 14.27

ZA1 635 7.64 1900 7.03 6.95 7.09

ZA2 403 15.69 1617 11.46 11.84 13.20

H1ZA 72 1.73 88 1.16 1.58 0.70




JUNE 2020 14-35

14.8.3 Swath Plots

Swath plots were also generated as part of the block model validation. A swath plot is a graphical

display of the grade distribution derived from a series of bands (or swaths), generated in several

directions throughout the deposit. Using the swath plots, grade variations from the OK model are

compared to the distribution of grade interpolated with the NN and ID2 methods and to the

composite grades. This validation method also works as a visual mean to identify possible bias in

the interpolation.

Figure 14-19 to Figure 14-21 illustrate a series of swath plots in the three directions. Generally,

the grades estimated in the blocks are close to the average grades provided by the data source;

no bias was found in the resource estimate in this regard.

Figure 14-19: Block model validation swath plot along strike (X-direction) for the ZA Zone




JUNE 2020 14-36

Figure 14-20: Block model validation swath plots across strike (Y-direction) for the ZA Zone




JUNE 2020 14-37

Figure 14-21: Block model validation swath plots along elevation (Z-direction)

Based on visual and statistical reviews, it is the QP’s opinion that the Kiena Mine area block

model provides a reasonable estimate of in situ gold resources.

14.9 Polygonal Mineral Resource Estimate methodology

The polygons were created on inclined longitudinal sections for each individual zone. All drillholes

intersecting gold-bearing zones (composites) were identified on the sections and assigned as the

polygon grades. Each polygon was assigned a unique name. Polygon limits were defined by the

mid-distance between two drillhole intercepts, or the maximum distance from the pierce point,

defined as 30 m or 40 m for the Wesdome Deposit (Figure 14-22).

The polygon tonnage was estimated for each polygon using the area of the polygon on the

inclined longitudinal section multiplied by the true thickness of each intercept and the specific

gravity.




JUNE 2020 14-38

Figure 14-22: Longitudinal section 12200 N +/- 50m looking north of the Dubuisson zone showing the polygonals for the 2019 MRE

14.10 Mineral Resource Classification

The mineral resources for the Kiena Mine Complex were classified in accordance with CIM

Standards.




JUNE 2020 14-39

14.10.1 Mineral Resource Definition

The “CIM Definition Standards for Mineral Resources and Reserves” prepared by the CIM

Standing Committee on Resource Definitions and adopted by the CIM council on May 10, 2014,

provides standards for the classification of Mineral Resources and Mineral Reserves estimates as

follows:

Inferred Mineral Resource:

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.

Indicated Mineral Resource:

An Indicated Mineral Resource is that part of a Mineral Resource for which quantity,

grade or quality, densities, shape and physical characteristics are estimated with

sufficient confidence to allow the application of Modifying Factors in sufficient detail to

support mine planning and evaluation of the economic viability of the deposit.

Geological evidence is derived from adequately detailed and reliable exploration,

sampling and testing and is sufficient to assume geological and grade or quality

continuity between points of observation.

An Indicated Mineral Resource has a lower level of confidence than that applying to a

Measured Mineral Resource and may only be converted to a Probable Mineral Reserve.

Measured Mineral Resource:

A Measured Mineral Resource is that part of a Mineral Resource for which quantity,

grade or quality, densities, shape, and physical characteristics are estimated with

confidence sufficient to allow the application of Modifying Factors to support detailed

mine planning and final evaluation of the economic viability of the deposit.

Geological evidence is derived from detailed and reliable exploration, sampling and

testing and is sufficient to confirm geological and grade or quality continuity between

points of observation.

A Measured Mineral Resource has a higher level of confidence than that applying to

either an Indicated Mineral Resource or an Inferred Mineral Resource. It may be

converted to a Proven Mineral Reserve or to a Probable Mineral Reserve.




JUNE 2020 14-40

14.10.2 Mineral Resource Classification for the Block Model MRE

Following the previous definitions, the estimated block grades were classified into Inferred and

Indicated Mineral Resource categories using drill spacing, a minimum number of drillhole and

recognition of grade and geological continuity within the zones.

No Measured resources were defined for the Kiena Mine area at this stage.

The following parameters were used for the classification:

▪ Inferred Mineral Resources were defined for blocks within the units that have been informed

by a minimum of two drillholes within 40 m of a drillhole (80 m of drill spacing);

▪ Indicated Mineral Resources were defined for blocks within the units that have been

informed by a minimum of three drillholes within 12.5 m of a drillhole (25 m of drill spacing).

When needed, a series of clipping boundaries were created manually in longitudinal views to

either upgrade or downgrade classification in order to homogenize the groups of resources by

removing artificial features and isolated blocks or group of blocks due to automatically generated

classification. All remaining estimated but unclassified blocks were flagged as “Exploration

Potential”.

Figure 14-23 shows an example of the classification.




JUNE 2020 14-41

Figure 14-23: Mineral Resource classification example for the S50_102 Zone

14.10.3 Mineral Resource Classification for the Polygonal MRE

The classification for the Polygonal MRE was established in 2009 by Turcotte and Pelletier for the

Wesdome deposit and by Turcotte et al. in 2015 for the other seven zones.

For this MRE, the QP reviewed the classification and undertook the following changes since the

last MRE:

▪ Removing isolated polygons for lack of plausible economic viability (Figure 14-24);

▪ Downgrading the measured resources into indicated resources as no recent drillholes nor

drift resampling were conducted on this part of the property to assess the quality of the

historical data;

▪ Discarding zones with less than three drillholes;

▪ Discarding the resources that were disclosed within a 100 m crown pillar (Figure 14-24 and

Figure 14-24).




JUNE 2020 14-42

For all remaining polygons, the classification was defined by the following rules:

For the Wesdome deposit:

▪ Indicated Mineral Resources were defined for polygons with a maximum radius of 15 m from

drillhole intercepts (equivalent to a 30 m of drill spacing) where a cluster of DDH with similar

results was observed;

▪ Inferred Mineral Resources were defined for polygons with a maximum radius of 40 m from

drillhole intercepts (equivalent to 80 m of drill spacing) where a cluster of DDH with similar

results was observed.

For the other zones:

▪ Indicated Mineral Resources were defined for polygon with a maximum radius of 30 m from

drillholes intercepts (60 m of drill spacing) where a cluster DDH with similar results is

present. The average radius from drillhole intercepts for the indicated resources is 14.81 m;

▪ No Inferred Mineral Resources were estimated.




JUNE 2020 14-43

Figure 14-24: Polygonal MRE and discarded area example for the 2019 MRE Longitudinal section 12200 N +/- 50 m looking north of the Dubuisson Zone




JUNE 2020 14-44

Figure 14-25: Polygonal MRE and of discarded area example for the 2019 MRE

Longitudinal section looking N30E of the Wesdome zone E1

14.11 Cut-off Grade

According to CIM’s Definition Standards, in order for a deposit to be considered a Mineral

Resource it must be proven that there are “reasonable prospects for eventual economic

extraction”. This requirement implies that the quantity and grade estimates meet certain economic

thresholds and that the Mineral Resources are reported at an appropriate cut-off grade that takes

into account extraction scenarios and processing recoveries.

The underground cut-off grade used for the Mineral Resource Estimate was 3.0 g/t Au for the

zones with >40° dip and 4.0 g/t Au for the shallow-dipping zones (<40° dip). The cut-off grades

were calculated using a gold price of USD1,300 per ounce, a CAD:USD exchange rate of 1.31,

mining cost of $110/t (>40° dip); $150/t (<40° dip), processing cost of $35/t, and G&A of $15/t.

The cut-off grades should be re-evaluated in light of future prevailing market conditions (metal

prices, exchange rate, mining cost, etc.).

It is the QP’s opinion that the cut-off grades are relevant to the grade distribution of this project

and that the mineralization exhibits sufficient continuity for economic extraction under the cut-offs

applied.




JUNE 2020 14-45

14.12 Kiena Mine Complex Mineral Resource Estimate





Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Block Model MRE 968,900 14.46 450,400 1,121,200 11.02 397,100

Polygonal MRE 1,859,300 5.65 337,800 1,796,900 6.94 401,000

TOTAL 2,828,200 8.67 788,100 2,918,100 8.51 798,100

Notes to Table 14-11:

These mineral resources are not mineral reserves as they do not have demonstrated economic

viability.


Results are presented in situ and undiluted and considered to have reasonable prospects for

economic extraction, below 100 m crown pillar.

The estimation combined two estimation methods, ordinary kriging in the Kiena Mine Complex and

polygonal for other deposits on the property.

The Kiena Mine Complex resources encompass 20 zones with a minimum true thickness of 3.0 m

using the grade of the adjacent material when assayed or a value of zero when not assayed. High-

grade capping varies from 20 g/t to 200 g/t Au (when required) was applied to composited assay

grades for interpolation using an Ordinary Kriging interpolation method based on 1.0 m composite and

block size of 5 m x 5 m x 5 m, with bulk density values of 2.8 (g/cm3). In addition, a high grade limit or

second capping value was used for the second and third pass grade interpolation to restrict high grade

impact at greater distance from the drillhole intersect. Indicated resources are manually defined and

encloses areas where drill spacing is generally less than 25 m, blocks are informed by a minimum of

three drillholes, and reasonable geological and grade continuity is shown.

The zone outside the Kiena Mine Complex encompasses eight zones with a minimum true thickness

of 1.5 m using a polygonal estimation method. Indicated resources were estimated from drillhole

results using the mid distance between drillhole or a maximum of 30 metres, 12.5 metres in some

areas. The high-grade capping was fixed at 34.28 g/t Au with a bulk density value of 2.8 (g/cm3).

The estimate is reported for potential underground scenario at cut-off grades of 3.0 g/t Au (> 40° dip)

and 4.0 g/t Au (< 40° dip, Wesdome Zone). The cut-off grades were calculated using a gold price of

USD1,300 per ounce, a CAD:USD exchange rate of 1.31 (CAD1,700); mining cost $110/t (> 40° dip);

$150/t (< 40° dip); processing cost $35/t; G&A $15/t. The cut-off grades should be re-evaluated in light

of future prevailing market conditions (metal prices, exchange rate, mining cost, etc.).

The number of metric tons and ounces were rounded to the nearest hundred and the metal contents

are presented in troy ounces (tonne x grade / 31.10348).







JUNE 2020 14-46

Table 14-12 and Table 14-13 show the breakdown of the polygonal and block model MRE at the

selected cut-off grade of 3.0 g/t Au and at 4.0 g/t for the Wesdome deposit. All tables for the

Polygonal MRE can be found in Appendix C.

Table 14-12: Indicated and Inferred block model Mineral Resource Estimate per lens

Cut-off grade


Lens Tonnage

(t) Grade (g/t)

Ounces Au (oz)

Tonnage (t)

Grade (g/t)

Ounces Au (oz)

S50_100 3.0 9,906 4.13 1,315 - - -

S50_101 3.0 22,768 4.42 3,235 9,611 4.23 1,308

S50_102 3.0 89,151 4.95 14,182 8,326 4.70 1,259

S50_103 3.0 6,644 4.45 951 796 3.56 91

S50_104 3.0 4,384 3.67 517 62,282 3.84 7,679

VC1_111 3.0 43,833 5.21 7,348 3,338 4.03 433

VC1_112 3.0 12,818 5.01 2,065 13,134 5.89 2,487

VC1_113 3.0 25,741 5.93 4,909 24,801 7.47 5,954

VC1_114 3.0 5,650 4.91 892 30,172 6.92 6,713

VC6_123 3.0 30,851 4.59 4,552 22,617 4.78 3,476

ZS_130 3.0 5,821 3.64 681 43,921 4.58 6,466

ZS_131 3.0 9,885 4.26 1,355 86,744 4.02 11,215

ZS_132 3.0 4,006 4.21 543 32,056 3.34 3,438

ZS_133 3.0 6,267 4.68 943 25,088 4.43 3,574

ZS_135 3.0 11,975 4.68 1,804 15,476 4.06 2,022

ZB_140 3.0 - - - 66,581 4.24 9,067

ZA_200 3.0 430,771 20.97 290,366 166,213 18.99 101,470

ZA1_210 3.0 185,961 13.26 79,259 187,331 10.60 63,845

ZA2_220 3.0 59,620 18.31 35,105 322,650 16.06 166,637

H1ZA_300 3.0 2,847 4.38 401 - - -

Total block model MRE 968,900 14.46 450,400 1,121,200 11.02 397,100




JUNE 2020 14-47

Table 14-13: Indicated and Inferred Polygonal Mineral Resource Estimate per zone

Zones Cut-off grade


Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Martin Zone 3.0 236,400 4.86 37,000 - - -

Dubuisson Zone 3.0 281,500 5.46 49,400 - - -

Dubuisson North 1 Zone 3.0 193,700 7.67 47,800 - - -

Dubuisson North 2 Zone 3.0 80,600 6.71 17,400 - - -

Northwest Zone 3.0 467,400 3.79 57,000 - - -

Presqu'ile 1 Zone 3.0 91,800 6.64 19,600 - - -

Presqu'ile 2 Zone 3.0 51,200 8.67 14,300 - - -

Wesdome deposit 4.0 456,700 6.49 95,300 1,796,900 6.94 401,000

Total polygonal MRE 1,859,300 5.65 337,700 1,796,900 6.94 401,000

Table 14-14 shows the sensitivity of the block model estimate to grade cut-off for the in situ

underground MRE.

The reader is cautioned that the numbers presented in the following tables should not be

misconstrued with a mineral resource statement.

Table 14-14 shows a 3D view of the grade distribution and classification of the Kiena Deep A

Zone A.

Table 14-14: Kiena Mine Complex Block Model Indicated and Inferred Mineral Resource cut-off grade sensitivity table


Cut-off grade

Tonnage (000 t)

Grade (g/t)

Ounces Au oz)

Tonnage (000 t)

Grade (g/t)

Ounces Au (oz)

> 5.00 g/t 666,652 19.28 413,133 6,51,808 16.24 340,277

> 4.50 g/t 717,200 18.25 420,841 716,675 15.20 350,163

> 4.00 g/t 781,013 17.11 429,549 808,615 13.95 362,694

> 3.50 g/t 860,071 15.88 439,064 930,947 12.61 377,395

> 3.00 g/t 968,899 14.46 450,420 1,121,198 11.02 397,140

> 2.50 g/t 1,104,691 13.02 462,375 1,379,812 9.46 419,884




JUNE 2020 14-48

Figure 14-26: Example of 3D views showing grade distribution and classification of the Kiena Deep A Zone A




JUNE 2020 15-1

MINERAL RESERVE ESTIMATE

Since this Report summarizes the results of a Preliminary Economic Assessment (PEA), no

Mineral Reserves have been estimated for the Kiena Mine Complex project as per NI 43-101

guidelines.




JUNE 2020 16-1

MINING METHODS

16.1 Historical Mining

The Kiena Mine Complex, which is currently under care and maintenance, is situated 10 km west

of Val-d’Or, Quebec, on an 8 km by 9 km land parcel. The Kiena headframe and 2,000 tpd mill are

on Parker Island, located towards the southwest end of the 8-km-wide Lac de Montigny and are

accessible by road via a causeway. Historical production of the mine is 1.8 million ounces since

1981.

The first gold discoveries were made on Parker Island (site of the present Kiena Mine surface

infrastructure) during the period 1911 to 1914. Kiena Gold Mines Ltd. (Kiena Mines) was

established in 1936 as part of the ventures group of Thayer Lindsley. A shaft with four levels was

developed. The Parker Vein had limited extensions. The North Zone was identified by surface

drilling and a drift was extended north from the shaft on the 120-metre level. This drift encountered

the VC Zone and continued north to provide a drilling platform to test the continuity of the North

Zone. In 1940, the operation was abandoned due to limited reserves, economic conditions and the

war effort.

Twenty years later, the property’s potential was re-examined by Falconbridge Ltd with drilling work

in 1961 establishing the S50 Zone. By 1963, exploration work had outlined a resource of 4.5 million

tonnes averaging 6.34 g Au/tonne for the S50 Zone. Shaft No.1 (the current shaft) was sunk to a

depth of 400 m and extensive underground exploration and development was completed. In 1965,

the project was again abandoned due to adverse mining conditions and the low gold price of

USD 35 per ounce.

In 1976, Wesdome Resources Ltd was created and in 1979 with an improvement in gold prices,

the project was successfully developed with commercial production commencing in October 1981.

In 1986, Kiena Mines was sold to Campbell Red Lake Mines, which subsequently merged with

Dome Mines Ltd. and Placer Development Ltd. to form Placer Dome Canada Ltd.

In 1997, Placer Dome sold Kiena to the now-defunct Montreal-based junior McWatters Mining.

McWatters ran the mine until 2002, when it was closed due to low gold prices. Initial production

was custom milled offsite until Kiena built their own mill, which commenced in September 1984.

The mill treated 1,300 tpd from the S50 Zone. In 1999, McWatters Mines Inc. decided to develop

the North Zone and expand the mill to 2,000 tpd capacity. The strategy was to increase production

and lower unit costs by providing 1,000 tpd from the S50 Zone and 1,000 tpd from the North Zone.

In all, between 1981 and 2002, Kiena produced 1.56 million oz of gold from 10.7 Mt grading

4.75 g/t Au, with virtually all the mineralized material coming from the S50 Zone. Between 1985

and 2002, output averaged 70,000 oz of gold annually at a cash cost of USD 234/oz.




JUNE 2020 16-2

In October 1999, Dynacor Mines Inc. and Western Quebec Mines signed an agreement to

consolidate the contiguous Siscoe and Siscoe-Extension (Dynacor Mines) and Wesdome,

Lamothe, Lamothe-Extension, Yankee Clipper and Callahan (Wesdome Resources) properties into

a new company, Wesdome Gold Mines Inc.. December 2003 saw the purchase of the Kiena Mine

Complex by Western Quebec Mines, which subsequently placed the property into Wesdome Gold

Mines Inc., thereby completing and consolidating Wesdome’s land package around Lac De

Montigny. Then, in July 10, 2007, a merger was completed with parent company Western Quebec

Mineson the basis of 1.45 shares of Wesdome for each share of Western Quebec Mines.

Wesdome was the surviving operating entity.

In 2006, Kiena Mine completed a pre-production development plan and brought the mine into

production in August 2006 until March 7, 2013. Wesdome suspended mining activities at the Kiena

mine on June 30, 2013 due to decreasing recovered grades, persistent industry cost pressures and

uncertainty in the gold price and exchange rates. At this time the Proven and Probable Mineral

Reserves were reclassified as Measured and Indicated Mineral Resources. Figure 16-1 illustrates

Kiena Mine production from 1981 until 2019.

Figure 16-1: Historical production 1981 to 2019

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Kiena Mine Historical Production (1981 to 2019)

Milled Tonnes Delivered Ounces Mill Grade




JUNE 2020 16-3

Since the suspension of operations, Wesdome has completed extensive underground diamond

drilling to enhance the mineral resource in four zones of the mine (South, South Deep, VC and

Kiena Deep). Figure 16-2 illustrates the four zones that are considered within the scope of this PEA.

Figure 16-2: Mining zones included in this PEA

16.2 Existing Material/Waste Handling Facilities

16.2.1 Shaft

The existing shaft (Shaft No. 1) is a three-compartment timbered shaft developed in 1963 to a depth

of 400 m. Compartment Nos. 1 and 2 house a cage/skip conveyance that has a capacity of 6.2 t

(refer to Figure 16-3) for skipping and 5,000 pounds for material in the cage. Compartment 3 is for

mine services (electrical cables, compressed air pipe, process water pipe and dewatering pipe) and

ladderway. Due to the size limitation of the cage, transport of consumable material such as pipe,

screen, ground support bolts, were suspended below the cage. Most levels of the mine are

accessible from an internal ramp that starts on 17 Level (17L). The main ventilation raise from

surface to 17L is used to sling mobile equipment and large items underground utilizing a surface

crane.




JUNE 2020 16-4

In the early 1990s, the shaft was deepened to 930 m with a fourth compartment added to enable

sinking without impacting production. The hoist is a Nordberg double-drum double-drive with GE

motors (refer to Figure 16-4).

Figure 16-3: Shaft No. 1 compartments

Figure 16-4: Hoist drives and motors




JUNE 2020 16-5

Hoisting of waste and mineralized material is a manual system and requires a hoist operator at the

hoist controls, as well as a skip tender at the loading pocket. The following information was

considered in assessing the hoisting system availability and capacity:

▪ Material movement;

▪ Personnel movement;

▪ Daily hoist inspections and maintenance;

▪ Monthly, quarterly and yearly inspections;

▪ Conveyance changeouts;

▪ Rope changes;

▪ Unplanned outages.

This information indicated that an average hoisting capacity per day of 2,000 tonnes can be

expected.

Hoisted mineralized material dumped into the headframe bin is conveyed to the storage bins

located between the headframe and crusher (Refer to Figure 16-5).

Figure 16-5: Surface mineralized material bins




JUNE 2020 16-6

16.2.2 Loading and Haulage

All mineralized material and waste loading and hauling is performed by diesel haulage trucks load

by Load Haul Dumps (LHD). The haulage trucks haul the mineralized material to the orepass

system that feeds the crusher directly, or to the truck dump on 79L, which also feeds the 81L

crusher. Any material dumped in the orepasses is sized to prevent damage to the pass or hang-

ups within the pass. Waste material is dumped into a waste pass system that ends on 81L, where

it passes through a grizzly with a rock breaker for sizing before passing through to the loading

pocket.

Mobile Equipment

As of writing of this Report, the existing mobile equipment underground is not suitable or

mechanically fit to support the life of mine (LOM) plan. Any equipment required for exploration

development is contractor owned.

16.2.3 Crushing

A 48” jaw crusher is located on 81 Level. Two passes feed a scalper/breaker before the crusher to

allow fines to bypass the crusher and break larger material before entering the jaws (refer to

Figure 16-6). The bin below the crusher feeds directly to the loading pocket located on 88 Level.

Figure 16-6: Crusher

16.2.4 Waste Disposal

Any waste created underground that is not disposed of into underground workings will be hoisted

to surface and stored at the current waste facilities.




JUNE 2020 16-7

16.2.5 Existing Services Facilities

Ventilation System

The ventilation system at Kiena Mine is a push-pull system.

Fresh Air

Two surface fresh air fans operated in parallel (refer to Table 16-1) push fresh air into the mine via

raises; surface to 12L and 12L to 17L. From 17L the fresh air travels via the main ramp system

down to the lower works on 100L.

Table 16-1: Surface fresh air fans

Fan Diameter Blade setting Horsepower

(hp) Volume KCFM

Pressure in WG

Sheldon Axico 200-8-10 79” 47o 200 170 4

Joy 2000 72-20-1200 72” 15o 250 172 4.5

Return Air

A single return air fan pulls air from the mine via a network of raises. There is a blockage in the

return air raises system between 17L and 33L that is currently forcing a higher than design amount

of return air into the shaft. By-pass raises are being driven in 2020 to resolve this issue.

Table 16-2: Return air fan

Fan Diameter Blade setting Horsepower

(hp) Volume KCFM

Pressure in WG

Joy M- 84-58-1200 84” 36o 600 300 7.6

The current surface fresh air and return air fan installations have capacity on the fan operating

curves and should not require upgrading for the project to go forward. However, these fans have

been operating for several years, in particular the Sheldon brand fan, therefore a full mechanical,

structural, and electrical inspection of all the fans should be completed before the project moves

forward. Of main concern is the lack of available parts for the Sheldon fan as this company is no

longer operating.

The current mine air heater systems may require upgrading, based on the heater capacities, before

the project proceeds. This is a mixed glycol and natural gas system and it should be determined if

they are properly functioning by way of a full mechanical, structural, and electrical inspection.




JUNE 2020 16-8

Repair Facilities

Repair facilities located on 64L are available to complete mobile equipment repairs and services

(wash bay and service bay), as well as a small shop to repair hand held drills and air pumps. A

separate area exists for welding on equipment (refer to Figure 16-7). Electrical parts and minor

electrical repairs are performed in the small facility also located on 64L

Figure 16-7: Repair facilities

Fuel and Oil Facilities

Located on 64L is the main fueling station and oil storage. There is a fuel line located in the shaft

that is used to fill the fuel tanks in this facility. Construction of a second fuel facility was started

before the closure of the mine in 2013 on 82L, and is planned to be completed upon the restart of

the mine with fuel delivery via a lined borehole from 64L. Currently, the area is being utilized for

storage while developing the exploration ramp.

Figure 16-8: Second fueling station

Backfill Facilities

An existing hydraulic backfill plant exists at the mine and will be utilized for creating backfill for all

stoping activities. Historically 30% to 35% of the tails has been utilized for backfill material.




JUNE 2020 16-9

Dewatering Facilities

The current dewatering system at the mine is a clean water system. The current pumping stations

are located on 38L, 52L, 82L and 94L. Run of mine (ROM) water flows to decant sumps with slimes

being disposed of underground and the clean water decanted into storage sumps that feed

standard multi-stage dewatering pumps (refer to Figure 16-9). All pumping stations have backup

pumps installed to maintain the continual operation of the pumping system. The main dewatering

pipe line is a mild steel Schedule 40 6” diameter line installed in the service compartment (No. 3)

of the shaft.

Figure 16-9: Dewatering pumps

Compressed Air Facilities

Compressed air for the underground workings at 690 kPa (100 psi) comes from the surface

compressors (2 x Sullair LS-25S compressors) located in a building attached to the hoist room

(refer to Figure 16-10 and Figure 16-11). A diesel standby compressor located in a separate

building is available if there is any issue with the main compressors. Distribution underground is via

a mild steel Schedule 40 8” diameter pipe located within the service compartment in the shaft

(Compartment No. 3). The pipeline size in the underground workings varies in size from 2” to 4”

diameter mild steel Schedule 40 pipe.




JUNE 2020 16-10

Figure 16-10: Surface compressor room

Figure 16-11: Surface compressors

Process Water Facilities

Process water from the mine is obtained from a well situated near the shore of Lac de Montigny.

Water is pumped to the shaft for distribution underground and to surface facilities. A 3” diameter

mild steel Schedule 40 pipe with Pressure Reducing Valves (PRV’s) distributes process water to

the levels underground (refer to Figure 16-12).

For drinking water underground, bottled water is supplied for individuals and refuge/lunchrooms.




JUNE 2020 16-11

Figure 16-12: Process water distribution




JUNE 2020 16-12

Underground Electrical Facilities

Power for underground operations comes from the surface electrical room located behind the hoist

room and feeds from Transformers T7 and T11 (refer to Figure 16-13). There are currently four

circuits dedicated to underground power.

Figure 16-13: Transformers T7 and T11

The cables for the four electrical circuits (72, 73, 1102 and 1103) feeding underground equipment

are located within the service compartment in the shaft. Each feed is 1,200 A and services specific

levels within the mine.

Communications Facilities

Radio communications is available throughout the active mining areas utilizing a leaky feeder

system. Hard wire telephones are also in use at specific locations, such as: refuge stations, work

shops, and shaft stations.

Central blasting is utilized throughout the mine with the main cable located within Compartment 3

of the shaft.

Refuge Facilities

Existing refuge facilities consist of excavations in drift walls with concrete walls and steel doors.

Compressed air is supplied via the compressed air system underground.

Second Egress Facilities

The main ramp is the second egress from the Mine along with the Fresh Air Raise with ladder ways

for workings below the shaft and from 17L to surface.




JUNE 2020 16-13

16.3 Geotechnical

16.3.1 Structure and Lithology

The Kiena Deep Zone and S50/B Zone are two discrete zones, located below existing Kiena mine

workings. Kiena Deep consists of three coplanar Sectors (ZA-200, ZA1-210, ZA2-220) with dips

that become shallower with depth, from ~80°(±) to 50°(±). In plan, the Kiena Deep Sectors follow

an arched shape due to an isoclinal fold associated with the Marbenite Fault (MF). The S50/B zones

consists of six Sectors with planar shape (ZB-140 and the S50-100, S50-101, S50-102, S50-103

S50-104), dipping at ~60°(±). Fault zones are present in the hanging wall of both A and B Lenses.

The Kiena Deep and S50/B zones are hosted within the contact of the basalt rock and schist, with

talc-chlorite schist associated with faulting. Ultramafic komatiite rock may be present in the hanging

wall of the Marbenite Fault. Zones are typically laminated quartz Sectors (shear veins) hosted within

sheared basalt, sheared ultramafic or sheared talc-chlorite schist.

16.3.2 Rock Strength

Laboratory unconfined compressive strength (UCS) testing of representative core from Kiena Deep

and S50/B zones has not been performed. As an initial reference, Table 16-3 provides historic

uniaxial compressive testing (σC), summarized in Golder (2010). As no core testing of ultramafic

rock has been performed at the site, anticipated range of rock strength is assumed.

In general, ISRM characterization of the rock units would assume the following:

▪ Basalt (ISRM: R4+), Strong rock;

▪ Ultramafic (ISRM: R3), Medium strength;

▪ Talc-Schist (ISRM: R2 to R3), Weak to Medium strength.

Table 16-3: Rock strength assumptions

(Golder, 2010)

Litho unit Compressive

strength, σC (MPa) Tensile strength,

σt (MPa) Youngs modulus, E

(GPa) Poisson ratio, ν

Basalt, V3 152 -19 66.3 0.29

Basalt stockwork, V10 103 -18 64.9 0.25

Ultramafic, V4 40 to 60 (estimate) No data

Talc-Schist, M8 Tc-Ca 34 -4 31.6 0.34




JUNE 2020 16-14

16.3.3 Rock Mass Characteristics

Measurements of rock strength, rock quality designation (RQD), and structural joint sets form a

basis for classifying the various rock types Kiena Deep Zone and S50/B Zone rock mass was

characterized using the industry-standard Barton et al. (1974) Rock Tunnelling Quality Index, Q’

methodology. The modified Rock Tunnelling Quality Index, Q’, is defined as:

Q’ = RQD/Jn × Jr/Ja Equation 1

Where:

▪ RQD represents the Rock Quality Designation. Logged RQD values from the drill database

were used;

▪ Jn is the Joint Set number. Typical values range from three sets (Jn = 9) to three + random

(Jn = 12) joint sets;

▪ Jr is the Joint Roughness number. Typical values ranges from Jr = 1 (smooth, planar joint

surfaces) to Jr = 1.5 (irregular to rough planar joint surfaces);

▪ Ja (Joint Alteration) values from Ja =1 (clean joint surfaces) to Ja = 2 (slightly altered joint

surfaces) in the basalt rock units. Ja values for mineralized areas reflect the presence of

material coating on the joint surfaces, ranging between Ja = 2, Ja = 3 (non-softening joint

coatings) to Ja = 4 (cohesive material coating on joint surfaces).

The rock mass characteristics for Kiena Deep and S50/B zones were assessed from available

diamond drilling data (RQD, lithology) and corresponding core photos (Jn, Jr and Ja). Core data

from diamond drillholes (DDH) 6322, 6326, 6363, 6409, 6446 and 6456 was examined in detail.

Core data and photos of the hanging wall and footwall rock masses, within 25 m of the mineralized

zones, was examined in detail. All holes were drilled into Kiena Deep Zone, with the exception of

DDH 6363, which targeted S50/B Zone.

Results are summarized in Table 16-4 and Table 16-5. Rock mass characteristics vary with

lithology. Lithology of Kiena Deep Zone includes the presence of ultramafic rock. Pillars located

between the discrete Sectors in Kiena Deep Zone are characterized as a unique unit (Inter-vein

Pillar). Of the three lithologies present in the Inter-vein Pillar, ultramafic and schist are dominant.




JUNE 2020 16-15

Table 16-4: Kiena Deep Zone, typical rock mass characteristics

(DDH 6322, 6326, 6409, 6446, 6456)

Location Lithology RQD (%) Jn Jr Ja Q’

Footwall

Schist 72 (53) 9 (12) 1 3 (4) 2.7 (1.1)

Ultramafic (sheared)

48 (38) 9 (12) 1.5 (1) 3 (4) 2.6 (0.8)

Hanging wall Basalt 91 (77) 9 (12) 1.5 1 (2) 15.1 (4.8)

Ultramafic 89 (65) 9 (12) 1.5 (1) 3 (4) 4.9 (1.4)

Mineralization Basalt 92 (79) 9 (12) 1.5 1 (2) 15.3 (4.9)

Schist 75 (60) 9 (12) 1 3 (4) 2.8 (1.3)

Inter-vein Pillar

Basalt 90 (80) 9 (12) 1.5 1 (2) 15 (5)

Schist 67 (46) 9 (12) 1 3 (4) 2.5 (1.0)

Ultramafic 91 (84) 9 (12) 1.5 (1) 3 (4) 5.1 (1.8)

NOTE: Brackets represent low-range values.

Table 16-5: S50/Deep B Zone, typical rock mass characteristics

(DDH 6363 core)

Location Lithology RQD (%) Jn Jr Ja Q’

Footwall Basalt 76 (63) 9 (12) 1.5 1 12.7 (7.9)

Hanging wall Basalt 59 (33) 9 (12) 1.5 1 9.8 (4.1)

Schist 59 (33) 9 (12) 1 3 (4) 2.2 (0.7)

Mineralization Basalt 70 (60) 9 (12) 1.5 (1) 1 (2) 11.7 (2.5)

NOTE: Brackets represent low-range values.

16.3.4 Stress Assumptions

In situ stress measurements have been performed at the Kiena mine site in the mid-1990s. Results

are summarized in Arjang (1996) and Corthésy et al. (1996). Stress measurements performed by

Corthésy et al. (1996) at 810 m depth obtained the following values: σ1=61.2 MPa (subhorizontal);

σ2=44.9 MPa (subhorizontal); σ3=26 MPa (subvertical).

The 810 m depth is considered representative of the stress setting in the Deep A and Deep B

zones. The anticipated principal stress tensor trend (Azimuth 035°; subhorizontal) is based on

mean principal stress assumptions from mines within the Cadillac Fault region (McKinnon and

Labrie, 2006). Stress assumptions, based on the 810 stress measurements, are provided in

Table 16-6. The Kiena Deep and S50/B mineralized zones strike sub-perpendicular to the principal

stress tensor, see Figure 16-14.




JUNE 2020 16-16

Table 16-6: Stress assumptions, Kiena Deep and S50/B zones

Stress tensor Trend

σ1 (MPa) = 2.35 • σ3 Horizontal trend on AZ 035°

σ2 (MPa) = 1.72 • σ3 Horizontal trend, normal to σ1

σ3 (MPa) = σv = 0.032 • D

(D = depth below surface, m) Vertical

Figure 16-14: Trend of the principal stress tensor (Azimuth 035°) relative to the geometry of Kiena Deep and S50/B zones

16.3.5 Anticipated Rock Mass Behaviour

Based on the information currently available, it is anticipated that the rock mass behaviour will be

mainly controlled by stress, as well as spacing of discontinuities and potentially locally by fault and

shear features. The occurrence, persistence and characteristics of the major geological structures

and rock mass fabric will control the rock mass behaviour around mine openings.

With depth and increased extraction in the lower Kiena zones, mining induced stresses are

expected to concentrate around the excavation abutments and are anticipated to be generally

observable as localized fracturing and light to moderate spalling in the high range rock masses.

Moderate to significant wall deformation may develop in low-range, foliated rock masses.

Figure 16-15 shows a schematic of expected rock mass failure modes for anticipated range of rock

mass quality and stresses. The dashed box highlights rock mass behaviour at increased depth.




JUNE 2020 16-17

Figure 16-15: Excavation behaviour matrix as a function of rock mass quality and stress (depth)

Kaiser et al., (2000). The dashed line illustrates the anticipated range of rock quality and stress for the lower Kiena zones.

16.3.6 Ground Support Requirements

Anticipated ground support requirements vary with lithology.

a) Development in basalt rock:

▪ Anticipated rock mass behaviour: Stiff, blocky ground. Stress accumulation may

generate surface spalling and minor crushing at drift corners;

▪ Primary support: Install resin-grouted rebar with weldmesh screen across the roof and

walls to 1.5 m above floor. Split sets (1.5 m length) may be an alternative for support of

the mid and lower walls;

▪ Secondary support options for wide-span intersections and stope brows include:

Grouted cable bolts, resin (spin) cables, Super Swellex or 3.6 m grouted steel bar.




JUNE 2020 16-18

b) Development in schistose and ultramafic rock:

▪ Anticipated rock mass behaviour: Non-linear deformation under load. Stress

accumulation may generate drift convergence and potential unravelling of under-

supported ground;

▪ Primary support: Install friction bolts (Standard Swellex in the roof and walls) with

weldmesh screen across the roof and walls to 1.5 m above floor. Splitsets (1.5 m length)

may be an alternative for support of the mid and lower walls. Alternative bolts, such as

D-Bolt or Hybrid Bolt, may have an application;

▪ Secondary support options for wide-span intersections include: Grouted cable bolts or

Super Swellex.

16.4 Mining Method

16.4.1 Longhole Mining

Future mining will continue utilizing the historical longhole mining method as illustrated in

Figure 16-16.

Figure 16-16: Longhole mining




JUNE 2020 16-19

Development will be performed utilizing standard development methods with electric hydraulic

jumbos, LHD and scissor lifts for services and ground support. Sills will be developed utilizing the

same equipment and will be driven longitudinally along the lenses from either the existing

development or new ramps (VC and Kiena Deep zones).

Bottom-up mining is planned with level mining retreating longitudinally along the lenses towards

the access. Production mining will utilize electric-hydraulic top hammer drills (downhole and uphole

capable) and LHDs for mucking. Mining activities for a stope will consist of:

▪ Slot raise drilling;

▪ Production hole drilling;

▪ Slot blasting and mucking;

▪ Production blasting and mucking;

▪ Backfilling.

Once completed, adjacent stopes can be mined after the backfill set period.

16.4.2 Underhand Longhole Mining

Underhand longhole mining will be utilized when mining under previously mined areas. This will

occur in the Kiena Deep Zone when mining fronts advance under a completed mining front (refer

to Figure 16-17). The same equipment will be utilized for mining, and, at this time, upper drilling for

the stopes.

In the PFS, a review is required to identify if the development of a top sill through the backfill of the

previously mined stopes is a more suitable methodology.




JUNE 2020 16-20

Figure 16-17: Crown recovery

16.4.3 Mineable Resource

Twenty resource wireframes have been identified (refer to Table 16-8) as potential mining areas

and combined into four mining zones as shown in Figure 16-18. Each wireframe was evaluated

using Deswik Stope Optimizer (DSO), which creates mineable shapes based on specific

parameters input into Deswik for each lens (refer to Table 16-7). These mineable shapes based on

this level of study (PEA) have not been proven and are preliminary in nature and cannot be

considered for inclusion as reserve calculations.




JUNE 2020 16-21

Table 16-7: Wireframe evaluated using Deswik Stope Optimizer




JUNE 2020 16-22

Table 16-8: Resource wireframes utilized

Zone Lens

South

ZS-130

ZS-131

ZS-132

ZS-133

ZS-135

S50/ B

ZB-140

S50-100

S50-101

S50-102

S50-103

S50-104

VC

VC1-111

VC1-112

VC1-113

VC1-114

VC6-123

Kiena Deep

ZA-200

ZA1-210

ZA2-220

H1ZA-300

Figure 16-18: Mining zones




JUNE 2020 16-23

The key parameters used are illustrated in Table 16-9.

Table 16-9: DSO key parameters

Parameter Unit Value

Price of Gold(1) USD/ounce 1,300

Exchange Rate CAD:USD 1.31

Mining Cost CAD/tonne 100.00

Processing Cost CAD/tonne 40.00

G&A Cost CAD/tonne 25.00

Processing Recovery % 97

Kiena Deep Mining Recovery % 90

South, VC and South Deep (S50) Mining Recovery % 80

Note: (1) Financial evaluations have been completed using USD1,532/oz

The results of this work identified potential Inferred Resource (refer to Table 16-10) and Indicated

Resource (refer to Table 16-11) that are available to mine and are summarized in Table 16-12.

Table 16-10: Potential Inferred Resource


(tonne)

Contained gold in DSO shapes

(Oz's)

DSO grade (g/t)

Inferred

South 146,785 18,367 3.89

VC 93,525 16,954 5.64

S50 & B 110,160 13,128 3.71

Kiena Deep 863,911 324,172 11.67

Total 1,214,381 372,621 9.54

Table 16-11: Potential Indicated Resource


(tonne)


(Oz's)

DSO grade (g/t)

Indicated

South 26,654 3,267 3.81

VC 114,522 17,103 4.65

S50 & B 117,931 16,059 4.24

Kiena Deep 909,042 395,763 13.54

Total 1,168,149 432,192 11.51




JUNE 2020 16-24

Table 16-12: Total Resource


(tonne)


(Oz's)

DSO grade (g/t)

All

South 173,439 21,633 3.88

VC 208,047 34,057 5.09

S50 & B 228,091 29,187 3.98

Kiena Deep 1,772,954 719,935 12.63

Total 2,382,530 804,813 10.51

A review of the shapes and their location, with respect to existing infrastructure and planned

infrastructure, was completed and DSO shapes that did not contain enough value for development

and mining costs have been removed (refer to Table 16-13) from the resource as shown in

Figure 16-19, Figure 16-20 and Figure 16-21. Two lenses, ZS-132 and H1ZA-300, were completely

removed as there were insufficient values to develop them at the time of this study.

Table 16-13: Resource removed


(tonne)


(Oz's)

DSO grade (g/t)

All

South 39,157 4,283 3.40

VC 9,996 1,214 3.78

S50 & B 19,140 2,122 3.45

Kiena Deep 1,240 161 4.04

Total 69,532 7,780 3.48




JUNE 2020 16-25

Figure 16-19: Resource removed from South Zone

Figure 16-20: Resource removed from VC Zone




JUNE 2020 16-26

Figure 16-21: Resource removed from South Deep (S50)

A recovery factor was applied to the remaining resources specific to each zone and is illustrated in

Table 16-14. This recovery accounts for mining loss of tonnes expected within each zone. The

internal dilution reported in Table 16-14 is the percentage of waste included in the resource

calculations.

Table 16-14: Resource recovered by zone

Zone Resource Recovered (%) Dilution (%)

South 80 7

VC 80 12

South Deep 80 8

Kiena Deep 90 21

Table 16-15 illustrates that a total of 2.07 Mt grading 10.64 g/t resource remains to be included in

the mine schedule.




JUNE 2020 16-27

Table 16-15: Resource for schedule

Class Zone Scheduled

(tonnes) Gold in schedule

(Oz's) Schedule grade

(g/t)

Inferred

South 96,552 12,237 3.94

VC 74,329 13,252 5.55

S50 & B 77,592 9,228 3.70

Kiena Deep 794,823 291,755 11.42

Sub-total 1,043,296 326,473 9.73

Indicated

South 12,661 1,643 4.04

VC 88,828 13,023 4.56

S50 & B 93,040 12,424 4.15

Kiena Deep 836,530 356,042 13.24

Sub-total 1,031,059 383,131 11.56

Total

South 109,213 13,880 3.95

VC 163,157 26,275 5.01

S50 & B 170,631 21,652 3.95

Kiena Deep 1,631,354 647,797 12.35

Total(1) 2,074,354 709,604 10.64

Note: (1) Table 16-15 has an additional 4,111 tonnes at 4.08 g/t from the South Zone that was not included

in Table 16-20..

16.5 New and Upgraded Material Handling Facilities

16.5.1 Shaft

The existing facilities will be utilized for mining of the South, South Deep (S50), VC and Kiena Deep

zones. Included in the financials is an upgrade of the hoists and addition of instrumentation to

support the automation of the mineralized material and waste rock hoisting system.




JUNE 2020 16-28

16.5.2 Loading and Hauling

Mobile Equipment

Table 16-16 illustrates the new fleet of mobile equipment that will be purchased to support the mine

plan. All LHDs, jumbos and production drills have been identified for use in autonomous operation

and can be converted to such operational parameters as the need arises.

Table 16-16: New fleet of mobile equipment

Type Class Price 2021 2022 2023 2024 2025 2026 2027 2028 2029

LHD's

LH307 $745,427 1

LH410 $880,326 2 1

LH514 $1,327,068 1

Trucks TH430 $950,643 1 1 2

Jumbo's DD421 $1,267,305 2 1

Production Drill DL-411-15 $1,307,683 1 1

Boom Truck A64 Boom $535,440 1 1

Scissor Lift A64-SL $455,166 2 2

Misc. Personnel Carriers

A64-PC $388,285 1 1 1

U/G Grader Grader M845 $450,000 1

Surface Loader CAT 988 $1,200,000 1

Development and production drills are diesel driven and electric/hydraulic operated. All prime

movers (LHDs and trucks) are diesel as well as all miscellaneous equipment. Use of electric and/or

battery powered equipment will be investigated during the prefeasibility study.

16.5.3 Crushing

The existing facilities will be utilized for mining of the South, South Deep (S50), VC and Kiena Deep

zones. Funding has been allocated for completing repairs and changing of key parts on the crusher.

16.5.4 Waste Disposal

Any waste created underground that is not disposed of into underground workings, as backfill or

abandoned headings, will be hoisted to surface and stored at the current waste facility.




JUNE 2020 16-29

16.5.5 New and Upgraded Service Facilities

As the mine is currently under care and maintenance, all existing infrastructure is considered

operational and fit for use unless otherwise indicated. Table 16-17 indicates the number of new

facilities that will be required for mining of the identified resource.

Table 16-17: New service facilities

U/G Construction Item Unit price 2021 2022 2023 2024 2025 2026 2027 2028 2029

Kiena Second Egress (metres) $2,550 90 180 180 180 140

VC Second Egress (metres) $2,550 100 100 100 60

Refuge Stations (each) $215,000 1 1 1

Fuel/Oils Station (each) $180,000 1 1

Communications/Automation (each)

$75,000 1 1 1 1

Backfill Stations (each) $15,000 2 5 8 5

Kiena Deep Electrical Stations (each)

$535,000 1 2 2 2

VC Electrical Stations (each) $535,000 1 1

Repair Bay (each) $624,623 1

Latrine (each) $36,710 1 2 2

Fuse Mag. (each) $63,776 1

Explosive Mag. (each) $137,954 1

Kiena Deep Booster Fan Installation

$150,000 1

VC Zone Booster Fan Installation $150,000 1

Ventilation Facilities

Based on the proposed development for both the Kiena Deep and the VC Zone two additional

return air booster fan installations will be required. The VC booster fan will be in a transfer drift

located on 81 Level (Refer to Figure 16-22). This fan will pull approximately 110,000 cfm from

12 Level through a series of old and new raises and through the VC Zone. This fan will have a duty

of approximately 100 hp @ 3.0 inches total pressure.




JUNE 2020 16-30

Figure 16-22: VC Zone booster fan location

The Kiena Deep booster fan will be in a transfer drift at the top of a new dedicated exhaust raise

(Refer to Figure 16-23). This raise is sized at 10 ft x 10 ft, with an overall length of approximately

2,500 feet. The transfer drift located on 94 Level will feed into the main Kiena exhaust air system.

The selected fan will pull approximately 130,000 cfm through the Kiena Deep fresh air ramp and

have a duty of approximately 250 hp @ 7.5 inches total pressure.




JUNE 2020 16-31

Figure 16-23: Kiena Deep Zone booster fan location

As the new booster fans will be increasing the overall pressure on the system it is recommended

that a thorough inspection be made of all doors and bulkheads, and where required these ventilation

infrastructure items be repaired, replaced, or upgraded.

It is possible that cooling may be required as the Kiena Mine goes progressively deeper. As is

evident with other operations in the Abitibi area cooling may be required, based on diesel equipment

usage, re-use of air, and seasonal fluctuations. Going forward a heat study with placed diesel

equipment will be required to confirm these requirements for cooling. However, the initial choice

would be to use increased air volume as the primary cooling mechanism. If required, a refrigeration

system for Kiena Deep could be located near the 81 Level Booster station which would afford a

location for exhausting the hot reject air, whilst cooling the ramp air.




JUNE 2020 16-32

Repair Facilities

Existing facilities will continue to be used as the main areas to complete repairs. An excavation will

be created in Kiena Deep to serve as a small repair area for minor repairs (hoses, change tires,

change filters etc.).

Fuel and Oil Facilities

Temporary oil facilities will be created in the Kiena Deep Zone only. These facilities will be limited

in quantity as they are designed to support equipment when limited amount of oil is required.

Backfill Facilities

The existing backfill plant will continue to be utilized and the existing network of underground

pipelines and boreholes will be expanded to support the new mining areas. In addition, a process

must be established to meet Section 84 of the Quebec “Regulation Respecting Occupational Health

and Safety in Mines Act respecting occupational health and safety (chapter S-2.1, ss. 223, 286,

294 and 310)” dated April 1, 2019 illustrated below.

“84. Where tailings are used for backfilling underground excavations, the water contained

in such residues and leaking therefrom may not have a cyanide content higher than

0.005%, expressed in potassium cyanide.”

Expansion of the system into the VC and Kiena Deep Zones will be required via boreholes and

lateral pipe networks. Waste from ramp and level access development will be utilized as backfill

when permitted. In the Kiena Deep Zone, consolidated fill will be required in areas of multi-veins

and in the sills of each Mining Front to allow mining under previously mined areas.

As the amount of waste development begins to decline (after 2024), the amount of tails required

for backfill will increase to the extent where most backfill will be from consolidated tails.

Dewatering Facilities

Future dewatering of zones below the current dewatering system will utilize a dirty water pumping

system. As the mine deepens, the amount of ground water inflowing into the mine is not expected

to increase and the new pumping system is designed to handle the process water created by the

proposed equipment, dust suppression, and backfill.

Compressed Air Facilities

No additional upgrade is anticipated to the compressed air system. Required pipe and installation

is included in the cost of development.




JUNE 2020 16-33

Process Water Facilities

No upgrade is anticipated to the process water system. Required pipe and installation is included

in the cost of development.

Electrical Facilities

Electrical services will be fed from the existing circuits in the mine for South, VC and a portion of

the South Deep zones. A dedicated circuit consisting of a 15 kV 500 MCM cable from the surface

electrical room to 100L via the shaft is required for the Kiena Deep Zone and portions of the South

Deep Zone. Seven electrical stations have been considered for Kiena Deep Zone and two for VC

Zone.

Communications Facilities

The existing communications systems will be advanced to support mining operations.

Teleoperation of equipment from surface has not been considered for inclusion at this time.

Refuge Facilities

Kiena Deep and VC are the two zones that will require the installation of additional refuge facilities.

Second Egress Facilities

Internal ladder ways will be installed in the ventilation raises for Kiena Deep and VC zones to act

as a second egress.

16.6 Development

Development is scheduled to start on the face locations identified by Kiena Mine personnel and will

commence in 2021. During the year 2020, some development is planned for exploration purposes

and will be completed by a contractor. The year 2021 will be a year of transition for the mine as it

will move from contractor development to Company development crews. This transition period will

take place over 18 months as equipment arrives for the mine crews. Based on the zone wireframes,

ramps and level access drifts have been designed to access the identified potential mineable

resource. In the South and South Deep zones, some mineable shapes can be accessed from

existing ramps and levels. Table 16-18 indicates the required lateral ramp and access development

required to support the Mine Production Schedule.




JUNE 2020 16-34

Table 16-18: Capital development schedule

Zone Unit Maximum monthly

Total 2021 2022 2023 2024 2025 2026 2027 2028 2029

South m 84 3,803 985 986 986 847 0 0 0 0 0

Avg. m/d 2.8 2.7 2.8 2.8 2.8 2.6 0.0 0.0 0.0 0.0 0.0

VC m 84 4,513 985 985 986 988 569 0 0 0 0

Avg. m/d 2.8 2.8 2.8 2.8 2.8 2.8 2.7 0.0 0.0 0.0 0.0

S50 & B m 84 1,846 986 841 19 0 0 0 0 0 0

Avg. m/d 2.8 2.5 2.8 2.4 0.6 0.0 0.0 0.0 0.0 0.0 0.0

Kiena m 167 6,565 762 1,548 1,710 1,658 883 0 0 0 0

Avg. m/d 5.6 3.9 2.1 4.3 4.8 4.6 3.7 0.0 0.0 0.0 0.0

Total m 395 16,727 3,718 4,360 3,701 3,493 1,452 0 0 0 0

Avg. m/d 13.2 9.9 10.4 12.2 10.4 9.8 6.1 0.0 0.0 0.0 0.0

Lateral development sizing is based on existing drift/ramp sizing and equipment requirements to

meet mining regulations and the production/development schedule. Table 16-19 illustrates planned

typical opening sizes.

Table 16-19: Development sizes

Type Width (m) Height (m)

Ramp 4.5 4.5

Access 5.0 4.5

Truck loading 7.0 7.0

Ventilation access 4.5 4.5

Electrical rooms 5.0 4.5

Sills 4.5 4.5

Ramp and access designs have been completed for all zones and are shown in Figure 16-24 to

Figure 16-27.




JUNE 2020 16-35

Figure 16-24: South Zone access

Figure 16-25: VC Zone access




JUNE 2020 16-36

Figure 16-26: South Deep Zone access

Figure 16-27: Kiena Deep Zone access




JUNE 2020 16-37

16.7 Production Plan

The production plan is based on a mill start-up of mid-2021; therefore, only sill production is planned

in the first half of 2021, which will be stockpiled. Each zone was reviewed, and a composite mining

rate developed in tonnes per day (tpd) that accounts for sill development and the mining activities

(drill, blast, muck and fill). This composite rate was applied to the zone tonnage to produce the LOM

production schedule (refer to Table 16-20). Kiena Deep being the largest of the four zones has

been divided into Seven Mining Fronts (A to G) for scheduling purposes (refer to Figure 16-28).

Each Mining Front consists of five levels to enable flexibility in the mining plan.




JUNE 2020 16-38

Table 16-20: Life of mine production

Kiena Deep Zone Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029


Mining Rate 585 tpd 45 250 569 610 627 684 761 729 609

Mined Grade Kiena Deep 12.35 g/t 9.77 11.42 12.23 11.40 11.01 12.62 13.36 12.54 14.81

Mined 647,797 ounce 5,052 32,672 79,676 79,848 79,025 98,812 116,351 105,010 51,351

VC Zone Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029


Mining Rate 93 tpd 13 75 99 99 99 72 0 0 0

Mined Grade VC 5.01 g/t 4.86 5.69 4.94 4.40 5.43 4.69 0.00 0.00 0.00

Mined Gold 26,275 ounce 733 4,900 5,621 5,004 6,179 3,839 0 0 0

S50 & B Zone Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029


Mining Rate 80 tpd 80 80 80 80 80 79 0 0 0

Mined Grade South Deep 3.95 g/t 3.71 3.91 3.42 3.92 4.22 4.52 0.00 0.00 0.00

Mined Gold 21,652 ounce 3,393 3,577 3,133 3,601 3,862 4,087 0 0 0

South Zone Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029

Trammed Mineralized Material 105,102 tonne 0 28,235 28,470 28,548 19,849 0 0 0 0

Mining Rate 79 tpd 0 79 80 80 56 0 0 0 0

Mined Grade South 3.95 g/t 0.00 3.82 3.93 4.05 4.01 0.00 0.00 0.00 0.00

Mined Gold 13,341 ounce 0 3,468 3,595 3,718 2,560 0 0 0 0

All Zones Total Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029

Operating Days 3,027 day 356 356 356 357 356 356 356 357 177


Mining Rate 684 tpd 138 484 829 869 862 834 761 729 330

Mined Grade 10.65 g/t 5.80 8.05 9.70 9.24 9.28 11.18 13.36 12.54 14.81

Mined Gold 709,065 ounce 9,177 44,616 92,025 92,171 91,626 106,737 116,351 105,010 51,351

Backfill Required 1,219,930 tonne 29,018 101,636 173,845 182,865 180,911 175,050 159,609 153,451 63,546

Backfill (from Tails) 790,804 tonne 2,902 25,409 121,692 128,005 126,638 122,535 111,726 107,415 44,482




JUNE 2020 16-39

Figure 16-28: Kiena Deep Mining Fronts

Figure 16-29, Figure 16-30 and Figure 16-31 show the tonnes, ounces and grade profile by year

for each zone for the LOM Plan.




JUNE 2020 16-40

Figure 16-29: Mining production - tonnes by year for each zone

Figure 16-30: Mining gold ounces by year for each zone




JUNE 2020 16-41

Figure 16-31: Mining grade by year for each zone




JUNE 2020 17-1

RECOVERY METHODS

This chapter presents information based on the “NI 43-101 Technical Report and Mineral Resource

Estimate for the Kiena Mine Complex" in Val-d'Or, Quebec dated November 7, 2019 (Richard and

Torrealba, 2019). The report presented the available information on historical mineral processing

at the Kiena mine (Anon, 1983 and Dupont, 1996).

17.1 Process Description

The Kiena Mine processing plant became operational in September 1984. A conventional gold

recovery process involving cyanidation and conventional CIP was used. The principal process

steps included crushing, grinding, leaching by cyanidation, gold adsorption and desorption,

electrolysis, cyanide destruction, melting and casting of doré bars. Figure 17-1 shows the process

flowsheet of the Kiena plant as at December 2011.

The following sections provide a detailed description of the Kiena processing plant.




JUNE 2020 17-2

Figure 17-1: Kiena Mine process plant flowsheet (December 2011)




JUNE 2020 17-3

17.2 Process Plant Design Criteria

The Kiena Mine Complex was designed with a plant throughput of 1,250 tpd, however crushing

capacity was added, and the plant has run as high as 2,000 tpd in the past. Figure 17-2 presents

the throughput versus year based on the latest LOM as described in Chapter 16 (Table 16-21).

Figure 17-2: Throughput versus year based on LOM

A new tailings storage facility (TSF) will be available in year 2024. It is envisioned that, during the

first three years of operation (namely Phase 1), the process plant will operate “as-is” at a maximum

throughput of 829 tpd until the new TSF is operative. In year 2024 (Phase 2), the throughput will

increase (maximum 869 tpd) based on the latest LOM, and a new tailings plant facility will be

operative (see Section 18.10). The processing plant flowsheet will be modified to include the

addition of a filtration plant and the cyanide destruction circuit will be refurbished.

According to the current production plan, 38% of the tailings will be used for backfilling on average

over the life of the mine, with the remaining being stored at the TSF. This value is lower for the

years 2021 and 2022 (6% and 15% respectively) as more waste is available. From 2023 to 2029,

it is estimated that 41% of the tailings will be used for backfilling. There may be an opportunity to

store most of the tailings for the first two years underground behind hydrostatic plugs (to be

investigated in the PFS).




JUNE 2020 17-4

The design criteria is based on a process plant design throughput of 829 tpd (years 2021 to 2023)

and 869 tpd (year 2024 and beyond). The design values were selected from the highest throughput

from each phase (Figure 17-2).

Table 17-1 presents an overview of the main criteria parameters of the original design as well as

the criteria for the first and second future phases (years 2021 to 2023, and year 2024 and beyond).

Table 17-1: Project process design criteria for different scenarios

Parameter Unit Original design

Phase 1 (years 2021-2023)

Phase 2 (year 2024 and

beyond)

Plant yearly throughput tpy 400,000 295,010 310,316

Plant daily throughput tpd 1,250(3) 829 869

Average Au feed grade g/t 7.85 11.7

Crushing plant utilization % N/A 65 65

Process plant utilization % 95 92 92

Plant hourly throughput @ % utilization tph 54.8 37.5 39.4

Gravity circuit - No TBD TBD

Grind size to Leach/CIP, P80 microns 75 75(1) 75(1)

Leaching Retention time h 44 34.1 32.4

CIP Retention time h 4 6.0 5.7

Overall Au recovery % 96.2 95.7-100(2) 95.7-100(2)

Carbon stripping, regeneration capacity t 2 2 2

Residual cyanide concentration at plant discharge, (max)

- N/A TBC TBC

Final tailings density target % 32 53 85

(1) Based on the former operation. To be confirmed in the prefeasibility stage.

(2) Based on the latest testwork. To be confirmed in the next phase of testwork.

(3) Throughput of up to 2,000 tpd has been processed in the past.

Figure 17-3 and Figure 17-4 show the block flow diagrams and preliminary mass and water balance

for Phase 1 (years 2021 to 2023), and Phase 2 (year 2024 and beyond).




JUNE 2020 17-5

.

Figure 17-3: Phase 1 (years 2021 to 2023) block flow diagram and mass balance




JUNE 2020 17-6

Figure 17-4: Phase 2 (year 2024 and beyond) block flow diagram and mass balance




JUNE 2020 17-7

17.3 Process Plant Facilities Description

17.3.1 Crushing Circuit

The crushing circuit starts underground with a Birdsboro Buchanan jaw-crusher, reducing the rock

size to 6 inches. The mineralized material entering the plant has therefore a maximum grain size

of 6 inches. The former design plant throughput is 1,250 tpd, however, with the addition a screen

and cone crusher, the plant has run as high as 2,000 tpd. The crushed material receiving facilities

start with a 35-tonne capacity hopper equipped with a 30" x 10' vibrating feeder and a 30-inch belt

conveyor, which transfers the crushed material onto the existing No. 1 belt conveyor. Conveyor

No. 1 will transport the crushed material in two coarse rock silos, both with a capacity of 600 tonnes.

The material discharged from the silos is conveyed, screened and crushed to approximately 1¼"

(32 mm). The crushing unit is a standard cone crusher (150 kW) operating in an open circuit. The

crushed and screened material is then stored in a 1,800-tonne silo. The crushed material is then

conveyed by vibrating feeders to the grinding circuit via a belt conveyor.

17.3.2 Grinding Circuit

The mineralised material will be ground in a 1,000 hp semi-autogenous (SAG) mill (11'6" x 18'8")

operating in an open circuit, followed by a 900 hp ball mill (10'6" x 13') operating in closed circuit

and two stages of cyclones for classification. If necessary, cyanide and lead nitrate can be added

at the SAG mill (Figure 17-5) and ball mill. The pH is controlled using quicklime in the feed to the

SAG mill.

Figure 17-5: SAG mill at the Kiena Mine Complex




JUNE 2020 17-8

The underflow of the cyclones will be redirected to the ball mill for further grinding. Based on

potential future gravity testwork, a gravity circuit (currently not shown on the PFDs) may be installed

and fed by a portion of the feed or underflow of the cyclone (to be confirmed by testwork). The

cyclone overflow will then be pumped to a vibrating 20-mesh screen to remove trash (wood chips,

plastic, particles, etc.). The screen underflow will be directed to a 65-ft diameter process thickener

to increase the density of the pulp.

17.3.3 Cyanidation

Thickener overflow will be pumped to a series of three carbon columns tanks where the gold in

solution will be adsorbed by the carbon. The last carbon column in the series will overflow into a

carbon screen where the oversize will be sent to the second CIP tank. The screen undersize

(solution with cyanide, collected in the grinding solution tank) will be recycled back to the grinding

circuit as grinding solution.

Thickener underflow will be forwarded to a series of three leach tanks. Cyanide and quicklime are

added to the leach feed. Additionally, lead nitrate and oxygen will be added to optimize the gold

dissolution and control the pH. The cyanidation reaction is the following:

4 Au + 8 NaCN + O2 + 2 H2O ↔ 4 NaAu(CN)2 (aq) + 4 NaOH

Retention time in the carbon columns and leach tanks is approximately 11 minutes and 36 hours

respectively, based on a plant feed of 1,250 tpd.

17.3.4 Carbon-in-pulp (CIP) Process

The leach tails are pumped to the next unit operation, the CIP circuit. The CIP circuit is composed

of five CIP tanks in series. At the CIP circuit, the slurry is fed at the first CIP tank in the row. The

activated carbon is fed at the last CIP tank and pumped in counter-current mode relative to the

slurry. The NaAu(CN)2 molecules in solution in the CIP tanks are held by the activated carbon by

adsorption.

The discharge of the CIP tanks is pumped to a 28-mesh safety screen to recover any carbon

particles escaping from the tanks.

Figure 17-1 presents the tailings handling configuration (as at December 2011), where the safety

screen undersize can be treated to produce backfill or sent to tailings, but there was no cyanide

destruction option (Section 17.3.6 presents an alternative proposed configuration). In Phase 1, the

slurry is sent to primary and secondary backfill cyclones. The cyclones' underflow (coarse particles)

is then sent to the backfill plant. The fine particle overflow from the cyclones is directed to the 75-ft

diameter waste-thickener, from where the fine slurry is sent to the tailings. The dilution water is

recycled toward the grinding circuit.




JUNE 2020 17-9

The loaded carbon is pumped to the elution vessel, where the gold desorption takes place using

the Zadra process at a temperature of 140°C and a pressure of 80 psi. The pregnant solution

obtained is pumped through an electrowinning cell where gold is recovered by plating on steel wool

cathodes. The cathodes are washed under pressurized water and the dried sludge is then melted

in an induction furnace to produce doré bars.

17.3.5 Acid Wash and Carbon Regeneration

Once desorption is complete, the carbon will be transferred to the wash vessel where it will be

washed with hydrochloric acid. It will then be forwarded to the regeneration furnace where it will be

heated up to a temperature of 1,050°C. Finally, the carbon will be screened to the desired grain

size in order to be reused in the process.

17.3.6 Cyanide Destruction Circuit and Tailings Treatment

To support the planned higher tonnage in Year 3 (Phase 2), the current process plant cyanide

destruction circuit which is not operational will be re-furbished. A filtered tailings circuit will also be

added. The cyanide destruction (CND) circuit will treat the thickened CIL residue slurry (tailings

thickener underflow) at 60% (w/w) solids and the overflow of the backfill cyclones. Addition of

process water will be required to set the CND feed at 50% solids. Cyanide destruction is completed

using the Inco SO2/Air process.

The concept for CND process considers that the reaction of cyanide destruction occurs in two tanks

arranged in series (or parallel), providing a retention time of an hour (TBC in future testwork). Liquid

SO2 is added to the tanks and process air is injected through cone spargers located at the bottom

of each tanks and oxidize the cyanide species present. Copper sulphate is added to catalyze the

destruction reaction. Hydrated lime addition controls the pH in each tank. An agitator in each tank

ensures adequate mixing and gas dispersion.

As part of the dry stack initiative for storing process plant tailings a filter plant will be installed in

Phase 2. Figure 18-2 presents the proposed location for the filtering plant. The feed to the filter

plant (30 tpd capacity) at approximately 50% solids will be processed by a pre-filter thickener that

will increase the percentage solids to 60%. The underflow of the thicker will be feed to a stock tank

that will act as a buffer for the filter press circuit (two in operation and one stand-by). The dry product

at 85% solids will be trucked to the TSF and the filtrate will be pumped back to the pre-filter to

remove suspended solids from the water. The overflow (O/F) of the pre-filter thickener will be fed

to the thickener O/F tank and the water will be recirculated back to the process water reservoir or

to the fresh water tank.




JUNE 2020 17-10

17.3.7 Reagents Systems

A summary of the reagents required in the process plant is presented in Table 17-2 along with the

expected form of supply and mixing requirements.

Table 17-2: Reagent mixing system

Reagent Delivery Preparation

Sodium Cyanide Tankers liquid No preparation required

Quicklime Trucks solid Mixing tank, water addition

Flocculant Bags solid Eductor, mixing tank, water addition to in-line mixer

Activated Carbon Super sacks solid Attrition tank mixing with water to remove fine carbon

Anti-scalant Drum liquid No preparation required

Caustic Soda Tanker liquid Dilution tank, water addition

HCl Totes liquid Dilution tank, water addition

Lead Nitrate Super sacks solid Mixing tank, water addition

SO2 Liquid Tanker liquid No preparation required (pressurized system)

Copper sulphate (CuSO4.5H2O)

Super sacks solid Mixing tank, water addition

Due to the lower than design throughput, it is assumed that the site storage capacities are enough

for each reagent or consumables.

17.4 Energy, Water and Consumable Requirements

17.4.1 Energy Requirements

The average electrical energy requirements for the process plant were derived from Kiena’s

historical reports, basic principles, and equipment list.

Each motorized item of equipment was assigned utilization, efficiency, and load factors to derive

the data presented in Table 17-3.




JUNE 2020 17-11

Table 17-3: Process plant power demand by phase and area

Area

Phase 1 (years 2021 to 2023) Phase 2 (year 2024 and beyond)

Connected load (kW)

Yearly consumption (GWh)

Connected load (kW)

Yearly consumption (GWh)

Crushing and grinding 366 2.4 607 4.0

Pumps and compressors 157 0.9 375 2.1

Material handling, screening, agitation and other equipment

109 0.6 216 1.2

Heating 812 4.5 812 4.5

Filtration plant - - 124 0.7

Network loss (2%) - 0.2 - 0.2

Tolerance (+/- 10%) - 0.9 - 1.3

Total 1,448 9.4 2,134 13.9

17.4.2 Water Requirements

The water requirements for the plant are divided into two main areas, fresh water and process

water.

Fresh water is sourced from the lake and it is used in the following areas: carbon elution, reagent

preparation, gland seal water.

Process water is used through the plant and is separated into two distinct systems: cyanide

bearing (grinding solution: tailings thickener overflow) and non-cyanide bearing (clarifier overflow –

Phase 2). The recycling of cyanide bearing water reduced the fresh cyanide requirements.

17.4.3 Consumables Requirements

The main consumables for the process plant include the grinding media and the liners for the SAG

and ball mills, as well as the reagents used in leaching, gold recovery and future CND circuit. It was

assumed that, for Phase 1, only one mill (i.e. SAG mill) is required to produce a P80 of 75 microns.

This assumption will be validated in the next stage of work for the Kiena Mine Complex project.

The grinding media consumption for the SAG and ball mill was estimated using benchmarking data

for similar projects and adjusted using power calculations (Bond equation or Moly-Cop Tools

software). The average media consumption for both grinding applications is presented in

Table 17-4.




JUNE 2020 17-12

Table 17-4: Estimated grinding media consumption

Area Type Size (mm)

Consumption (kg/t)

Phase 1 (years 2021 to 2023)

Phase 2 (year 2024 and beyond)

SAG Mill Forged steel 127 0.8 0.3

Ball Mill Forged steel 51 - 0.6

The average reagents consumptions are indicated in Table 17-5

Table 17-5: Reagents and consumables – Applications and consumption

Area Use

Consumption (tpy)

Phase 1 (years 2021 to

2023)

Phase 2 (year 2024 and

beyond)

Liners – SAG Mill Mills 10.0 5.0

Liners – Ball Mill (BM) Mills - 11.8

Grinding Media – SAG Mills 143.5 72.3

Grinding Media – BM secondary Mills - 150.7

Sodium Cyanide Gold lixiviant and gold eluant 91.6 136.8

Quicklime pH modifier 161.0 240.4

Flocculant Flocculation of solids in thickeners 3.5 5.3

Activated Carbon Gold adsorption 3.5 5.3

Anti-scalant Scale control 4.1 6.2

Caustic Soda Carbon stripping/washing 19.2 28.6

HCl Carbon wash 16.7 24.9

Lead Nitrate Gold leaching catalyst 35.5 53.0

Quicklime pH modifier in CND - 240.0

SO2 Liquid CND agent - 28.8

Copper Sulphate CND reaction catalyst - 27.4




JUNE 2020 18-1

PROJECT INFRASTRUCTURE

18.1 Introduction

The Kiena Mine Complex produced gold up until June 2013 and has been under care &

maintenance since then. The complex was suspended with the belief that operations would

resume eventually. Wesdome has maintained a site presence since 2013 and is currently

developing the underground ramp while using the existing infrastructure. It is therefore expected

that most of the surface infrastructure can be reused as is to support the planned production rate

as the future operations will typically be at a rate below historical operating rates. An exception is

the power distribution system, which will need refurbishing prior to restarting operations.

Also, a new filtered tailings facility is planned to be built in 2023 and to be operational in 2024. It

will replace the current tailings facility that is expected to reach full capacity around that

timeframe. This new tailings management facility (TMF) will include a filter plant, water treatment

plant, and stacking area. A new electrical substation may be required to power these new

facilities.

18.2 Site Access

The Kiena Mine Complex is located next to Highway 117, 10 km west of Val-d’Or, on the De

Montigny Lake. It can be accessed by trucks via the Kienawisik road by turning north from

Hwy 117. On site, various roads and parking lots are present. The nearest airport with scheduled

flights from Montreal is Val-d’Or Airport.

18.3 Service and Administration Buildings

Existing service and administration buildings include:

▪ Administrative building;

▪ Warehouses and maintenance shops;

▪ Dry house and engineering office;

▪ Assay laboratory;

▪ Reagents and hazardous material storage facility.

No new administration or service buildings are envisioned for the future operation of the site, with

the possible exception of a mine rescue station.

18.4 Personnel and Accommodation

As the complex is located within the limits of the town of Val-d’Or, there are no requirements for

camps or lodging.




JUNE 2020 18-2

18.5 Power and Electrical

Power is supplied to the Kiena Mine Complex through the Hydro-Québec transmission network

via a 25 kV line. On the Kiena property, the 25 kV line splits off to several substations where it is

stepped down to supply power to the various consumers on surface and underground.

A 1,000 kW diesel-powered generator provides emergency power to critical areas within the

Kiena Mine Complex in the event of a major electrical disruption.

The mine consumed about 7 MW at peak when it was last operating in 2013. The current forecast

for power is approximately the same; however, power availability will need to be discussed with

Hydro-Québec. In addition, a detailed power study will be required during the prefeasibility study

to further define the load distribution at site. It is expected that a small substation can be located

next to the new tailings facility that can be connected directly to the regional grid if the future

power study shows a lack of capacity at the existing main substation. This has been included as a

risk item.

All electrical equipment will have to be inspected and tested before the restart. Many components

have been identified as needing refurbishments and/or modernization during BBA’s audit (Boily

and Gangadin, 2019).

18.6 Communications

Site communications infrastructure includes:

▪ Internet access (both satellite and hardwired from a local provider);

▪ A fiber optic control network using the Modbus Plus technology;

▪ An emergency communication line for the mine;

▪ Three separate CCTV networks, one for the site security, one for the gold room, and one for

process monitoring.

18.7 Cyanide Destruction

The cyanide destruction (CND) system will be refurbished and operative in year 2024. The

system consists in an Inco SO2/Air. The thickener will treat the underflow of the tailings thickener

(at 60% solids) and the overflow of the backfill cyclones. Process water will be added to maintain

50% solids in the CND feed. The tails from the CND system will be split between the backfill

cyclones and the filter plant in a 40/60 percentage ratio. The equipment will include two tanks

(90 min retention time), agitators, reagent preparation and storage. A new testwork program will

be conducted to validate the process design criteria and equipment requirements.




JUNE 2020 18-3

18.8 Filter Plant

A new filter plant will be built and operational by 2024 as part of the tailings management system

upgrades. Figure 18-2 shows the proposed location for the filter plant. The feed to the filter plant

at approximately 50% solids will be processed by a pre-filter thickener that will increase the

percentage solids to 60%. The underflow of the thickener will be fed to a stock tank that will act as

a buffer for the filter press circuit (two in operation and one stand-by). The dry product at 85%

solids will be trucked to the TSF and the filtrate will be pumped back to the pre-filter to remove

suspended solids from the water. The overflow of the pre-filter thickener will feed to the thickener

O/F tank and will be recirculated back to the process water reservoir or the fresh water tank.

18.9 Existing Tailings Storage Facility

The Kiena mine contains an existing tailings storage facility (TSF) that is located on the western

part of the site (see Figure 18-1). The TSF consists of two main storage cells that have a total

remaining design capacity of 1.78 Mm³ (357,400 m³ in the North Cell and 1,424,860 m³ in the

South Cell). The current water management strategy (tailings and process water) considers the

utilization of the two cells as well as a polishing pond (Coppola and Gagnon, 2019).

Based on the current mine and backfill plan, deposition can continue in the existing TSF for up to

50 months. A new facility is therefore required to store tailings that will be generated by the future

operations, based on the identified constraint that the South facility cannot be used for additional

storage. However, in order to be conservative and allow for risk mitigation at this early stage of

design, the new TSF has been scheduled to come online in 2024, roughly 30 months after start-

up. An opportunity exists to eliminate the new facility or to delay its construction if the South Cell

can be used without significant compromises and should be investigated in future studies, once

the mine plan is further refined.




JUNE 2020 18-4

Figure 18-1: Kiena mine site - Existing surface layout

18.10 Additional Tailings Management Facility

The limits of the new tailings management facility (TMF) were previously defined by Stantec

(Coppola and Gagnon, 2019) and presented to the Ministère de l’Environnement et de la Lutte

contre les changements climatiques (MELCC) in the report entitled "Avis au MELCC concernant

l’emplacement de la nouvelle aire d’accumulation au Complexe Kiena afin de réaliser une

codéposition des stériles et des résidus miniers". The new TMF considers that tailings will be

stored as a dry stack in an area that is located nearby the existing waste rock stockpile (see

Figure 18-2:). The existing waste rock is considered as non acid generating (Coppola and

Gagnon, 2019).

BBA proposes a design that uses the existing waste rock to build a perimeter berm that will

confine the tailings dry stack. This structure will be developed prior to the commencement of

tailings deposition.




JUNE 2020 18-5

Figure 18-2: Proposed tailings management facility layout

18.10.1 Design Considerations

The conceptual design performed by BBA considers the following:

▪ The new facility will become operational 30 months after the restart of operations. The

tailings for the first 30 months will be deposited in the current facility (client provided design

parameter). This assumption has not been validated by BBA;

▪ Since physical properties of the filtered tailings are currently under evaluation, some

assumptions were made by BBA for the geotechnical analysis of the facility;

▪ Expected filtered tailings moisture content is around 15%;

▪ Future tailings are assumed to be non-acid generating material (see Section 20.2.3);

▪ It is assumed that the protection of groundwater by means of a geomembrane on the

footprint of the future facility will not be required. This is because cyanide destruction is

expected to occur before tailings are sent to the new TSF and filtered tailings have low water

content. This assumption must be validated in the next phase of the Project;




JUNE 2020 18-6

▪ BBA has assumed that future tailings will have a density ranging from 1.6 t/m³ to 1.8 t/m³.

For volumetric calculations, as well as for conceptual stability analysis, a compacted in-place

density of about 1.65 t/m³ has been used;

▪ The new facility is expected to store 1.0 Mt (0.61 Mm³), which corresponds to six years of

anticipated production (see Table 18-1). Tailings dry-stacking operation is planned to begin

approximately two and a half years after the mine operations resume. Facility capacity has

been evaluated at conceptual level and will need to be validated in further stages of the

project. Throughout the TSF operations, dust control activities will need to be carried out.

Table 18-1: Surface tailings production – TSF required capacity

Year

Tailings to be stored

(tonnes)

Tailings to be stored

(m³)

Dry stack cumulative tailings

(m³)

Total cumulative tailings

(m³)

2021 46,340 28,154 0 28,154

2022 147,064 89,348 0 117,502

2023 173,318 105,299(1) 52,650(1) 222,801

2024 182,311 110,762 163,412 333,564

2025 180,363 109,579 272,991 443,143

2026 174,519 106,029 379,019 549,171

2027 159,126 96,676 475,696 645,848

2028 152,986 92,946 568,642 738,793

2029 63,353 38,490 607,132 777,284

Note: (1) Half of the 2023 produced tailings are assumed to be stored at the new dry stack. Dry-

stacking tailings operation will commence 30 months after start-up.

▪ The tailings slurry that will be produced by the process plant will be transported by pipelines

(see Figure 18-2:) to a filter plant that will be located, as proposed by BBA, north of the

existing polishing pond. From the filter plant, tailings will be trucked to the TSF and will be

disposed of by lifts of 300 mm using bulldozers and compactors.

18.10.2 Facility Configuration

BBA has performed a conceptual design of the proposed TSF. The approximate area that will be

occupied by the new TSF and its perimeter berm and drainage structures will be 11.1 ha (see

Figure 18-3 and Figure 18-4). Given the current location of the core shack, this facility will need to

be relocated. BBA has proposed a preliminary relocation that will have to be validated in the next

phase of the Project. Refer to Appendix D for the layout of the proposed tailings management

system and for details of the dry stack area concept. The main characteristics of the facility are

listed as follows:




JUNE 2020 18-7

▪ The final crest of the TSF is at elevation 319 m a.s.l. (metres above sea level). The expected

maximum height is 21 m;

▪ The filtered tailings stack is expected to be built following a global slope of 5H:1V, above

berm elevation;

▪ The perimeter berm will be built out of the existing waste rock, which is assumed to be a self-

draining material. The crest width of the berm is 4 m, with an inner slope of 2.5H:1V, an

outer slope of 3H:1V and an average height of 5.3 m. The approximative berm volume is

100,000 m³, whereas the existing waste rock pile volume is 175,000 m³. This means that

there will be approximately 75,000 m³ of waste rock that could be used as a bed material for

new filtered tailings or for other project needs;

▪ Surface water is proposed to be managed by two perimeter ditches, both with a total length

of approximatively 1.4 km. the water will be conveyed by gravity to the polishing pond and

then disposed of into the environment via the final effluent point, or pumped to existing

tailings ponds if quality does not meet the environmental design criteria;

▪ BBA’s preliminary evaluation shows that the facility’s raising rate will be between 1 m to 3 m

per year. A detailed stacking sequence and planning should be developed in the next phase

of the Project. Zones designated for winter and summer deposition must be evaluated.

Frozen tailings zone should be avoided within the stack;

▪ Tailings compaction is assumed to be no less than 95% of maximum dry density as

determined by the Modified Proctor Test. In order to better asses the dry stack construction

parameters, it is recommended that a complete tailings characterization and testing be

performed in the next phase of the Project;

▪ Long term stability has been validated by BBA; however, more analyses and modelling are

required and should be performed in the next phase of the Project. BBA recommends to:

- Gather additional geotechnical data to refine the stability model;

- Consider pore pressure dissipation between deposition sequences, for short term slope

stability analyses. It is important to note that the stress-strain-pore water pressure

model needs to be consistent with the final deposition plan of the TSF;

- Conduct a detailed seepage analysis to define the position of the phreatic surface within

the TSF;

- Conduct liquefaction analysis for the loose and saturated silt beneath the clay layer;

- Analyze different options to improve clay foundation strength (pre-load, wick drains,

other).

▪ A tailings line will be required to convey water from the process plant to the filter plant. It is

assumed that this will be installed on the existing road corridor.




JUNE 2020 18-8

Figure 18-3: TSF configuration and typical cross-section




JUNE 2020 18-9

Figure 18-4: Berm location and typical cross-section

18.11 Water Management

A preliminary approach to manage impacted surface water from the TSF was developed by BBA.

The following design considerations were assumed:

▪ The new facility will impact surface water quality on the existing footprint. According to

Directive 19, contact water from the new TSF must be managed and treated prior to disposal

into the environment;

▪ Water filtered from the tailings will be returned to the process plant;

▪ The principal contaminant expected on surface water from the TSF will be total suspended

solids (TSS). This assumption must be validated in the next phase of the Project;

▪ A surface gravity drainage system is to be preferred;




JUNE 2020 18-10

▪ Preliminary design flood criteria have been defined in accordance with Directive 019. The

expected water volumes should be calculated considering a 24-h rainfall with a recurrence of

1,000 years combined with a 30-day snowmelt with a recurrence of 100 years;

▪ Precipitation data for the design events is shown in Table 18-2. This information is based on

BBA's experience from previous projects executed in the Val-d’Or area and should be

validated at the PFS level study.

Table 18-2: Project design precipitation data

Precipitation Recurrence Value

24-h rainfall 1: 1,000 year 93.7 mm

30-day snowmelt 1: 100 year 521.0 mm

BBA has advanced a conceptual design of the water management infrastructure. Detailed design

must be performed in the next phase of the Project. The resulting water management

infrastructure is described below (Figure 18-5 and Figure 18-6):

▪ A drainage network comprising two ditches, A and B, with respective lengths of 790 m and

560 m. Conceptual cross-section considers internal slopes of 3H:1V, a minimum base width

of 1.5 m and a height of 1.0 m. Ditches will be built in cut and fill;

▪ A prefabricated pumping station located at the end of the ditch network will be required; in

the event drainage water does not meet the discharge quality criteria, water will have to be

pumped to the existing tailings pond.

Water volumes resulting from the design flood over the TSF area has been evaluated by BBA at

68,550 m³. The impact of managing this additional amount of water with the existing infrastructure

must be evaluated in the next phase of the Project.




JUNE 2020 18-11

Figure 18-5: TSF water management infrastructure

Figure 18-6: Drainage ditch conceptual cross-section

18.12 Specific Infrastructure Upgrades for Restart

Several upgrades are required for the restart of the Project due to the difference in mineralized

material location and depth. These upgrades are covered in Chapter 16 Mining Methods

(Dewatering, UG distribution, etc.) and in Chapter 17 Recovery Methods (Reagents).




JUNE 2020 19-1

MARKET STUDIES AND CONTRACTS

19.1 Market Studies

No market studies have been conducted by Wesdome Gold Mines or its consultants in relation to

the gold doré that will be produced by the Kiena Mine Complex project. Gold is a freely traded

commodity on the world market for which there is a steady demand from numerous buyers.

19.2 Assumptions

Table 19-1 outlines the terms used in the economic analysis.

Table 19-1: NSR assumptions used in the economic analysis

Assumptions Unit Value

Au payable % 99.95

Au refining charge (including transportation cost) CAD/oz 3.00

Gold price USD/oz 1,532.00

The long-term price of gold was taken as the average price projection for the period 2020-2029 and

is expressed in 2019 dollars. Independent gold price projections were prepared by CPM Group for

Wesdome Gold Mines and are dated March 30, 2020 (refer to Appendix E). Figure 19-1 shows the

historical and projected gold price (nominal and real in 2018 US dollars).

Figure 19-1: CPM’s gold price projections to 2029




JUNE 2020 19-2

19.3 Contracts

Wesdome Gold Mines Ltd. currently has a gold refining agreement in place with the Royal Canadian

Mint in Ottawa. In addition, several contracts have been awarded to local businesses. Table 19-2

lists the contracts that are already in place at the Kiena Mine Complex.

Table 19-2: Existing contracts at the Kiena Mine Complex

Supplier Category 2020 estimated value ($ million)

CMAC-Thyssen Labour Supply 13

Forage Orbit Drilling 7

Laboratoire ALS Services 1.5

AC Nord Transport 0.2

Services MNG Labour Supply 0.05

Hydro-Quebec Electrical 1.6

Energir Services 0.1

Garda Services 0.23




JUNE 2020 20-1

ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

It is planned to restart production at the Kiena Mine Complex in 2021. It is envisioned that during

the first three years of operation (namely Phase 1), the process plant will operate "as-is", at a

maximum throughput of 829 tpd until the new tailings management facility (TMF) is operative. In

year 2024 and beyond (Phase 2), the throughput will increase (maximum 869 tpd), and tailings will

be stored at the new TMF.

Information relevant to the status of existing permits has been provided by Wesdome.

20.1 Environmental Studies and Issues

20.1.1 Environmental Studies

Various environmental studies have been produced to obtain the existing Certificate of

authorization (CoA) dated August 2, 1984. The following baseline studies must be carried out

before the establishment of the new TMF:

▪ Soil quality;

▪ Hydrogeology and groundwater quality;

▪ Hydrology and surface water quality;

▪ Fish habitats;

▪ Wetlands;

▪ Vegetation.

20.1.2 Environmental Issues

The Project has been under care and maintenance since mid-2013. Effluent is being controlled and

analytical results are submitted to federal and provincial authorities who have determined that the

effluent, based on the available data, complies with regulation requirements.

Most of the Project's infrastructure was permitted under the 1984 Certificate of Authorization, which

allows for the mining and milling of 2,200 tpd. However, the projected new TMF and the relocation

of the waste rock storage dumps must be authorized by new Certificates of Authorization under the

2012 version of the Directive 019 on mining industry.

The main challenge for the implementation of the new TMF is the lack of available surface space

due to the proximity of the lake and the CN railroad. Social resistance from some stakeholders is a

possibility.




JUNE 2020 20-2

20.2 Waste Rock, Mineralized Material, Tailings and Water Management Strategy

20.2.1 Waste Rock Management

Geochemical Characterization

Geochemical characterizations of waste rock samples were carried out in 2001 and 2012 by

Stavibel, and in 2016 by Stantec. A summary including all the results was presented in a report by

Stantec (2017). Results showed that the waste rock is not potentially acid generating.

According to Directive 019 on mining industry, a mining waste is considered «leachable» if, for a

given contaminant, the concentration measured in the leachate obtained with the toxicity

characteristic leaching procedure (TCLP) exceeds the criteria for resurgence in surface water and

the content (mg/kg) is higher than soil quality Criteria A. For the 2012 study, results from the TCLP

static leaching test indicated that some of the five samples were «leachable» for chromium, copper

or nickel. However, metal concentrations in leachates were lower than the resurgence criteria in

leachates obtained with the synthetic precipitation leaching procedure (SPLP) and CTEU-9 more

representative leaching tests.

For the 2016 study, leachates from the TCLP static leaching test showed manganese

concentrations higher than the resurgence criteria for all 13 samples. However, manganese

contents of the samples were not measured. Therefore, it is not possible to clearly classify the

samples as «leachable» for manganese. It should be noted that manganese contents measured in

all five samples from the 2012 study were lower than Criteria A. For the 2016 study, detection limits

for various metals (chromium, copper, lead, nickel, zinc) were higher than the corresponding

resurgence criteria and therefore those results are useless.

Consequently, it is likely that waste rock should not leach contaminants. However, complementary

geochemical testing should be carried out in order to confirm that waste rock is not leachable.

Management

The Project involves the mining of new deposit zones. As a result, the waste rock in the new zones

could be slightly different from that mined in previous phases of the Kiena project.

All waste rock produced will be used for backfilling or stored in underground workings. Moreover,

following complementary ARD and ML testing results, waste rock could probably be stored in

surface areas near the existing TMF and/or the projected dry-stack.




JUNE 2020 20-3

20.2.2 Mineralized Material Management


As the mineralized material and tailings are similar for gold mines, no geochemical characterization

of the mineralized material has been carried out.

Management

Mineralized material will be stored in a silo before being milled and will not be stored at the surface

of the site. However, a temporary mineralized material storage area of 300 m² will be used.

20.2.3 Tailings Management


Geochemical characterizations of waste rock samples were carried out in 1999, 2000, 2001, 2002,

2012 and 2013. Results showed that tailings were not potentially acid generating. Results from the

TCLP static leaching test also showed that some tailings' samples were leachable as per Directive

019 classification for various metals (chromium, copper, manganese, nickel and lead). However,

metal concentrations in leachates were lower than the resurgence criteria in leachates obtained

with the SPLP and CTEU-9 more representative leaching tests. Therefore, leaching of metals from

tailings management facilities is unlikely.

The Project includes the mining of new deposit zones. Therefore, it is possible that mineralized

material from the new zones could be different from that mined in previous phases of the Kiena

project. In this context, a geochemical characterization study must be carried out on representative

mineralized material samples or, ideally on representative tailings samples produced during

metallurgical studies.

Management

Approximately 38% of the tailings produced at the process plant will be sent in underground

workings for backfilling purposes. For the remaining tailings, the technology selected for the PEA

is the dry stack method. A comprehensive site selection and technology selection study will be

carried out at the PFS level. For the PEA, a site located east from the existing tailings storage area

has been considered for the location of the dry stack.

20.2.4 Water Management

Mine

Dewatering water from underground workings will be sent to the process plant. This will include

water from backfill leaching, underground operations and wall infiltration.

Cyanide destruction will be carried out on tailings used for backfilling and on tailings sent to the

new TMF.




JUNE 2020 20-4

Process Plant

Most of the water required for the process plant will be reclaimed from the tailings filtration

installations at the mill and from dewatering water. Freshwater for potable and sanitary uses, as

well as for reagents and gland seal purposes, will be pumped from a well located on the site.

Tailings Management Facilities

Water from the tailings filtration installations will be recirculated to the process plant or used in

backfill. Water from upstream of the dry stack will be diverted by a system of ditches and berms.

Precipitation falling on the dry stack (including snow) will be pumped to the existing TMF for control

of suspended solids.

In order to respect Directive 019 requirements, cyanide destruction will be carried out at the process

plant. Since tailings are not expected to generate acidity or metal leaching, runoff waters from the

dry stack should not contain contaminants other than suspended solids. Effluent from the polishing

pond will be monitored and discharged in the environment.

20.3 Legal Aspects and Permitting

20.3.1 Federal

Canadian Environmental Assessment Act (CEAA)

An environmental and social assessment is required by the federal government for projects covered

under the Impact Assessment Act, 2019 (IAA). The IAA applies to projects described in the Physical

Activities Regulations (SOR/2019-285).

(19) The expansion of an existing mine, mill, quarry or sand or gravel pit in one of the

following circumstances:

(c) in the case of an existing metal mine, other than a rare earth element mine, placer

mine or uranium mine, if the expansion would result in an increase in the area of

mining operations of 50% or more and the total ore production capacity would be

5 000 t/day or more after the expansion;

The planned production rate of the Kiena project will be approximately 800 tpd to 900 tpd, therefore

the Project is not subject to the IAA Environmental Assessment Process, even if the mining

operations area was to be increased by 50% or more.

Fisheries Act

The Fisheries Act (R.S.C., 1985, c. F-14) amended by An Act to Amend the Fisheries Act and other

Acts in consequence (2019) and the Metal and Diamond Mining Effluent Regulations (SOR/2002-

222) apply to mining projects. Environment Canada Guidelines for the Assessment of Alternatives

for Mine Waste Disposal (2016) must be used for tailings location selection study.




JUNE 2020 20-5

Others

Various other Acts and Regulations could apply to mining projects:

▪ Canadian Environmental Protection Act (S.C. 1999, c. 33):

- PCB Regulations (SOR/2008-273);

- Environmental Emergency Regulations (SOR/2003-307);

- Federal Halocarbon Regulations (SOR/2003-289); and

- National Pollutant Release Inventory.

▪ Species at Risk Act (S.C. 2002, c. 29);

▪ Canada Wildlife Act (R.S.C., 1985, c. W-9):

- Wildlife Area Regulations (C.R.C., c. 1609).

▪ Migratory Birds Convention Act, 1994 (S.C. 1994);

▪ Nuclear Safety and Control Act (S.C. 1997, c. 9):

- Radiation Devices Regulations (SOR/2000-207).

▪ Hazardous Products Act (R.S.C., 1985, c. H-3);

▪ Explosives Act (R.S.C., 1985, c. E-17);

▪ Transportation of Dangerous Goods Act (1992):

- Transportation of Dangerous Goods Regulations (SOR/2001-286).

20.3.2 Provincial

Ministry of Environment and Fight against Climate Change

Environmental Impact Assessment and Review

The Project is located in southern Quebec and is therefore subject to Division II, § 4 of the

Environmental Quality Act.

According to Section 22 of Part 2 of the Regulations Respecting Environmental Impact Assessment

and Review (Q-2, r.23.1), mining projects subject to the environmental impact assessment and

review include:

(2) the establishment of a mine whose maximum daily capacity for extracting any other

metal ore is equal to or greater than 2,000 metric tons;

(5) any increase of the maximum daily extraction capacity of a mine referred to in

subparagraph 2 or 3 to reach or exceed, as the case may be, a threshold provided for

therein;

(6) any expansion of 50% or more of the mine operation area in the following cases:

(b) the maximum daily extraction capacity of a mine referred to in subparagraph 2 or

3, as the case may be, is reached or exceeded




JUNE 2020 20-6

The planned production rate of the Kiena Mine Complex project will be approximately 800 tpd to

900 tpd and therefore the Project is not subject to the Environmental Impact Assessment and

Review Process even if the mine operation area was to be increased by 50% or more.

However, Certificates of Authorization under Section 22 of the Environmental Quality Act will be

required for the new tailings storage area and for the relocation of the waste rock storage dumps.

Modifications to the existing CoA issued in 1984 will also be required for various topics.

Applicable Acts and Regulations

Applicable Acts and Regulations from various ministries includes:

▪ Environmental Quality Act (c. Q-2):

- Regulation Respecting the Application of Section 32 of the Environment Quality Act

(Q-2, r. 2);

- Regulation Respecting the Application of the Environment Quality Act (Q-2, r. 3);

- Clean Air Regulation (Q-2, r. 4.1);

- Regulation Respecting Industrial Depollution attestations (Q-2, r. 5);

- Regulation Respecting Pits and Quarries (Q-2, r. 7);

- Regulation Respecting the Declaration of Water Withdrawals (Q-2, r. 14);

- Regulation respecting Mandatory Reporting of Certain Emissions of Contaminants into

the Atmosphere (Q-2, r. 15);

- Regulation Respecting Halocarbons (Q-2, r. 29);

- Regulation Respecting Hazardous Materials (Q-2, r. 32);

- Protection Policy for Lakeshores, Riverbanks, Littoral Zones and Floodplains (Q-2,

r. 35);

- Water Withdrawal and Protection Regulation (Q-2, r. 35.2);

- Land Protection and Rehabilitation Regulation (Q-2, r. 37);

- Regulation Respecting the Charges Payable for the Use of Water (Q-2, r. 42.1).

▪ Directive 019 sur l’industrie minière (2012);

▪ Protection and Rehabilitation of Contaminated Sites Policy (1998);

▪ Threatened or Vulnerable Species Act (c. E-12.01):

- Regulation Respecting Threatened or Vulnerable Wildlife Species and their Habitats (E-

12.01, r.2);

- Regulation Respecting Threatened or Vulnerable Plant Species and their Habitats (E-

12.01, r.3).

▪ Compensation Measures for the Carrying out of Projects Affecting Wetlands or Bodies of

Water Act (M-11.4);




JUNE 2020 20-7

▪ Watercourses Act (c. R-13):

- Regulation Respecting the Water Property in the Domain of the State (R-13, r. 1).

▪ Sustainable Forest Development Act (c. A-18.1):

- Regulation Respecting Standards of Forest Management for Forests in the Domain of

the State (A-18.1, r. 7).

▪ Conservation and Development of Wildlife Act (c. C-61.1):

- Regulation Respecting Wildlife Habitats (C-61.1, r. 18).

▪ Lands in the Domain of the State Act (c. T-8.1);

▪ Act Respecting the Preservation of Agricultural Land and Agricultural Activities (P-41.1):

- Preservation of Agricultural Land and Agriculture Activities Regulation (P-14-1, r. 1).

▪ Building Act (c. B-1.1):

- Safety Code (B-1.1, r. 3);

- Construction Code (B-1.1, r. 2).

▪ Explosives Act (c. E-22):

- Regulation under the Act Respecting Explosives (E-22, r. 1).

▪ Cultural Heritage Act (c. P-9.002);

▪ Occupational Health and Safety Act (c. S-2.1):

- Regulation Respecting Occupational Health and Safety in Mines (S-2.1, r. 14).

▪ Highway Safety Code (c. C-24.2):

- Transportation of Dangerous Substances Regulation (C-24.2, r. 43).

Ministry of Energy and Natural Resources

The Mining Act (c. M-13.1) and the Regulation Respecting Mineral Substances other than

Petroleum, Natural Gas and Brine (M-13.1, r. 2) contain requirements for mines' development,

operation and closure.

Commission des normes, de l'équité, de la santé et de la sécurité du travail

As per the Regulation Respecting Occupational Health and Safety in Mines (S-2.1, r. 14), where

tailings are used for backfilling, the water contained in such residues and leaking therefrom may

not have a potassium cyanide (KCN) content higher than 50 mg/L.

Commission de protection du territoire agricole du Québec

All the Project infrastructure facilities are located outside of agricultural zones as defined by the

Commission de protection du territoire agricole du Québec (CPTAQ).




JUNE 2020 20-8

20.3.3 City of Val-d’Or

The Kiena property is located inside the limit of the city of Val-d’Or. If required, construction permits

must be obtained from the city.

20.4 Social and Communities Issues

Wesdome has kept the municipal authorities informed about their mining and exploration activities

but no official consultation has been conducted with other interested stakeholders. Noise has been

the main concern in the past. Another concern was groundwater quality as it is the source of potable

water for nearby dwellings. Wesdome monitored the well water quality up until operations ceased

in 2013. Resuming use of cyanide would trigger the resumption of the well water monitoring

program.

20.5 Closure and Rehabilitation

20.5.1 Concepts

Rehabilitation works will include buildings and infrastructure dismantling, site safety, ground ripping

and revegetation of impacted area such as the infrastructure’s footprints and some roads.

Rehabilitation works will also include revegetation of tailings management facilities and waste rock

piles.

Dismantling Buildings and Infrastructure

Buildings and infrastructure specifically erected for the operation of the mine will be dismantled to

retrofit the site to a state compatible with the surrounding environment. Some infrastructure could

be maintained for the benefit of the local communities. The cost estimation considers that all

buildings will be dismantled.

The following buildings/infrastructure have been considered:

▪ Process plant;

▪ Conveyors and silo for material handling;

▪ Tailings management facilities and corresponding water management infrastructure;

▪ Heavy equipment;

▪ Office;

▪ Maintenance garage;

▪ Pumps;

▪ Gasoline and diesel fuel tanks;

▪ Fencing and lighting towers;

▪ Electric station and electric transport line.




JUNE 2020 20-9

During the dismantling operations and disposal of the Project's facilities, all buildings and surface

infrastructure not required for the closure plan follow-up process will be taken apart by a certified

contractor. Waste material resulting from the dismantling operations will be transported to

authorized recycling sites. During the dismantling operations of the buildings and infrastructure,

rehabilitation work will include the following activities:

▪ Salvageable material and equipment will be set aside and then donated or sold to recycling

sites;

▪ Any process, production or service equipment, such as silos, reservoirs, tanks, pipelines and

pumps will be drained and cleaned. The wash water will be collected for treatment (settling,

water/oil separation if needed) before being discharged into the environment;

▪ Any equipment containing oils or other potentially contaminating liquids, such as electrical

equipment and vehicles, will be drained and cleaned before being discarded;

▪ Management of chemical products, waste materials, and dangerous goods will be disposed

of safely according to regulations in effect. All solids, liquids, pulps and sludges located

inside the buildings will be characterized, if needed, and their disposal sites will be approved

by the Project environment representative;

▪ The walls and floors of the buildings will be cleaned, if needed, before the buildings are

dismantled. The wash water will be collected for treatment (settling, water/oil separation if

needed) before being discharged into the environment.

Rehabilitation of Impacted Areas

All impacted areas such as roads, laydown areas and industrial work bays, as well as the various

footprint areas of dismantled buildings, will be scarified to improve drainage and revegetation

results. Ripping will be carried out with a scraper box in order to rework the ground surface on a

thickness of 150 mm. An overburden and organic soil layer will then be placed on the restored

ground before seeding.

Dry Stack Tailings

Revegetation will be carried out by hydro-seeding. TMF will be operated by successive cells and

progressive rehabilitation/revegetation will be carried out.

Temporary Mineralized Material Stockpile

The surface of the stockpile area will be ripped, and a 15 cm layer of overburden and organic soil

will be placed on the surface before seeding.

Water Management

A breach will be realized in the polishing pond, and, to the fullest possible extent, the original

drainage will be restored.




JUNE 2020 20-10

Site Safety

At closure, blocks of waste rock will be used for plugging the entry of the underground mine.

Concrete will be used for plugging all openings to the surface (ventilation or backfill purposes).

Heavy Mobile and Stationary Surface Equipment

Whenever possible, heavy mobile and stationary surface equipment will be sold on the used

equipment market. The remaining unwanted equipment will be sold as scrap metal or disposed of

at designated dump sites. Heavy mobile and stationary surface equipment located in underground

workings will be hauled outside, drained of any liquids, and tagged as either saleable or scrap.

Excessively worn or old parts will be sent to scrap metal recyclers or disposed of at designated

dump sites.

New and Used Controlled Products

Petroleum products, fuels, diesel, oils and greases will be spent at the end of the LOM. All

petroleum products reservoirs and associated piping used on site to store will be drained, cleaned

and dismantled. Soils contiguous to the reservoirs or containers will be characterized and corrective

measures will be taken in compliance with the policy on the protection of soils and the rehabilitation

of contaminated lands.

All reagents and other chemical products will be spent at the end of the LOM, except those required

for water treatment during the environmental post-closure follow-up period. Residual reagents and

chemical products not required for that purpose will be put into properly labelled containers and

transported to approved sites for recycling.

Management of residual dangerous goods is regulated, and the disposal of such products must be

done in compliance with the Regulations on Dangerous Goods. No residual hazardous materials

should be found on the Property after the cessation of the mining operations. All used oils will be

sent to an approved recycling/burning site and other residual dangerous goods will be collected,

packaged, labelled and transported at approved sites for elimination.

Residual non-dangerous materials will be sorted; recyclable materials will be sent to an authorized

recycling facility.




JUNE 2020 20-11

Soils and Contaminated Materials

At cessation of mining activities, the property will be characterized and rehabilitated if the

characterization study reveals presence of contamination. Incidents associated with handling of

petroleum products or other chemical products could occur, especially at the following sites:

▪ Petroleum products storage facility;

▪ Point of use locations of petroleum products;

▪ Reagents and chemical products storage facility;

▪ Near plants and mechanical shops;

▪ On the path between the treatment plant and the TMF.

All soils affected by petroleum hydrocarbons shall be excavated and disposed of at an authorized

site.

20.5.2 Cost Estimates

The total cost of reclamation (and the guarantee) is estimated at $10.0M. This cost includes direct

and indirect costs of site rehabilitation as well as post-closure monitoring, engineering costs and

the mandatory 15% contingency.

A closure and restoration plan for the existing project has been submitted to the MERN and

approved on June 19, 2017. The total cost of reclamation was estimated at $7.0M. Wesdome has

fed a financial guarantee fund for this amount.

The additional reclamation costs not covered by the actual financial fund is estimated at $3.0M.

This amount is to be provided in three installments constituting of 50%, 25% and 25% of the

additional restoration costs. The first payment is to be provided within 90 days of receiving the

approval of the next version of the restoration plan. The second and third installments (25%) are

due on the anniversary date of the restoration plan approval.




JUNE 2020 21-1

CAPITAL AND OPERATING COSTS

The capital and operating cost estimates presented in this PEA for the Kiena Mine Complex project

are based on the restart of operations at the mine after a period of inactivity, and comprise various

refurbishments and upgrades, as well as the addition of a new tailings storage facility.

All capital and operating cost estimates cited in this Report are referenced in Canadian dollars.

21.1 Capital Costs

21.1.1 Summary

The total pre-production capital cost for the Kiena Mine Complex project is estimated to be $43.7M

(including indirect costs, contingencies, pre-production operating costs, and sunk costs). The

cumulative life of mine capital expenditure including costs for pre-production and sustaining is

estimated to be $164.5M.

Table 21-1 provides a summary of the Project capital costs.

Table 21-1: Project capital costs summary




($M)


3000 Mine 29.4 92.8 122.2


6000 Process Plant 2.4 1.2 3.6


0.02 16.65 16.7


9000 Project Indirect Costs (9800 excluded)

2.7 4.0 6.7

9800 Contingency 2.9 5.6 8.5


Total 43.7 120.8 164.5







JUNE 2020 21-2

21.1.2 Scope of Capital Cost Estimate

The following items are included in the capital cost estimate:

All refurbishments and restart costs that were identified as necessary in the plant restart

audit report (BBA document 3767002-000000-80-ATR-0001-R00) and during BBA’s site

visits.

Refurbishment of the headframe as recommended in the headframe structure inspection

report (Concept DB document 2342-RA-20190712).

Required refurbishments, upgrades, additions, and development in the mine as estimated by

BBA based on the mine plan, site visit, and discussions with Wesdome.

Construction of a new dry stack tailings facility and related infrastructure.

Project indirect costs and owner’s costs.

Reclamation and closure costs related to the new tailings management facilities (Wesdome

already has a bond with the government of Quebec that covers the closure of the existing

facilities).

Operating costs related to the pre-production phase (until the end of June 2021).

A contingency of 25%. This contingency was not applied to items whose scope is considered

well defined, such as mine development, mining equipment, and several mine infrastructure

items.

The following items were excluded from the capital cost estimate:

Licensing and financing costs.

Project development costs incurred to date, including studies and early works.

Permitting costs.

Taxes.

Operating costs after start of production.

Work stoppages.

Scope changes or an accelerated schedule.

Hydrological, environmental or hazardous waste issues.

21.1.3 Structure and Basis of Capital Cost Estimate

The capital cost estimate was developed in accordance with BBA’s work breakdown structure

(WBS) presented in Table 21-2:




JUNE 2020 21-3

Table 21-2: BBA work breakdown structure (WBS)

WBS Area WBS Description

2000 Administration and Services

3000 Mine

5000 Stockpiling and Conveying

6000 Process Plant


8000 Owner's Costs (8900 excluded)

9000 Project Indirect Costs (9800 excluded)

- Pre-production Operating Costs

9800 Contingency

8900 Site Reclamation and Closure

The cost schedule estimated for the LOM was divided into two timeframes: pre-production and

sustaining. Pre-production capital includes capital costs incurred from 2020 to Q2 2021. Sustaining

capital includes capital costs incurred from Q3 2021 to 2029.

The overall capital cost estimate developed in this Preliminary Economic Assessment study meets

the AACE Class 4 requirements and has an accuracy range of between -35% and +35%. The

capital cost estimate for this study forms the basis for the approval of further development of the

Project by means of a prefeasibility study (PFS). The various sections have been developed at

different levels of engineering definition.

The subcomponents of the estimate are based on the following basis:

Well defined items, such as items identified in the audit report, the headframe inspection

report, and the new tailings facilities, have been estimated using a bottom-up approach.

Less defined items, such as mine infrastructure items, have been estimated by using

benchmarking.

Mine development costs are based on the preliminary mine plan and actual cost data

provided by Wesdome.

The estimated costs of some items, such as field programs and crusher refurbishment, have

been provided by Wesdome based on quotes or budget allocation.

Project indirect costs and owner’s costs were estimated on a percentage basis over the

associated direct costs.




JUNE 2020 21-4

For bottom-up estimates, the following parameters were used for labour costs:

a. A 5-day workweek and 10-hour workdays for a total of 50 hours per week;

b. Contracts are assumed to be awarded to local contractors on a cost-plus basis;

c. Regulated rates are used mostly for electrical, fire protection, civil and concrete work,

and are based on the industrial collective agreement 2017-2021, data published by

"Association de la construction du Québec" (ACQ), the "Commission de la construction

du Québec" (CCQ), the "Commission de la Santé et de la Sécurité du Québec" (CSST)

as well as the "Direction générale des acquisitions du Centre des services partagés du

Québec";

d. Unregulated rates are used mostly for mechanical work and are based on historical data

from similar projects;

e. Different productivity rates have been used depending on the type of work, based on

various factors.

Contingency: for this level of study a 30% contingency is typical; however, given the fact that

Wesdome is currently developing the mine, many costs are known, and the contingency was

reduced to 25%.

21.1.4 Mining Capital Costs

Refer to Table 21-3 for a breakdown of mining capital expenditures by year and area.




JUNE 2020 21-5

Table 21-3: Mining capital cost expenditures

Item Total Unit 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029

Lateral Development - Project 16,085 $k 8,891 7,194 0 0 0 0 0 0 0 0

Lateral Development - Sustaining 63,249 $k 0 7,103 17,747 16,431 15,508 6,447 0 14 0 0

Vertical Development - Project 0 $k 0 0 0 0 0 0 0 0 0 0

Vertical Development - Sustaining 5,664 $k 0 1,680 2,256 1,200 480 48 0 0 0 0

U/G Construction 5,796 $k 0 806 2,097 1,270 977 430 215 0 0 0

Infrastructure - Hoisting 1,942 $k 271 1,661 10 0 0 0 0 0 0 0

Infrastructure - Dewatering 3,060 $k 260 350 700 700 700 350 0 0 0 0

Infrastructure - Ventilation 500 $k 0 0 100 100 100 100 0 100 0 0

Infrastructure - Electrical 5,724 $k 375 1,604 1,070 1,605 1,070 0 0 0 0 0

U/G Crushing 725 $k 500 225 0 0 0 0 0 0 0 0

U/G Mobile Equipment 19,895 $k 0 9,133 7,482 2,825 0 0 0 0 0 0

Total Capital Expenditures 122,185 $k 10,297 29,756 31,462 24,131 18,835 7,375 215 114 0 0




JUNE 2020 21-6

21.1.5 Pre-production Capital Costs

The pre-production capital costs incurred from 2020 to the first half of 2021 from the Kiena Mine

Complex project are estimated to total $43.7M of project-related capital expenditures. The pre-

production capital costs include a contingency of 25%.

The Project pre-production capital cost summary is outlined in Table 21-4 and shown as a pie chart

in Figure 21-1.

The pre-production ‘Forecast to Spend’ amount of $36.3M (as of March 31, 2020) includes the

addition of reclamation and closure bonding costs, as well as the subtraction of sunk costs. The

pre-production mine development costs for the year 2020 are considered as sunk costs.

Table 21-4: Project pre-production capital cost summary


capital cost ($M) CAPEX (%)

2000 Administration and Services 0.5 1.2

3000 Mine 29.4 67.2

5000 Stockpiling and Conveying 0.1 0.2

6000 Process Plant 2.4 5.6

7000 Tailings Storage Facility & Water Management 0.02 0.1

8000 Owner's Costs 3.4 7.8

9000 Project Indirect Costs 2.7 6.1

9800 Contingency 2.9 6.7

Pre-production Operating Costs 2.2 5.1

Total 43.7 100.0

Less Sunk Costs -8.9 -

8900 Site Reclamation and Closure 1.5 -

Total - Forecast to Spend 36.3 -




JUNE 2020 21-7

Figure 21-1: Distribution of pre-production capital costs (%)

21.1.6 Sustaining Capital Costs

The sustaining capital costs incurred over the 8 years of production (Q3 2021 to Q2 2029) from the

Kiena Mine Complex project is estimated to total $120.8M of project-related capital expenditures,

excluding reclamation and closure bonding. The sustaining capital costs include a contingency of

25%.

The sustaining ‘Forecast to Spend’ amount of $122.3M includes the addition of reclamation and

closure bonding costs.

The breakdown of LOM sustaining capital expenditures by area is provided in Table 21-5 and

Figure 21-2, while a detailed sustaining capital schedule is provided in Table 21-6.

1.2

67.20.2

5.6

0.1

7.8

6.1

6.7 5.1

Administration and Services Mine

Stockpiling and Conveying Process Plant

Tailings Storage Facility & Water Management Owner's Costs

Project Indirect Costs Contingency

Pre-production Operating Costs




JUNE 2020 21-8

Table 21-5: Project sustaining capital cost summary

Area Cost area description Sustaining capital

cost ($M) CAPEX (%)

2000 Administration and Services 0.0 0.0

3000 Mine 92.8 76.8

5000 Stockpiling and Conveying 0.0 0.0

6000 Process Plant 1.2 1.0

7000 Tailings Storage Facility & Water Management 16.6 13.8

8000 Owner's Costs 0.6 0.5

9000 Project Indirect Costs 4.0 3.3

9800 Contingency 5.6 4.6

Total 120.8 100.0

8900 Site Reclamation and Closure 1.5 -

Total - Forecast to Spend 122.3 -

Figure 21-2 Project sustaining capital cost summary (%)

76.8

1.0

13.8

0.53.3 4.6

Mine Process Plant

Tailings Storage Facility & Water Management Owner's Costs

Project Indirect Costs Contingency




JUNE 2020 21-9

Table 21-6: Sustaining capital cost summary by year

Area Description 2021 2022 2023 2024 2025 2026 2027 2028 2029 Total %

3000 Mine 10.7 31.5 24.1 18.8 7.4 0.2 0.1 - - 92.8 75.9

6000 Process Plant - 0.3 0.9 - - - - - - 1.2 1.0

7000 Tailings Storage Facility & Water Management - - 16.6 - - - - - - 16.6 13.6

8000 Owner's Costs (8900 excluded) - - 0.6 - - - - - - 0.6 0.5

9000 Project Indirect Costs (9800 excluded) - 0.1 3.9 - - - - - - 4.0 3.3

9800 Contingency - 0.1 5.5 - - - - - - 5.6 4.6

Subtotal 10.7 32.0 51.6 18.8 7.4 0.2 0.1 - - 120.8 98.8

8900 Site Reclamation and Closure - 0.8 0.8 - - - - - - 1.5 1.2

Total 10.7 32.8 52.4 18.8 7.4 0.2 0.1 - - 122.3 100.0




JUNE 2020 21-10

21.2 Operating Costs

21.2.1 Summary

The average operating cost over the 8-year mine life is estimated to be $162.66/t mined or

$493.1/oz (CAD). Table 21-7 provides a summary of the projected operating costs for the Kiena

Mine Complex project.




($/t mined) Average

LOM ($/oz) OPEX (%)

U/G Mining 214.1 26.8 104.15 315.7 64.0

Processing & Lab 48.6 6.1 23.62 71.6 14.5



HSE & Training 11.5 1.4 5.61 17.0 3.4

Administration 24.8 3.1 12.07 36.6 7.4


5.6 0.7 2.75 8.3 1.7

Total 334.4 41.8 162.66 493.1 100.0

Table 21-8 provides the breakdown of operating costs per year for the life of the mine (from

Q3 2021 to Q2 2029).




JUNE 2020 21-11

Table 21-8: Operating costs breakdown per year for LOM

Cost area Unit 2021 2022 2023 2024 2025 2026 2027 2028 2029 Total

Overall - Mineralized Material tonnage

kt 35.1 172.5 295.0 310.3 307.0 297.1 270.9 260.4 107.8 2,056.1

U/G Mining M$ 3.7 18.0 30.7 32.3 32.0 30.9 28.2 27.1 11.2 214.1

Processing & Lab M$ 0.8 3.8 7.0 7.4 7.3 7.1 6.4 6.2 2.6 48.6

Surface Operations M$ 0.3 1.5 2.5 2.6 2.6 2.5 2.3 2.2 0.9 17.4

Technical Services M$ 0.2 1.0 1.8 1.9 1.8 1.8 1.6 1.6 0.6 12.3

HSE & Training M$ 0.2 1.0 1.7 1.7 1.7 1.7 1.5 1.5 0.6 11.5

Administration M$ 0.4 2.1 3.6 3.7 3.7 3.6 3.3 3.1 1.3 24.8

Tailings Management (new facilities)

M$ 0.0 0.0 0.9 0.9 0.9 0.9 0.8 0.8 0.3 5.6

Total Operating Costs M$ 5.6 27.3 48.1 50.6 50.1 48.5 44.2 42.5 17.6 334.4




JUNE 2020 21-12

21.2.2 Basis of Operating Cost Estimate

The operating costs for all cost areas, except for tailings management and related process costs

(both linked to the new tailings facilities) are based on detailed historical data from the Kiena Mine

Complex provided by Wesdome for the years 2011 and 2012, which are the last two full calendar

years of operation before production stopped. Historical costs were escalated based on inflation

rates observed in the industry over this period to obtain the current estimated operating costs. The

resulting numbers are in line with comparable mining operations.

All operating cost estimates are in constant Q4 2019 Canadian dollars (CAD or $).

21.2.3 Mining

The historical operating costs provided by Wesdome in the mining category include costs related

to mining operations and underground services.

Mining operations' costs comprise level development, raise boring, mucking, ground support,

backfill, underground construction, sump maintenance, heating, utilities, supply services, lighting,

ventilation, road and level maintenance, waste disposal, supervision, crushing, hoisting, cage

services, shaft maintenance, underground rehab, and miscellaneous costs.

Underground services comprise underground transportation, trucking, service vehicles, supply

truck, scissor lift, bolters, jumbos, drills, jacklegs, shotcrete sprayer, rail loaders, tramming, fans,

pumps, work shops, maintenance supervision, rock breaker, crusher, and electrical and mechanical

services.

The average mining operating cost for the years 2011 and 2012 was $84.25/t. It has been escalated

with a rate of 2.5% per year to obtain a 2019 cost of $100.15/t.

As definition drilling was not included in the historical costs, an allowance of $4.00/t was added to

the escalated historical costs to account for it. The total operating costs for the U/G Mining cost

area is thus $104.15/t.

21.2.4 Processing and Laboratory

The historical operating costs provided by Wesdome in the processing and lab category include

costs related to the mill operations as well as the assay lab.

Mill operating costs comprise labour costs, supervision, temporary workforce, conveyor transport,

mineralized material storage, grinding, leaching and thickening, residue rejection, backfill, pump

stations, carbon absorption and desorption, acid washing, carbon regeneration, refining, security

guard and security systems, ingot transportation, mill laboratory, office costs, process control, trials,

mill utilities, compressors, workshops, structural maintenance, and environment services.




JUNE 2020 21-13

Costs related to the assay lab comprise workforce and supplies, supervision, maintenance, HVAC,

quality control, and analysis revenues.

The average processing and lab operating costs for the years 2011 and 2012 were $17.91/t. This

number has been escalated with a rate of 3% per year to obtain a 2019 cost of $22.02/t.

When the planned tailings management facilities come into operation in 2023, additional operating

expenses will be incurred due to the new cyanide destruction circuit. These additional processing

costs will mostly result from the consumption of reagents. Based on the projected reagent

consumption, it is estimated that the cyanide destruction will add $1.77/t to the processing costs.

This amount is added to the escalated historical cost of $22.02/t for the Processing & Lab cost area

from Q1 2023 onwards.

Table 21-9 provides the breakdown of costs for the reagents related to the destruction of cyanide.

Table 21-9: Cyanide destruction reagents cost

Reagents Average $/year $/t Mined

Quick Lime 46,780 0.26

SO2 Liquid 210,041 1.18

Copper Sulphate 57,523 0.32

Total 314,344 1.77

21.2.5 Surface Operations

The historical operating costs provided by Wesdome in the surface operations category include the

costs of technical services such as generators’ and electric compressors’ maintenance, mobile

surface equipment, wheel loaders, transport vehicles, road and yard maintenance, maintenance

supervision, surface structures, surface workshop, utilities, and powerline maintenance.

The average surface operating cost for the years 2011 and 2012 was $6.88/t. This number has

been escalated with a rate of 3% per year to obtain a 2019 cost of $8.47/t.

21.2.6 Technical Services

The historical operating costs provided by Wesdome in the technical services category include

engineering, geology, surveying, and environmental services costs.

The average technical services cost for the years 2011 and 2012 was $4.88/t. This number has

been escalated with a rate of 3% per year to obtain a 2019 cost of $6.00/t.




JUNE 2020 21-14

21.2.7 HSE & Training

The historical operating costs provided by Wesdome in the HSE and training category comprise

the underground and surface radio, mine rescue, industrial safety services, first aid and medical

tests, prevention and training services, health and safety committee, courses and workshops,

conferences, safety rewards, safety materials, customized training, major events, and emergency

measures.

The average surface HSE and training cost for the years 2011 and 2012 was $4.56/t. This number

has been escalated with a rate of 3% per year to obtain a 2019 cost of $5.61/t.

21.2.8 Administration

The historical operating costs provided by Wesdome in the administration category include costs

related to administration, human resources (HR), and corporate activities.

Administration costs comprise management costs, accounting, warehousing, office furniture,

telephone, internet, information technology, refining costs, mobile equipment insurance, general

insurance, permits, property taxes, subscription fees, membership and association fees, travel

expenses, and meals and entertainment fees.

Human resources costs comprise HR service, recruitment fees, employee relations, public

relations, retirement committee, donations and sponsorships, and collective agreement.

Corporate costs include long-term debt interest, interest and penalties, bank fees, and financing

fees.

The average administration operating cost for the years 2011 and 2012 was $9.82/t. This number

has been escalated with a rate of 3% per year to obtain a 2019 cost of $12.07/t.

21.2.9 Tailings Management

The new tailings facilities are projected to come online at the beginning of 2023. Until then, the

mine's tailings will be directed to the existing tailings pond. Operating costs related to the

management of tailings with the existing infrastructure are captured in the historical costs provided

by Wesdome under ‘Surface Operations’.

From Q1 2023, the operation of the new tailing management facilities will generate additional costs.

The operating costs for the new tailing facilities will include the costs to operate the filter plant,

stacking area, and water treatment plant. These costs have been estimated based on a study from

the Mine Environment Neutral Drainage (MEND) Program entitled ‘Study of Tailings Management

Technologies’, from 2017. According to this study, the typical relative operating cost per tonne of

tailings stored for dry stack installations is $5.20/t. Considering that that 41% of the tailings are

used for backfilling, we obtain an operating cost of $3.06 per tonne mined. This value is in line with

detailed estimates that have been developed for similar projects.




JUNE 2020 21-15

21.3 Site Personnel Summary – All Areas

The project is divided in two phases:

▪ In Phase 1, the mine will operate with the existing tailings storage facility (from the restart of

production in 2021 until 2023);

▪ In Phase 2, the mine will operate with a new dry stack tailings storage facility and with the

cyanide destruction circuit (2024 and beyond).

A total facility workforce of 147 employees is estimated for the Kiena Mine Complex project until

the end of Phase 1; at that point an estimated additional five employees will be required for Phase

2. A summary of labour in all areas is shown in Table 21-10.

Table 21-10: Summary of personnel – All areas

Employee count

Category Phase 1

(2021-2023)

Phase 2

2024-2029

Total Exploration 1 1

Mine Ops Total 70 70

UG Service Total 20 20

Mill Total 15 17

Assay Lab Total 5 5

Surface Ops Total 11 14

Technical Services Total 11 11

HSE & Training Total 6 6

Administration Total 7 7

HR Total 1 1

Total 147 152




JUNE 2020 22-1

ECONOMIC ANALYSIS

The economic/financial assessment of the Kiena Mine Complex project for Wesdome Gold Mines

was carried out using a discounted cash flow approach on a pre-tax and after-tax basis, based on

independent long-term projections for the gold price in United States currency and cost estimates

in Canadian currency. An exchange rate of USD 0.76 per CAD 1.00 was assumed to convert USD

market price projections and components of the capital cost estimates into Canadian dollars (CAD),

based on the three-year trailing average as of Q1 2020. No provision was made for the effects of

inflation.

Future tax liabilities were estimated by using a simplified tax model provided by Wesdome. This

model uses current Canadian tax regulations to assess the corporate tax liabilities and the most

recent provincial regulations to assess the Quebec mining tax liabilities.

The internal rate of return (IRR) on total investment was calculated based on 100% equity financing,

even though Wesdome may decide in the future to finance part of the Project with debt. The net

present value (NPV) was calculated from the cash flow generated by the Project, based on a

discount rate of 5%. The simple payback period and payback period after the start of production,

which are based on the undiscounted annual cash flow of the Project, are also indicated as a

financial measure. Furthermore, a sensitivity analysis has been performed for the after-tax base

case to assess the impact of variations in the Project capital costs, operating costs, USD to CAD

exchange rate, and the price of gold.

The economic analysis presented in this section contains forward-looking information about the

mineral resource estimates, commodity prices, exchange rates, proposed mine production plan,

projected recovery rates, operating costs, construction costs and project schedule. The results of

the economic analysis are subject to several known and unknown risks, uncertainties and other

factors that may cause actual results to differ materially from those presented here. The reader is

cautioned that this PEA is preliminary in nature and includes the use of Inferred Mineral Resources

that are considered too speculative geologically to have the economic considerations applied to

them that would enable them to be categorized as Mineral Reserves and, as such, there is no

certainty that the PEA economics will be realized.

22.1 Assumptions and Basis

The economic analysis was performed using the following assumptions and basis:

▪ The deposits included are the resources located within the Kiena mine area that contain

enough value for development and mining costs, as presented in the production plan (see

Chapter 16);

▪ The Project Executive Schedule developed in Chapter 24, taking into consideration key

project milestones;

▪ Commercial production start-up is scheduled to begin in the third quarter (Q3) of 2021. The

first full year of production is therefore 2022. Operations are estimated to span a period of

approximately eight years;




JUNE 2020 22-2

▪ The base case gold price is USD 1,532/oz.;

▪ The long-term price of gold was taken as the average price projection for the period 2020-

2029 and is expressed in 2019 dollars. Independent gold price projections were prepared by

CPM Group for Wesdome Gold Mines and are dated March 30, 2020. No price inflation or

escalation factors were considered. It is understood that commodity prices can be volatile

and that there is the potential for deviation from the LOM forecasts;

▪ The United States to Canadian dollar exchange rate has been assumed to be

USD 0.76: CAD 1.00 over the life of mine (CAD:USD exchange rate of 1.31), corresponding

to the three-year trailing average as of Q1 2020;

▪ All cost estimates are in constant Q4 2019 Canadian dollar with no inflation or escalation

factors considered;

▪ All metal products are assumed sold in the same year they are produced;

▪ Project revenue is derived from the sale of gold into the international marketplace.

This financial analysis was performed on both a pre-tax basis and after-tax basis. The general

assumptions used for this financial model, LOM plan tonnage and grade estimates are summarized

in Table 22-1.

Table 22-1: Financial model parameters

Description Unit Value

Long Term Gold Price USD/oz 1,532

Exchange Rate USD:CAD 0.76

Discount Rate % 5

Mine Life year 8

Total Mined Mt 2.1

Gold Grade g/t 10.65

Process Overall Gold Recovery % 97.0

U/G Mining Operating Costs $/t mined 104.15

Processing & Lab Operating Costs (Phase 1, Years 0 to 3) $/t mined 22.02

Processing & Lab Operating Costs (Phase 2, Year 3+) $/t mined 23.80

Surface Operations Operating Costs $/t mined 8.47

Technical Services Operating Costs $/t mined 6.00

HSE & Training Operating Costs $/t mined 5.61

Administration Operating Costs $/t mined 12.07

Tailings Management Operating Costs (New Tailings Facility) $/t mined 3.06

Royalties % NSR 0.00

Pre-production Capital Costs $M 34.8

Sustaining Capital Costs $M 120.8

Reclamation and Closure Costs $M 3.0




JUNE 2020 22-3

22.2 Gold Production

Over the life of mine (LOM), a total of 687,449 oz of gold (payable) (average annual: 85,931 oz)

will be produced. Figure 22-1 provides a summary of the payable gold production by year.

Figure 22-1: Annual payable gold production (oz)

0

100,000

200,000

300,000

400,000

500,000

600,000

700,000

800,000

0

20,000

40,000

60,000

80,000

100,000

120,000

2021 2022 2023 2024 2025 2026 2027 2028 2029

Cum

ula

tive G

old

Pro

duction (

oz)

Gold

Pro

duction (

oz)

Gold Production (oz) Cumulative Gold Production (oz)




JUNE 2020 22-4

22.3 Capital and Sustaining Costs

All capital costs (pre-production, sustaining, reclamation and closure) for the Project have been

distributed against the development schedule to support the economic cash flow model.

Figure 22-2 presents the planned annual and cumulative LOM capital cost profile, excluding sunk

costs.

Figure 22-2: Overall Kiena Mine Complex project capital cost profile

22.4 Royalties

Over the life of the Project, based on information from Wesdome’s 2018 financial statement, a 0%

NSR royalty has been assumed for the Kiena Mine Complex project.

22.5 Taxation

The taxation calculations for the Project were made based on a simplified spreadsheet model that

was provided by Wesdome.

At the effective date of the Report, it has been determined that the Kiena Mine Complex project will

be subject to the following tax regime:

▪ The Canadian Corporate Income Tax system consists of the federal income tax (modelled at

a rate of 15%) and the provincial (Quebec) income tax (modelled at a rate of 11.5%);

▪ The Quebec Mining Tax was modelled using a statutory tax rate of 16.5%.

0

20

40

60

80

100

120

140

160

180

0

10

20

30

40

50

60

2020 2021 2022 2023 2024 2025 2026 2027 2028

Cum

ula

tive C

apital C

osts

(m

illio

n C

AD

)

Annual C

apital C

osts

(m

illio

n C

AD

)

Pre-production Capital Costs Sustaining Capital Costs

Reclamation and Closure Costs Cumulative Capital Costs




JUNE 2020 22-5

The tax calculations are underpinned by the following key assumptions:

▪ The calculations are based on the tax regime at the date of the PEA. Future changes in tax

laws could impact the calculations;

▪ A total of $73M in tax attributes has been considered for the calculation of income taxes

based on information provided by Wesdome;

▪ A total of $47M in tax attributes has been considered for the calculation of mining taxes

based on information provided by Wesdome;

▪ The Project is held 100% by a corporate entity and the after-tax analysis does not attempt to

reflect any future changes in corporate structure or property ownership;

▪ Assumes 100% equity financing and therefore does not consider interest and financing

expenses;

▪ Quebec Mining Tax is deductible for federal and provincial income tax purposes;

▪ Actual taxes payable will be affected by corporate activities, and current and future tax

benefits, with respect that these activities have not been considered.

The combined effect on the Project of the two levels of taxation, including the elements described

above, is an approximate cumulative effective tax rate of 33%, based on project earnings (earnings

before interest and tax (EBIT)). It is anticipated, based on the Project assumptions, that Wesdome

will pay approximately $186M in Canadian Corporate Income Tax and $102M in Quebec Mining

Tax over the life of the Project.

22.6 Financial Analysis Summary

A 5% discount rate was applied to the cash flow to derive the NPV for the Project on a pre-tax and

after-tax basis. Cash flows have been discounted to 2020 under the assumption that the Project

will start in Q1 2020 and that major project financing would be carried out at this time. The summary

of the financial evaluation for the base case of the Project is presented in Table 22-2.




JUNE 2020 22-6

Table 22-2: Financial analysis summary (pre-tax and after-tax)

Description Unit Base case

Pre

-ta

x

Net Present Value (0% disc) $M 883.3


Internal Rate of Return % 125.7

Payback Period (simple) year 3.1

Payback Period (after start of operations) year 1.6

Aft

er-

tax



Internal Rate of Return % 101.6

Payback Period (simple) year 3.2

Payback Period (after start of operations) year 1.7

The pre-tax base case financial model resulted in an IRR of 125.7% and an NPV of $620.4M with

a discount rate of 5%. The pre-tax payback period after the start of operations is 1.6 years. On an

after-tax basis, the base case financial model resulted in an IRR of 101.6% and an NPV of $416.1M

with a discount rate of 5%. The after-tax payback period after the start of operations is 1.7 years.

The summary of the Kiena Mine Complex project discounted cash flow financial model (pre-tax and

after-tax) is presented in Table 22-3.




JUNE 2020 22-7

Table 22-3: Kiena Mine Complex project financial model summary

Year -2 -1 1 2 3 4 5 6 7 8 9

Total Unit 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030

Production Summary

Total Tonnes Processed kt 0.0 49.2 172.5 295.0 310.3 307.0 297.1 270.9 260.4 107.8 0.0 2,070.2

Head Grade Au g/t 0.00 5.80 8.05 9.70 9.24 9.28 11.18 13.36 12.54 14.81 0.00 10.65

Payable Gold koz 0.0 8.9 43.3 89.2 89.4 88.8 103.5 112.8 101.8 49.8 0.0 687.4

Revenue

Gold Price USD 1,532.00 1,532.00 1,532.00 1,532.00 1,532.00 1,532.00 1,532.00 1,532.00 1,532.00 1,532.00 1,532.00

Exchange Rate USD:CAD 0.76 0.76 0.76 0.76 0.76 0.76 0.76 0.76 0.76 0.76 0.76

Gross Revenue $M 0.0 17.8 86.6 178.6 178.9 177.9 207.2 225.9 203.8 99.7 0.0 1,376.4

Operating Expenditures

Underground Mining $M 0.0 3.7 18.0 30.7 32.3 32.0 30.9 28.2 27.1 11.2 0.0 214.1

Processing & Laboratory $M 0.0 0.8 3.8 7.0 7.4 7.3 7.1 6.4 6.2 2.6 0.0 48.6

Surface Operations $M 0.0 0.3 1.5 2.5 2.6 2.6 2.5 2.3 2.2 0.9 0.0 17.4

Technical Services $M 0.0 0.2 1.0 1.8 1.9 1.8 1.8 1.6 1.6 0.6 0.0 12.3

HSE & Training $M 0.0 0.2 1.0 1.7 1.7 1.7 1.7 1.5 1.5 0.6 0.0 11.5

Administration $M 0.0 0.4 2.1 3.6 3.7 3.7 3.6 3.3 3.1 1.3 0.0 24.8

Tailings Management $M 0.0 0.0 0.0 0.9 0.9 0.9 0.9 0.8 0.8 0.3 0.0 5.6

Total Operating Costs $M 0.0 5.6 27.3 48.1 50.6 50.1 48.5 44.2 42.5 17.6 0.0 334.4

Royalty Payments $M 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Capital Expenditures

Pre-production $M 8.6 26.2 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 34.8

Sustaining $M 0.0 10.7 32.0 51.6 18.8 7.4 0.2 0.1 0.0 0.0 0.0 120.8

Reclamation and Closure $M 0.0 1.5 0.8 0.8 0.0 0.0 0.0 0.0 0.0 0.0 0.0 3.0

Total Capital Costs $M 8.6 38.4 32.8 52.4 18.8 7.4 0.2 0.1 0.0 0.0 0.0 158.7

Changes in Working Capital $M 0.0 0.1 0.4 0.9 -0.1 0.0 0.6 0.5 -0.4 -1.0 -1.2 0.0

Pre-tax Cash Flow

Pre-Tax Cash Flow $M -8.6 -26.3 26.1 77.2 109.6 120.4 157.9 181.0 161.7 83.1 1.2 883.3

Cumulative Pre-Tax Cash Flow $M -8.6 -34.9 -8.8 68.5 178.0 298.4 456.3 637.3 799.1 882.1 883.3

Taxes and Duties

Mining Tax $M 0.0 -0.2 2.8 11.7 11.8 12.4 16.7 19.8 17.8 8.9 0.0 101.7

Income Tax $M 0.0 0.0 0.0 20.1 21.1 22.9 32.0 38.6 34.8 16.8 0.0 186.3

Total Taxes and Duties $M 0.0 -0.2 2.8 31.8 32.9 35.3 48.6 58.4 52.6 25.7 0.0 288.0

After-tax Cash Flow

After-tax Cash Flow $M -8.6 -26.1 23.3 45.5 76.6 85.1 109.3 122.6 109.1 57.3 1.2 595.3

Cumulative After-tax Cash Flow $M -8.6 -34.6 -11.3 34.1 110.8 195.9 305.1 427.7 536.8 594.2 595.3




JUNE 2020 22-8

Figure 22-3 shows the cumulative cash flows for the Project projected for the LOM on a pre-tax and

after-tax basis.

Figure 22-3: Life of mine cash flow projection (cumulative, pre-tax and after-tax)

22.7 Production Costs

A summary of the Project’s production costs is provided in Table 22-4. All costs are in USD. Total

cash costs are calculated per ounce on a payable basis using the costs of mining, material

transport, processing, tailings, waste and water treatment, on-site G&A, refining and smelting, and

royalties.

The LOM operating cash cost per ounce is USD 374/oz Au. The LOM all-in sustaining cost (AISC)

per ounce is USD 512/oz Au derived from the total cash costs plus sustaining capital and closure

costs. The operating margin over the LOM has been estimated to be USD 1,158/oz Au based on a

gold price of USD 1,532/oz.

-200

0

200

400

600

800

1,000

Year 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029

Cum

ula

tive C

ash F

low

(m

illio

n C

AD

)

Pre-tax (Non-Discounted) After-tax (Non-Discounted)




JUNE 2020 22-9

Table 22-4: Production cost summary

Description Unit LOM

Metal Production

Gold Moz 0.688

Costs and Royalties

U/G Mining USD M 163.6

Processing & Lab USD M 37.1

Surface Operations USD M 13.3

Technical Services USD M 9.4

HSE & Training USD M 8.8

Administration USD M 19.0

Tailings Management USD M 4.3

Refining and Smelting USD M 1.6

Royalties USD M 0.0

Total Operating Cost (after Credit) USD M 257.1

AISC and Profit Margins (per oz payable)

Gold Price USD/oz 1,532

Cash Cost (Operating) USD/oz 374

Sustaining and Closure Costs USD M 94.6

Total Costs (Operating and Sustaining) USD M 351.7

All-in Sustaining Costs (AISC) USD/oz 512

Operating Margin USD/oz 1,158

22.8 Sensitivity Analysis

A financial sensitivity analysis was conducted on the base case after-tax cash flow NPV ($M) and

IRR of the Project, using the following variables: capital costs, operating costs, USD:CAD exchange

rate, price of gold and discount rate. The after-tax results for the Project IRR and NPV ($M) based

on the sensitivity analysis are summarized in Table 22-5 through Table 22-9.




JUNE 2020 22-10

Table 22-5: NPV sensitivity results (after-tax) for metal price and exchange rate variations

USD:CAD Gold Price (USD/ounce)

1,100 1,200 1,300 1,400 1,532 1,600 1,700 1,800 1,900

0.60 359.8 414.4 469.0 523.4 595.0 631.9 686.2 740.5 794.7

0.65 313.6 364.0 414.4 464.8 531.1 565.1 615.2 665.3 715.4

0.70 274.1 320.8 367.6 414.4 476.2 507.9 554.4 600.9 647.4

0.76 231.1 273.9 316.7 359.5 416.1 445.3 488.2 530.8 573.4

0.80 210.0 250.8 291.6 332.5 386.6 414.4 455.4 496.3 537.0

0.85 183.5 222.0 260.4 298.8 349.7 375.9 414.4 453.0 491.5

0.90 159.6 196.3 232.7 269.0 316.9 341.6 378.0 414.4 450.8

Table 22-6: NPV sensitivity results (after-tax) for capital (LOM) and operating costs variations

OPEX CAPEX

-30% -20% -10% 0% 10% 20% 30%

-30% 505.2 491.7 478.2 464.7 451.2 437.7 424.3

-20% 489.0 475.5 462.0 448.5 435.0 421.5 408.1

-10% 472.8 459.3 445.8 432.3 418.8 405.3 391.9

0% 456.6 443.1 429.6 416.1 402.6 389.1 375.7

10% 440.4 426.9 413.4 399.9 386.4 372.9 359.5

20% 424.2 410.7 397.2 383.7 370.2 356.8 343.3

30% 408.0 394.5 381.0 367.5 354.0 340.6 327.1

Table 22-7: IRR sensitivity results (after-tax) for metal price and exchange rate variations

USD:CAD Gold Price (USD/ounce)

1,100 1,200 1,300 1,400 1,532 1,600 1,700 1,800 1,900

0.60 87.9% 101.2% 114.8% 127.8% 144.5% 153.1% 165.9% 178.8% 191.8%

0.65 76.9% 88.9% 101.2% 113.7% 129.6% 137.5% 149.2% 160.9% 172.8%

0.70 67.8% 78.6% 89.8% 101.2% 116.6% 124.2% 135.0% 145.8% 156.7%

0.76 57.9% 67.7% 77.6% 87.9% 101.6% 108.8% 119.6% 129.5% 139.4%

0.80 53.1% 62.4% 71.8% 81.4% 94.4% 101.2% 111.4% 121.6% 131.0%

0.85 47.2% 55.8% 64.6% 73.5% 85.5% 91.8% 101.2% 110.8% 120.5%

0.90 41.5% 50.0% 58.3% 66.6% 77.7% 83.6% 92.3% 101.2% 110.2%




JUNE 2020 22-11

Table 22-8: IRR sensitivity results (after-tax) for capital (LOM) and operating costs variations

OPEX CAPEX

-30% -20% -10% 0% 10% 20% 30%

-30% 180.7% 153.4% 132.3% 115.7% 102.3% 91.2% 81.9%

-20% 173.4% 147.1% 126.9% 111.0% 98.0% 87.4% 78.5%

-10% 166.2% 141.0% 121.6% 106.3% 93.9% 83.7% 75.1%

0% 159.0% 134.9% 116.3% 101.6% 89.8% 80.0% 71.8%

10% 151.9% 128.8% 111.1% 97.0% 85.7% 76.3% 68.5%

20% 144.9% 122.9% 105.9% 92.5% 81.7% 72.7% 65.2%

30% 138.0% 117.0% 100.8% 88.1% 77.7% 69.2% 62.0%

Table 22-9: NPV sensitivity results (after-tax) for discount rate

Discount Rate

0% 3% 5% 8% 10%

NPV 595.3 479.0 416.1 339.0 296.9

The graphical representations of the financial sensitivity analysis are depicted below in Figure 22-4

for the Project’s NPV and Figure 22-5 for the Project’s IRR.

The sensitivity analysis reveals that the price of gold has the most significant influence on both the

NPV and IRR compared to the other parameters, based on the range of values evaluated. After the

price of gold, the NPV and IRR were most impacted by changes in the USD:CAD exchange rate

and then, to a lesser extent, by variations in operating costs and capital costs. It should be noted

that the economic viability of the Project will not be significantly negatively impacted by variations

in the capital or operating costs, within the margins of error associated with the PEA cost estimates.

Overall, the NPV and IRR of the Project are positive over the range of values used for the sensitivity

analysis when analyzed individually.




JUNE 2020 22-12

Figure 22-4: Sensitivity of the net present value (after-tax) to financial variables

Figure 22-5: Sensitivity of the internal rate of return (after-tax) to financial variables

0

100

200

300

400

500

600

700

-40% -30% -20% -10% 0% 10% 20% 30% 40%

Aft

er-

Tax N

PV

@ 5

% (

CA

D-m

illio

ns)



0%

20%

40%

60%

80%

100%

120%

140%

160%

180%

-40% -30% -20% -10% 0% 10% 20% 30% 40%

Aft

er-

Tax IR

R (

%)






JUNE 2020 23-1

ADJACENT PROPERTIES

The Kiena Mine Complex is located in the Val-d’Or mining camp as illustrated in Figure 23-1.

Several mining and junior exploration companies are active in the area of the Kiena Mine

Complex. Pierre-Luc Richard has not been able to verify the information presented below and the

information is not necessarily indicative of the mineralization on the Kiena Mine Complex Property

area (the subject of this Report). Some information was also taken from the previous technical

report of Beausoleil et al. (2019).

23.1 Canadian Malartic Property

In June 2014, Yamana Gold and Agnico Eagle Mines bought all the common shares from Osisko.

They created the Canadian Malartic Partnership. It is located approximately 6 km west of the

Project. As of December 2018, proven and probable mineral reserves are estimated at 2.78 Moz

Au (Agnico Eagle Mines – Canadian Malartic, 2019).

Pierre-Luc Richard, QP, has not been able to verify the information presented above and

the information is not necessarily indicative of the mineralization on the Kiena Mine

Complex Property area.

The Canadian Malartic Property lies at the southern margin of the eastern portion of the Archean

Abitibi volcanic belt, mainly within the Pontiac Group of metasedimentary rocks. The property

covers a 16-km-long section of the Larder Lake Cadillac Fault Zone (LLCFZ) and is underlain by

mafic and ultramafic metavolcanic rocks of the Piché Group cut by intrusions, as well as

metasediments of the Cadillac Group north of the fault zone.

The Canadian Malartic mine is a large-tonnage, low-grade Archean gold system, consisting of a

broad shell of disseminated gold-bearing pyrite mineralization hosted by porphyritic felsic to

intermediate intrusions and altered metasediments. The system is open to the west and to the

south at depth. Mineralization in the Canadian Malartic Extension (Barnat deposit) is largely along

the southern edge of the LLCFZ. The two deposits contain the bulk of the current reserves and

are part of one large pit. The Jeffrey and Gouldie deposits, a few hundred metres east and south

of the pit, respectively, contain some of the mineral resources (Agnico Eagle Mines – Canadian

Malartic, 2019).

23.2 Dubuisson JV Property

The Dubuisson Property is a joint venture between Probe Metals Inc. and Agnico Eagle Mines. It

is located in Dubuisson Township to the south of the Project and consists of 31 contiguous claims

covering a surface area of approximately 748 ha.




JUNE 2020 23-2

The joint venture was announced in July 2010. Under the terms of the agreement, Agnico Eagle

Mines acquired 51% of the rights, title and interest in the property for $100,000 in cash and by

issuing 15,000 shares of Agnico Eagle Mines for a total value of approximately $1 million.

As part of the agreement and following the exercise of the option, Agnico Eagle Mines may

acquire an additional interest, which would bring Agnico Eagle Mines share up to 70% and Probe

Metals to 30% interest in the Property, by completing a bankable feasibility study. Agnico Eagle

Mines will act as the operator for all exploration work carried out on the Property during the option

period (Probe Metals-Dubuisson, 2019).

23.3 Agnico Eagle Mines Ltd. Properties

Agnico Eagle Mines Ltd. holds several properties and parts of historical properties northwest and

south of the Project: Goldex, Joubi, Bigué, Callahan and School Mine.

The Goldex Property, adjacent to the southeast limit of the Project, has proven and probable

reserves of 18.925 Mt at 1.58 g/t Au for 962,000 oz Au (Agnico Eagle Mines-Goldex 2019). It also

encompasses the Quebec Explorers Gold Deposit. The property straddles a 5-km segment of the

prolific LLCFZ. The gold system locally exceeds 20 m in thickness and can be traced for more

than 800 m along strike. The mineralization is known to a depth of more than 1,500 m (Deep

Zone 3).




The Goldex Property is located in the Dubuisson Township of Quebec in the southern Abitibi

Greenstone Belt. The intermediate to mafic and ultramafic volcanic sequence that underlies the

property dips steeply to the northeast. It is intruded by a large tabular-shaped quartz-diorite body

known as the Goldex Granodiorite that also dips steeply northeast.

Goldex is a large, relatively low-grade body defined by the intensity of stockwork veins and gold

grades rather than by individual veins. Most of the gold occurs as microscopic particles

associated with pyrite, while the rest occurs as coarse native gold grains. There are several

zones of gold mineralization with isolated mineralized intercepts over mineable widths on the

property, and all of them, except the South Zone, are hosted by the Goldex Granodiorite. The M

and E Zones and the Deep 1 Zone contain gold-bearing quart-tourmaline- pyrite veins and

veinlets. The South Zone consists of quartz veins that have higher grades than those in the

primary mineralized zones at Goldex (Agnico Eagle Mines – Goldex, 2019).

Past producers on the Goldex property include the Joubi Mine and some exploration shafts

(Goldex Shaft 1 and the School Mine).




JUNE 2020 23-3

The Callahan Property, adjacent to the northwest limit of the Project, hosts the Callahan deposit

in the Dubuisson Formation. In 1987, Falconbridge prepared a resource estimate. The latest

major work was in 2010 when Kinross Gold Corp., as owner of the claims, conducted a diamond

drilling campaign of 10,722 m (Beauregard and Gaudreault, 2010).

23.4 Tarmac Project Property

Globex Mining Enterprise holds six claims in the middle of the Project. Those claims are located

west of the closed Siscoe Mine. The claims have an area of 94.1 ha.

23.5 Harricana River Mining (O3 Mining) Property

O3 Mining Inc. acquired Harricana River Mining Corporation as per press release August 23,

2019. O3 Mining Inc. holds 117 claims in the vicinity of the closed Sullivan Mine. Those claims

are located approximately 6 km east of the Project.

During the period between 1934 and 1968, the Sullivan mine produced a total of 1,134,342 oz of

gold and 293,857 oz of silver from 4,613,500 t of mineralized material grading an average 7.65 g/t

Au and 1.98 g/t Ag. The veins mined at the Sullivan mine are found within the narrow west end of

the Bourlamaque granodiorite batholith. The granodiorite is considerably altered, but chemically it

is similar to the quartz-albite facies in the Siscoe stock (Sauvé et al., 1993).

23.6 Knick Exploration Property

Knick Exploration Inc. holds seven claims located 4 km west of the Project covering an area of

184.2 ha.

23.7 Metanor Resources (Bonterra) Property

Metanor Resources Ltd. (now Bonterra Resources Inc.) holds 21 claims located approximately

6 km southeast of the Project. The claims contain some showings and deposits (Nouvelle Zone

Aurifère and Zone No. 5).

23.8 Marban Block Property

The Marban Block Property held 100% by NioGold Mining Corporation (now Osisko Mining Inc.)

is located about 15 km west of the town of Val-dOr. The property consists of 42 claims and three

mining concessions for a total surface area of 9.8 km2. The Marban Block has three past

producing mines: Marban, Norlartic and Kierens. Those deposits have NI 43-101 resource

estimates effective as at June 1, 2013 (Gustin and Ronning, 2013), totalling 6.5 Mt at 1.4 g/t Au

for 296,000 oz of gold in measured resources and 25.6 Mt at 1.5 g/t Au for 1,235,000 oz of gold in

indicated resources (Osisko Mining – Marban Block, 2019).




JUNE 2020 23-4




The following description of the deposits is mostly modified and summarized from Trudel and

Sauvé (1992), Gustin and Ronning (2013), and references therein.

During the period between 1961 and 1974, the Marban mine produced a total of 330,000 oz of

gold from 1,983,000 t of mineralized material grading an average 5.27 g/t Au (Ducharme et al.,

2009).

During the periods between 1959 and 1966 and between 1990 and 1992, the Norlartic mine

produced a total of 188,000 oz of gold from 1.435 Mt of mineralized material at an average of

4.07 g/t Au (Ducharme et al., 2009).

During the period between 1965 and 1966, the Norlartic North Zone produced a total of 11,000 oz

of gold from 81,000 t of mineralized material grading an average 4.35 g/t Au. The North-North

Zone is located 500 m northeast of the Norbenite Fault. It is a near-surface intrusive-hosted

deposit with mineralized quartz-tourmaline stockwork. Gold mineralization is confined to a

conformable quartz-albite-carbonate-pyrite alteration envelope with a quartz-tourmaline-

carbonate vein stockwork localized in the central to lower portions of a 60-m-wide granodiorite sill

(Ducharme et al., 2009).

During the periods between 1965 and 1966 and between 1988 and 1992, the Kierens Mine

produced a total of 52,000 oz of gold from 251,000 t of mineralized material from the Kierens

Zone with an average grade of 6.30 g/t Au (Ducharme et al., 2009).

Three styles of gold mineralization were identified within the Kierens Zone by Aur Resources

geologists: 1) high-grade single-vein structures; 2) sill stockworks; and 3) laminated veins in

recrystallized mafic volcanic rocks.

A mineral resource estimate for the Kierens Zone was prepared by Gustin and Ronning (2013) on

behalf of NioGold Mining. The estimate, as at June 1, 2013, established an indicated resource of

1.437 Mt at 2.19 g/t Au for a total of 101,000 oz of gold, and an inferred resource of 1.178 Mt at

1.73 g/t Au.







JUNE 2020 23-5

23.9 Siscoe East Property

The Siscoe East Property is located 3 km northwest of Val-d’Or between the former Siscoe and

Sullivan mines. It hosts the former Stabell mine (Shaft No.1) that produced 13,629 oz of gold at

an average grade of 0.21 oz/t and Shaft No.2 that produced 1,782 oz of gold at an average grade

of 0.27 oz/t (Khobi and Frechette, 1987). The property is held 50% by NioGold Mining

Corporation (now Osisko Mining Inc.) and 50% by Alexandria Minerals Corporation (now

O3 Mining).

The Siscoe East Property is situated within the Malartic Group of mafic to ultramafic volcanic

rocks north of the LLCFZ. The property is characterized by the contact between mafic to

ultramafic volcanic rocks of the Dubuisson Formation (Lower Malartic Group) and the multi-

phased granodiorite-diorite of the Bourlamaque Batholith. In the summer of 2008, Alexandria

Minerals entered into a joint venture agreement with NioGold Mining allowing NioGold Mining to

earn a 50% interest in the property by issuing 650,000 shares (issued) and completing $750,000

of exploration work on the property.




JUNE 2020 23-6

Figure 23-1: Kiena Mine Complex adjacent properties




JUNE 2020 24-1

OTHER RELEVANT DATA AND INFORMATION

24.1 Project Execution Plan

Development and construction activities will occur at the mine and process plant site, located at the

Kiena Complex on Parker Island and at the tailings management facility, located east of Kienawisik

road and adjacent to the De Montigny Lake. The Project is divided in two phases:

▪ In Phase 1, the mine will operate with the existing tailings storage facility (from the restart of

production in 2021 until the end of 2023);

▪ In Phase 2, the mine will operate with a new dry stack tailings storage facility and with the

cyanide destruction circuit (from 2024 to 2029).

Accordingly, the Project execution is constituted of the following two components:

1. The restart of operations at the Kiena Complex to resume production. This component involves

the following:

- Operational readiness definition and preparation of master plans;

- Refurbishing and upgrades of the mine infrastructure;

- Refurbishing and upgrades of the processing plant and surface infrastructure;

- Mine pre-development;

- Plant commissioning.

The building of new dry stack tailings management facilities (TMF) to supplement the

existing tailings management infrastructure when it reaches full capacity. This component will

include:

- Engineering;

- Obtaining certificates of authorization and permits;

- Construction of new tailings management facilities (filter plant, dry stack area, water

treatment, upgrading the cyanide destruction circuit).

The major next steps in terms of engineering will include a prefeasibility study that will encompass

both components of the Project; however, the restart activities will occur concurrently.

24.1.1 Project Organization

Management

Wesdome Gold Mines will assemble a team to manage the Project technical studies and

construction management.




JUNE 2020 24-2

The Wesdome technical and environmental groups will supervise the engineering studies, the

environmental assessment and the EPCM services. Specialized engineering firms will be selected

for each portion of the Project to assemble a strong integrated design and execution team. They

will be responsible for the following functions in the respective phases:

▪ PFS:

- Preliminary technical specification and scope of work documents;

- Purchase package preparation for long lead equipment;

- 3D modelling, drawings production and material take-offs (MTO);

- Cost estimating of direct cost components.

▪ Execution:

- Definite specification and scope of work documents;

- Technical and economical evaluations;

- Short list meetings;

- Purchase order requisition preparation;

- Drawing management and approval.

The Wesdome technical team is responsible for the following activities in the respective phases:

▪ PFS:

- Budgetary/firm bid request;

- Addenda;

- Reception of bids;

- Indirect cost estimate;

- Project execution plan.

▪ Execution:

- Definite bid request;

- Addenda;

- Bid reception;

- Final negotiation;

- Contract award;

- Construction management;

- Purchase order release;

- Progressive payment;

- Shop visits;

- Site logistics.




JUNE 2020 24-3

24.1.2 Construction Management

During the engineering phase of the Kiena Mine Complex project, construction management will

contribute to the project design with constructability reviews. In the project execution phase,

Wesdome will assemble a construction management team (CMT) and will have the support of third-

party construction and contract administration specialists under the supervision of the construction

manager. The CMT will include the following services:

▪ Constructability reviews; ▪ Reporting;

▪ Site supervision; ▪ Health, safety and environment (HSE);

▪ Project cost control; ▪ Contract administration;

▪ Scheduling; ▪ Construction progress measurement.

It is recognized that an effective health and safety program during the Project is a necessity. The

success of the Construction Safety Program is contingent upon its enforcement at all stages of the

Project, including design, construction planning, construction execution, and start-up and

commissioning.

The CMT will also follow Wesdome’s procedures and work methods to ensure the protection of the

environment. Furthermore, the CMT will work closely with each department of the operations group

to ensure proper installation and functional results. During the construction phase, personnel from

operations will be integrated into the construction team as coordinators and supervisors.

The Wesdome Gold Mines operations group will support the CMT for the following services during

the construction phase:

▪ Staff payroll;

▪ Accounting support;

▪ IT support;

▪ Site security;

▪ Public relations;

▪ Environmental and permitting;

▪ Medical and first aid;

▪ Site logistics.

24.2 Project Execution Schedule

The preliminary project execution plan is developed to a PEA level and therefore conceptual in

nature. The execution plan and schedule will be further developed and detailed during the next

phases of project development. The preliminary project execution schedule, developed in this PEA

and described herein, covers the period from the end of the PEA (Q2 2020) up to the completion

of the construction of the new tailings management facilities in the fourth quarter (Q4) of 2023.

The major project milestones for the Project activities are shown in Table 24-1.




JUNE 2020 24-4

Table 24-1: Key milestones (preliminary)

Activity Start date Completion date

Preliminary Economic Assessment (PEA) 2020-04

Prefeasibility Study (PFS) 2020-04 2021-01

Operational Readiness Definition and Master Plans 2020-04 2020-11

Process Plant Restart 2020-04 2021-01

Upgrades and Production Ramp-up 2020-12 2021-09

New Tailings Management Facilities – EPCM 2022-06 2023-12

The Project is expected to require a provincial environmental assessment (EA) prior to the start of

new construction. In addition, the Project will require several permits, approvals and authorizations

from provincial, federal and municipal agencies. The ongoing baseline data collection will feed the

EAC application. The reception of the EAC certificate and Project permits is expected to be

completed in the first quarter (Q1) of 2021.




JUNE 2020 25-1

INTERPRETATION AND CONCLUSIONS

25.1 Overview

BBA was mandated by Wesdome to prepare a preliminary economic assessment conforming to

NI 43-101 standards to demonstrate the economic viability of the Kiena Mine Complex project. The

Project is based on the October 2019 mineral resources estimate prepared by Wesdome and

audited by BBA.

This NI 43-101 compliant technical report on the Kiena Mine Complex project was prepared by

experienced and competent independent consultants using accepted geologic and engineering

methodologies and standards. It provides a summary of the results and findings from each major

area of investigation including exploration, geological modelling, mineral resource, plant feed

estimations, mine design, metallurgy, process design, infrastructure, environmental management,

tailings and water management, capital and operating costs and economic analysis. The level of

investigation for each of these areas is considered to be consistent or surpassing what is normally

expected with a preliminary economic assessment for resource development projects.

The mutual conclusion of the QPs is that the Kiena Mine Complex project as summarized in this

PEA contains adequate detail and information to support the positive preliminary economic

outcome shown. The Kiena Mine Complex project contains substantial precious metal resources

that can be mined by underground methods and recovered using conventional processing

technologies. To date, the Qualified Persons are not aware of any fatal flaws in the Kiena Mine

Complex project and the results are considered sufficiently reliable to guide Wesdome

management in a decision to further advance the Project. This would typically involve the

preparation of a preliminary feasibility study or a feasibility study.

25.2 Data Verification and Mineral Resources

The October 2019 Kiena Mine Complex Mineral Resource Estimate (the “2019 MRE”) was

prepared by Karine Brousseau, P. Eng., Wesdome, and audited by Pierre-Luc Richard, P. Geo.,

BBA, using all available information including historical and recent diamond drillholes (DDH).

Pierre-Luc Richard, P. Geo., is the Independent and Qualified Person (QP) for the 2019 MRE and

all activities related to it. The effective date of the estimate is September 25, 2019 based on the

compilation status and cut-off grade parameters.

The resource database for the Project, as of August 6, 2019, consisted of 6,616 DDH (976,170.3 m)

with a cumulative length of 884,329.5 m, including 187 DDH from the 2018-2019 drilling program.

As of January 15, 2020, Wesdome had completed an additional 112 DDH for 32,413 m that are not

included in the herein MRE. Pierre-Luc Richard, QP, is of the opinion that while the addition of






JUNE 2020 25-2

Pierre-Luc Richard, QP, reviewed the drilling, sample preparation, analytical and security

procedures, as well as insertion rates and the performance of blanks, standards and duplicates for

the 2018-2019 drilling programs, and concluded that the observed failure rates are within expected

ranges and that no significant assay biases are present. Pierre-Luc Richard, QP, is of the opinion

that the protocols in place are adequate and have been followed. The database for the Project is

of good overall quality and adequate to industry standards. Pierre-Luc Richard, QP, is also of the

opinion that the database is appropriate for the purpose of the MRE and that the sample density

allows for a reliable estimate to be made of the size, tonnage and grade of the mineralization in

accordance with the level of confidence established by the Mineral Resource categories in the CIM

Standards.





Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Block Model MRE 968,900 14.46 450,400 1,121,200 11.02 397,100

Polygonal MRE 1,859,300 5.65 337,800 1,796,900 6.94 401,000

TOTAL 2,828,200 8.67 788,100 2,918,100 8.51 798,100

Notes to Table 25-1:

The independent qualified person for the 2019 MRE, as defined by NI 43 101, is Pierre Luc- Richard,

P. Geo., of BBA. The effective date of the estimate is September 25, 2019.

These mineral resources are not mineral reserves as they do not have demonstrated economic viability.


Results are presented in situ and undiluted and considered to have reasonable prospects for economic

extraction, below 100 m crown pillar.

The estimation combined two estimation methods, ordinary kriging in the Kiena Mine Complex and

polygonal for other deposits on the property.

The Kiena Mine Complex resources encompasses for 20 zones with a minimum true thickness of 3.0 m

using the grade of the adjacent material when assayed or a value of zero when not assayed. High-grade

capping varying from 20 g/t to 200 g/t Au (when required) was applied to composited assay grades for

interpolation using an Ordinary Kriging interpolation method based on 1.0 m composite and block size

of 5 m x 5 m x 5 m, with bulk density values of 2.8 (g/cm3). In addition, a high grade limit or second

capping value was used for the second and third pass grade interpolation to restrict high grade impact

at greater distance from the drillhole intersect. Indicated resources are manually defined and encloses

areas where drill spacing is generally less than 25 m, blocks are informed by a minimum of three

drillholes, and reasonable geological and grade continuity is shown.




JUNE 2020 25-3

The zone outside the Kiena Mine Complex encompasses for eight zones with a minimum true thickness

of 1.5 m using a polygonal estimation method. Indicated resources were estimated from drillhole results

using the mid distance between drillholes or a maximum of 30 m, 12.5 m in some areas. The high-grade

capping was fixed at 34.28 g/t Au with a bulk density value of 2.8 (g/cm3).

The estimate is reported for a potential underground mining scenario at cut-off grades of 3.0 g/t Au (>40°

dip) and 4.0 g/t Au (<40° dip, Wesdome Zone). The cut-off grades were calculated using a gold price of

USD 1,300 per ounce, a CAD:USD exchange rate of 1.31 (CAD 1,700); mining cost $110/t (>40° dip);

$150/t (<40° dip); processing cost $35/t; G&A $15/t. The cut-off grades should be re-evaluated

considering future prevailing market conditions (metal prices, exchange rate, mining cost, etc.).

The number of metric tons and ounces were rounded to the nearest hundred and the metal contents are

presented in troy ounces (tonne x grade / 31.10348).




It is Pierre-Luc Richard’s, QP, opinion that the cut-off grades are relevant to the grade distribution

of this Project and that the mineralization exhibits sufficient continuity for economic extraction under

the cut-offs applied. It should be noted that the MRE presented herein has an effective date prior

to the new CIM Best Practice Guidelines (published November 29, 2019) and, therefore, was

produced in accordance to the previous Guidelines. That being said, Pierre-Luc Richard’s, QP, is

of the opinion that, had this MRE been prepared under the new Guidelines, it would not have had

a material impact on the MRE presented in this Report.

25.3 Mining Methods

The mineral resource utilized in the mining plan is the Block Model MRE. A comparison with the

Mining Shapes developed using Deswik Stope Optimizer is shown in Table 25-2 and illustrates that

when applying mining parameters (stope dip, mining width, stope length, etc.) no additional

resource has been included in the mining plan.

Table 25-2: Comparison of block model MRE and mining shapes


Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Tonnage (t)

Grade (g/t)

Ounces Au (oz)

Block Model MRE 968,900 14.46 450,400 1,121,200 11.02 397,100

Mining Shapes 1,168,100 11.51 432,200 1,214,400 9.54 372,600

Note: These mineral resources are not mineral reserves as it includes inferred resource that are considered

too speculative geologically to have the economic considerations applied to them that would enable them to

be categorized as Mineral Resource, and there is no certainty that the PEA will be realized.




JUNE 2020 25-4

Kiena Deep will be the major contributor to the production tonnes and gold ounces for the LOM

Plan at the Kiena Mine. At this time, no changes to the historical mining method has been proposed.

As the mine is going deeper, there is the potential for higher Geomechanical impact to mining along

with potential ventilation constraints (heating) and recommendations have been made on work

required for further studies (refer to Chapter 26).

Continued drilling of the Kiena Deep Zone is focused on increasing the resource base and support

to move a quantity of the inferred resource to indicated resource.

25.4 Metallurgy and Processing

25.4.1 Metallurgy

Two metallurgical testwork programs were conducted with mineralized material from the Kiena

deposit in 2018 (Kiena Deep A from zones A and A1, and S50 zones) and 2019 (Kiena Deep A

zones A, A1 and A2). The testwork was focused on measuring comminution properties and Au

recovery in relation to the Kiena Mine Complex flowsheet.

The comminution test results have shown that all samples are moderately soft. Furthermore, the

Bond Abrasion test indicates that the samples are in the range of slightly abrasive to moderately

abrasive. Preliminary power calculation using the LOM production schedule and the comminution

results indicate that the mills at Kiena Complex are capable of grinding material with those

comminution properties.

Gravity testwork was evaluated during the Au leaching testwork: GTL (gravity plus leaching of

gravity tails) and WOL (whole ore leach). It was assumed that the WOL results at a preliminary

stage adequately represents the current leaching plus adsorption circuit at the Kiena Mine Complex

processing. Overall, Au recoveries were in a similar range. The WOL tests resulted in high gold

recoveries independent of the laboratory. Gold recoveries varied between 95.7% and 99.7% for a

leaching time of 48 hours and P80 = 75 microns.

In (Knelson/Mozley) gravity tests, high gold recoveries of 61%, 67% and 59% were achieved for

samples A, A1 and A2 respectively. Installation of a gravity circuit to the existing process has the

potential to recover gold faster and reduce the amount of coarse gold particles reporting to the

leach circuit for improved process economics.

NaCN consumptions were lower than 0.6 kg/t, which is considered low. The amount of consumed

CaO was found to be slightly lower than 1 kg/t in each test for all samples, which is considered

average. The flocculant reagent SNF Flomin 913 VHM was found to be the most effective and

efficient flocculant among the flocculants tested. The dosage range of 9-12 g/t was found to be

relatively low.

More comminution and metallurgical testwork is expected in the next phase of the Project to validate

the previous observations.

The preliminary testwork results confirm that the Kiena Mine Complex flowsheet is adequate to

process the mineralized material from the mine.




JUNE 2020 25-5

25.4.2 Process Flowsheet

The Kiena Mine Complex processing plant is a conventional gold recovery process involving

cyanidation and conventional CIP. The principal process steps included crushing, grinding, leaching

by cyanidation, gold adsorption and desorption, electrolysis, Inco SO2/Air cyanide destruction

system (not operative during the last years of operation), melting and casting of doré bars.

Before the operation shut down in 2013, the process tails were split between the backfill and the

tailings storage facility (in a 30/70 ratio). The cyanide destruction circuit was not operative. Cyanide

was destroyed by sunlight UV or by dilution at the tailings pond.

A new tailings storage facility (TSF) at the Kiena Mine Complex will be available in year 2024. It is

envisioned that, during the first three years of operation (namely Phase 1), the process plant will

operate “as-is” at a maximum throughput of 829 tpd until the new TSF is operative. In year 2024

(Phase 2), the throughput will increase (maximum 869 tpd) based on the latest LOM and a new

tailings plant facility will be operative (see Section 18.10). As part of the dry stack initiative for

storing process plant tailings a filter plant will be installed in Phase 2. The cyanide destruction circuit

will be refurbished to provide cyanide free tails to the filtration and backfill plants.

The filter plant will include a pre-filter thickener that will increase the feed percentage solids to 60%.

The underflow of the thicker will be feed to a stock tank that will act as a buffer for the filter press

circuit (two in operation and one stand-by). The dry product at 85% solids will be trucked to the TSF

and the filtrate will be pumped back to the pre-filter to remove suspended solids from the water.

The overflow (O/F) of the pre-filter thickener will be fed to the thickener O/F tank and the water will

be recirculated back to the process water reservoir or to the fresh water tank.

Based on the testwork results, the addition of a gravity circuit does not offer a significant

improvement in the project economics. This observation should be confirmed in the next project

phase with additional variability testwork.

It should be confirmed in the next phase of work that during Phase 1 the Kiena Mine Complex can

operate only with the SAG mill and not the ball mill due to the lower tonnage. Simulations could be

required.

It should be confirmed in the next phase of work that the Kiena Mine Complex can operate during

Phase 1 without a cyanide destruction circuit.

Regarding process plant power demand, the connected load is estimated as 1.4 MW and 2.1 MW

for Phase 1 and 2, respectively. The yearly consumption is estimated as 9.4 GWh and 13.9 GWh

for Phase 1 and 2, respectively.




JUNE 2020 25-6

25.5 Environmental Studies

In accordance with provincial laws, a rehabilitation and restoration plan has to be prepared and

approved by the MERN. The objective of the mine rehabilitation and restoration plan is to ensure

that the site does not present unacceptable risks to the health and safety of people, and to return

the site to an acceptable condition, ensuring that the environment will eventually return to normal

and that future generations can use the site.

The total cost of reclamation is estimated at $10.0M. This cost includes direct and indirect costs of

site rehabilitation as well as post-closure monitoring, engineering costs and the mandatory 15%

contingency. An estimated $3.0M will have to be added to the existing financial guarantee of $7.0M

to cover the reclamation and closure costs associated with the planned modifications to the site.

25.6 Infrastructure

It is expected that most of the existing site infrastructure can be reused as is to support the planned

production at the Kiena Complex; however, several upgrades have been identified as being

required, including the refurbishment of part of the power distribution systems.

In addition, the existing tailings storage facility is expected to reach its capacity by 2023 and a new

facility is therefore required to store the tailings that will be generated by the future operations. This

new tailings management system will be of the dry stack type, and will include a filter plant, stacking

area, and water treatment plant. It will also require the refurbishment of the existing cyanide

destruction system and the relocation of the core shack.

25.7 Capital and Operating Costs

The total capital costs (pre-production and sustaining) for the Kiena Mine Complex project were

estimated at $164.5M. The pre-production costs were calculated at $43.7M, including a $2.9M

contingency. The sustaining costs were calculated at $120.8M including a $5.6M contingency. Site

reclamation and closure bonding costs were estimated to be $3.0M. Items such as taxes,

permitting, licensing, and financing costs are not included in the cost estimate.

The overall capital cost estimate developed in this study meets the AACE class 4 requirements and

has an accuracy range of –35% and +35%.

The project capital cost summary is outlined in Table 25-3.




JUNE 2020 25-7

Table 25-3: Project capital cost summary




($M)


3000 Mine 29.4 92.8 122.2


6000 Process Plant 2.4 1.2 3.6


0.02 16.65 16.7


9000 Project Indirect Costs (9800 excluded) 2.7 4.0 6.7

9800 Contingency 2.9 5.6 8.5


Total 43.7 120.8 164.5




The average operating cost over the 8-year mine life is estimated to be $162.66/tonne mined. The

operating cost estimate developed in this study is deemed to be of an accuracy within alignment

with a PEA level of study. Table 25-4 presents the breakdown of the projected per-tonne mined

operating costs for the Kiena Mine Complex project.




($/t mined) Average

LOM ($/oz) OPEX (%)

U/G Mining 214.1 26.8 104.15 315.7 64.0

Processing & Lab 48.6 6.1 23.62 71.6 14.5



HSE & Training 11.5 1.4 5.61 17.0 3.4

Administration 24.8 3.1 12.07 36.6 7.4


5.6 0.7 2.75 8.3 1.7

Total 334.4 41.8 162.66 493.1 100.0




JUNE 2020 25-8

25.8 Indicative Economic Results

The financial analysis performed as part of this PEA using the base case assumptions results in an

after-tax NPV at 5% of $416.1M and an internal rate of return of 102% (base case exchange rate

of 0.76 CAD for 1.00 USD). The cumulative cash flow for the Project (after-tax) is $595.3M and the

payback period after start of operations is 1.7 years over the planned mine life of 8 years.

A sensitivity analysis indicates that the Project valuation would be most impacted by variations in

the gold price and exchange rate; however, the Project NPV and IRR remain positive over the range

of values evaluated, when analyzed individually. The economic viability of the Project will not be

significantly impacted by variations in the capital or operating costs.

The PEA plant feed is partly based on Inferred Mineral Resources that are considered too

speculative geologically to have the economic considerations applied to them that would enable

them to be categorized as Mineral Reserves, and there is no certainty that the preliminary economic

assessment based on these Mineral Resources will be realized.

25.9 Project Risks and Opportunities

As with most mining projects, there are risks that could negatively or positively affect the economic

viability of the Project. Many of these risks are based on a lack of detailed knowledge and can be

managed as more sampling, testing, design, and engineering are conducted at the next study

stages. Table 25-5 identifies what are currently deemed to be the most significant project risks and

opportunities.

Risks have not been included in the calculation of the contingency and are speculative at this level

of project development.




JUNE 2020 25-9

Table 25-5: Project risks and opportunities

Description and Potential Impact Group Probability Impact

Geology and Mineral Resources

There is some uncertainty in the spatial location of drillholes sampling data for longer historical drillhole for which there is not much downhole survey data available.

Economic Medium Low

The mineralized zones might be of slightly variable shapes due to the complex geometry of the deposit. Definition drilling will help define with more precision the shapes of the zones.

Economic Medium Low

The interpreted mineralized zones could be affected by some structures (faults or folds) that could displace or stop the mineralized zones. Definition drilling will improve the confidence in the interpretation.

Economic Low Medium

Increased resource associated with the up-dip potential of the Kiena Deep Zone could result in earlier than planned resource from this structure as it will be within access from existing infrastructure.

Economic Medium Medium

The calculation of the mineable resource was based on a gold price of USD1,300 per oz, Higher gold price results in current resource below cut-off becoming economical increasing resource tonnes.


Underground Mine

Cooling required in deep areas of mine. As is evident with other operations in the Abitibi area, cooling may be required, based on diesel equipment usage, re-use of air, and seasonal fluctuations. Going forward a heat study with placed diesel equipment will be required to confirm these requirements for cooling.

Economic Medium High

Bulkhead failure caused by crown pillar failure (bulkheads are being installed on levels with access to old workings). Bulkhead failure poses a safety risk to personnel in the mine.

Safety Low High

Bulkhead failure caused by crown pillar failure (bulkheads are being installed on levels with access to old workings). Bulkhead failure poses a safety risk to the general public. In the case of such a failure, boaters and fauna in the vicinity of the crown pillar may be drawn into the mine.

Community Low High

Faults of new deposit are more seismically active than anticipated (yet to be studied, PFS modelling is planned)

Safety Low High

Hanging wall failure in a stope increases dilution for duration of stope. Yet to be studied, PFS modelling is planned.

Economic Low Medium

Existing mine electrical circuits provide insufficient electrical capacity for the anticipated operation.

Economic Low Medium

Underground crusher cannot be refurbished with the capital allocated, and a new underground crusher installation is required.


Capital allocated for grizzly refurbishment is insufficient. Economic Low Low

Existing material in bins above crusher cannot be removed, and substantial or full rebuild of the bin is required.

Economic Low Medium




JUNE 2020 25-10


Loading pocket steelwork requires refurbishment. Economic Medium Medium

Mine development schedule not achieved. Production delayed by 4 months.


Late delivery of equipment, production schedule delayed. Economic Low Medium

Underground backfill system is inadequate, not usable, or fails (e.g. piping system failure, piping or equipment plugging, etc.).

Economic High High

Main ramp at level accesses is not flattened, an elevation differential on the downward side of the level access drift is created, and a small vehicle tips.

Safety Low Medium

One of the main fans is a Sheldon Fan and date of manufacture is unknown. This company is now defunct and spare parts unreliable. Any failure of a major component will impact mine performance.


Processing Plant

Cyanide exposure to personnel (note: contrary to the historical operation that employed cyanide bearing water in the ball mill, current tentative process designs consider usage of cyanide bearing waters in the leach tanks only). It is assumed that the international cyanide handling code will be adopted at site.

Safety Low High

Plant restart takes longer than anticipated. Cost incurred to continue restart efforts provides for refurbishments that were not anticipated. Increase mill refurbishment costs by 10% and delay production by 4 months

Economic High High

Availability of metallurgical expertise and key (historical) operating personnel within Wesdome is limited (resulting in inefficient plant operation). Profit loss 5% for 4 months.


Availability of process procedural documentation and access to key (historical) operating procedures is limited (resulting in inefficient plant operation). Profit loss 5% for 4 months.


Gravity circuit deemed necessary at a later stage, PFS testwork will include further study of this issue.


Need for variable speed drive in grinding mills due to variations in tonnage along the latest mine plan. Current configuration of the grinding mills is fix speed, this could cause overgrinding and not efficient use of the energy.


Need for cyanide destruction circuit to be operative during the first two years of operation. The current assumption is that the permit of the mine allows the operation to leach mineralized material with cyanide, without the use of a cyanide destruction plant to treat final tails.

Economic High Medium

Changes to process plant recovery estimates if the metallurgical recovery in some domains is less or greater than currently assumed, including the application of alternative processing methods.





JUNE 2020 25-11


Need of a larger filtering plant capable of treating 100% of the process plant tailings (for example, if backfill plant is not operative). Currently the filter plant was sized for 60% of the maximum throughput (850 tpd) indicated in the latest mine plan.


Contractors, personnel, and visitors

Theft of high-grade gold resources and related products, including mineralized material, liner residue, gravity concentrate if applicable, loaded carbon, doré, etc.

Economic Low Low

Availability of staff and contractors in the region is limited due to competition for such resources (failure to achieve production schedule). Profit loss 10% for 4 months.


Surroundings

Complaints from owners/occupants of residential properties. Community Low Medium

ATV riders trespass in the tailings dyke area or elsewhere, resulting in an injury to the ATV rider (see Stantec tailings impoundment report).

Community High Medium

Environment

Cyanide content in the mine exceeds limits permitted by regulations (possible consequence: cyanide destruction system required earlier than anticipated, adding related OPEX ~1.5 year earlier than anticipated).

Economic Medium Low

Ammonia content in tailings water exceeds limits permitted by regulations (possible consequence: a new treatment circuit and attendant CAPEX will be required).

Economic Low Medium

Existing cyanide destruction system cannot be refurbished with the allotted capital.

Economic Low Medium

Leakage of wastewater or reagents to Lac de Montigny (note that in 2014 waves damaged the tailings impoundment dyke).

Environmental Low High

Vehicle Infrastructure & Access Roads

Road access to Kiena Complex (rue Kienawisik) adversely affected by storm or lake water, thereby limiting access to the complex. Profit loss, no production for 3 days.

Environmental Low High

Facilities and Utilities

Electrical power supply from grid is irregular (i.e., Kiena load is increased, increased industrial activity in Val-d'Or since historical Kiena operation). Profit loss 5% for 4 months.

Economic Low High

Capacity of existing main substation insufficient to provide for new filter plant electrical load (possible mitigation - new substation directly connected to grid).


Remaining capacity in the existing tailings storage facility is greater than assumed in this study and the construction of the new dry stack facility can be delayed or eliminated.

Economic Low High


NI 43-101 – Technical Report ate


JUNE 2020 26-1

RECOMMENDATIONS

This NI 43-101 compliant technical report on the Kiena Mine Complex project was prepared by

experienced and competent independent consultants using accepted engineering methodologies

and standards. It provides a summary of the results and findings from each major area of

investigation including exploration, geological modelling, mineral resource, mine design,

metallurgy, process design, infrastructure, environmental management, tailings and water

management, capital and operating costs, and economic analysis. The level of investigation for

each of these areas is considered to be consistent with, or surpassing, the level expected of a

preliminary economic analysis.

The mutual conclusion of the QPs is that the Kiena Mine Complex project, as summarized in this

PEA, contains adequate detail and information to support the positive economic outcome shown.

The results of this study indicate that the Kiena Mine Complex project is technically feasible and

has financial merit at the base case assumptions considered.

In summary, the QPs recommend that the Project proceed to the prefeasibility study phase. It is

also recommended that environmental and permitting continue as needed to support the

development plans and the schedule of the Project.

An extensive work program including additional exploration drilling and the prefeasibility study has

been developed based on QP recommendations. The work program is estimated to cost

approximately $7.6M including a $1.2M contingency. A breakdown of this budget is summarized in

Table 26-1.

Table 26-1: Work program budget

Description Unit Cost ($)

Ongoing Exploration and Definition Drilling 50,000 m 5,000,000

Metallurgical Testwork 100,000

Geotechnical Drilling 130,000

Hydrogeological Program 150,000

Environment and Permitting 15,000

Conversion of the Polygonal Resources to Block Model 150,000

Mineral Resource Estimate and Prefeasibility Study 800,000

Contingency (20%) 1,269,000

Total 7,614,000

Analysis of the results and findings from each major area of investigation completed as part of this

preliminary economic assessment suggests numerous recommendations for further investigations

to mitigate risks and/or improve the base case designs. Sections 26.1 to 26.6 provide additional

details about the recommended work program outlined in Table 26-1.




JUNE 2020 26-2

Drilling and Geology

▪ Continue all ongoing exploration and definition drilling program of the Kiena Deep A Zones,

VC Zones and S50 Zones. Exploration drilling program should be done to continue

investigating any up and down plunge of the currently identified mineralization. A provision of

approximately 50,000 m should be considered.

▪ Convert the current Polygonal Resource to Block Modelling Resources in order to better

assess the potential economic viability. This should be preceded by a large database

compilation and validation as well as the modelling in 3D of all zones.

▪ A compilation of historic geological data adjacent to the Kiena Mine area and across the

entire Property is recommended to identify additional targets.

Metallurgical Testwork

High gold recoveries were achieved in the metallurgical testwork program regardless of the

flowsheet employed (WOL or GTL). However, additional tests need to be performed to confirm the

high gold recovery from mineralized material across the Kiena Mine Complex. The following actions

are recommended:

The following future testwork is recommended for the Kiena Deep deposit:

▪ A comminution testwork program to study the mineralized material hardness variability;

▪ A metallurgical testwork program to study the Au recovery variability with Au head grade

(based on the latest mine plan Au grade);

▪ Gravity recoverable gold (GRG) testwork to characterize the nature of the gravity gold in the

Kiena Deep A Zone. A cyanide leaching optimization program could be implemented

following a GRG testwork program;

▪ An optimization testwork program to study the optimization of leaching variables for the

option selected in the current testwork program (WOL or GTL):

- Stirred reactor tests could be conducted to validate or optimize process variables such

as cyanide addition, oxygen vs. air, lead nitrate addition, etc.

▪ A preliminary cyanide destruction and filtration testwork program based on the future tailings

handling system;

▪ A dynamic settling testwork program to optimize reagent addition.

It is recommended to build a preliminary grinding circuit model using information available to

estimate mill performance and final particle size. This information will be important to evaluate the

potential of operating the grinding circuit with only SAG mill (no ball mill) and also for designing the

filtration and tailings systems.




JUNE 2020 26-3

Mining

▪ Perform uniaxial laboratory studies on core samples from depth (basalt, ultramafic, schist);

▪ Update Geotech database, incorporating recent (2018-2020 diamond drilling). Review deep

core photos in Basalt for indicators of discing;

▪ Revisit stress assumptions, taking past history in existing Kiena mine workings at depth into

account;

▪ Closer examination of fault (Schist Zone) impact in proximity to HW and OZ. Kiena Deep A

Marbenite Fault (on FW side of zone) requires quantification;

▪ Examine impact of schist in Kiena Deep on stope wall stability;

▪ Numerical modelling (2D and 3D) of stress distribution associated with mining sequencing;

anticipated severity of ground damage:

- Impact of new mining zones on existing infrastructure;

- Seismicity potential at depth;

- Non-linear modelling of Deep A vein/intervein assemblage.

▪ There is limited drilling of the pillar area separating Kiena Deep Zone and Zones S50/B,

especially at depth. Are there major faults in this sector that may affect mining at depth?

What is the lithology?

▪ Review data from previous mining experience (lower parts of the old Kiena mine) to calibrate

proposed stope sizes;

▪ Ground support requirement to reduce dilution risk and to increase the stope sizes;

▪ Perform geotechnical drilling program;

▪ Complete ventilation design review with respect to heat accumulation to determine if or when

cooling is required at depth;

▪ Complete sill designs in Kiena Deep for each sector (lenses);

▪ Investigate the possibility of storing tailings behind hydrostatic plugs underground.

Recovery Methods

▪ Due to the differences in tonnage during the two operational phases, it is recommended to

investigate the feasibility of using only one grinding mill (i.e. SAG or Ball Mill) to operate

when the tonnage is less than 470 tpd. Additionally, it is recommended to explore the

addition of variable speed drives in the mills to accommodate variations in tonnage along the

LOM;

▪ It is recommended to conduct an audit to validate the absence of cyanide destruction circuit

during the first two years of operation of the process plant.




JUNE 2020 26-4

Infrastructure

▪ Investigate the remaining capacity of the existing tailings storage facility to confirm the need

and timing for a new facility;

▪ Investigate the need for a new mine rescue station;

▪ Perform a full inspection of the cyanide destruction equipment available at site;

▪ Investigate the location of future filter plant (determine options and optimal location);

▪ More analyses and modelling are required for the new dry stack tailings management facility

and should be performed in the next phase of the Project. It is recommended to:

- Gather additional geotechnical data to refine the stability model;

- Consider pore pressure dissipation between deposition sequences, for short term slope

stability analyses. It is important to note that the stress-strain-pore water pressure

model needs to be consistent with the final deposition plan of the TSF;

- Conduct a detailed seepage analysis to define the position of the phreatic surface within

the TSF;

- Conduct liquefaction analysis for the loose and saturated silt beneath the clay layer;

- Analyze different options to improve clay foundation strength (pre-load, wick drains,

other).

▪ Complete a detailed power study to further define the load distribution at site.

Environment and Permitting

▪ A geochemical kinetic test (Acid Rock Drainage and Metal Leaching) should be carried out

on representative mineralized material samples from the new mining zones or on

representative tailings samples produced during metallurgical testing for the new zones;

▪ Environmental baseline studies must be carried out early at the dry-stack location in order to

not delay obtainment of required operation permits.




JUNE 2020 27-1

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JUNE 2020

Appendix A: Detailed list of mineral claims

(verified on July 2, 2019)




JUNE 2020 Appendix A 1

Mineral claims detailed list (verified on July 2, 2019)

Claim No. Claim status

Issue date

Anniversary date

Area Ha

Owner Claim name

Land claim

403635183 Active 2014-11-26 2021-06-22 11.06 Wesdome Gold Mines Ltd. (84889) 100 % 2415481 CDC


403633306 Active 2014-10-14 2021-06-15 1.98 Mines Dynacor inc. (2146) 50 %; Wesdome Gold Mines

Ltd. (84889) 50 % 2411926 CDC


Ltd. (84889) 50 % 2411927 CDC


Ltd. (84889) 50 % 2411928 CDC


Ltd. (84889) 50 % 2411929 CDC


Ltd. (84889) 50 % 2411930 CDC


Ltd. (84889) 50 % 2411931 CDC

403637064 Active 2015-03-13 2021-04-24 32.63 9264-7890 Québec inc. (95234) 25 %; Wesdome Gold

Mines Ltd. (84889) 75 % 2421972 CDC


Mines Ltd. (84889) 75 % 2421973 CDC


Mines Ltd. (84889) 75 % 2421974 CDC


Mines Ltd. (84889) 75 % 2421975 CDC


Mines Ltd. (84889) 75 % 2421976 CDC


Mines Ltd. (84889) 75 % 2421977 CDC


Mines Ltd. (84889) 75 % 2421978 CDC


Mines Ltd. (84889) 75 % 2421979 CDC






Issue date

Anniversary date

Area Ha

Owner Claim name

Land claim

































Issue date

Anniversary date

Area Ha

Owner Claim name

Land claim

































Issue date

Anniversary date

Area Ha

Owner Claim name

Land claim

































Issue date

Anniversary date

Area Ha

Owner Claim name

Land claim

































Issue date

Anniversary date

Area Ha

Owner Claim name

Land claim

































Issue date

Anniversary date

Area Ha

Owner Claim name

Land claim

































Issue date

Anniversary date

Area Ha

Owner Claim name

Land claim






403542101 Active 1962-11-30 184.35 Wesdome Gold Mines Ltd. (84889) 100 % 494 CM




JUNE 2020

Appendix B: List of drillholes on the Kiena Mine

Complex Property




JUNE 2020 Appendix B 1

List of drillholes of the 2018-2019 Program on the Kiena Mine Complex Property included in this MRE

(UTM NAD 83 Zone 18N)

Survey Easting Northing Elevation Azimuth (°) Dip (°) Length (m) Status

6382 283221.64 5333993.22 -770.02 339.0 0.5 145.6 In MRE

6383 283579.08 5333823.37 -733.16 314.0 -73.0 594.0 In MRE

6384 283482.06 5333855.86 -741.98 12.1 -80.5 586.8 In MRE

6385 283578.17 5333822.79 -734.14 314.0 -67.0 534.3 In MRE

6386 283225.03 5333991.63 -770.29 46.0 -10.0 80.3 In MRE

6387 283224.81 5333991.49 -770.99 44.0 -41.0 100.0 In MRE

6388 283224.69 5333991.41 -771.30 37.0 -62.0 134.1 In MRE

6389 283136.76 5334114.16 -374.65 317.0 -9.0 350.0 In MRE

6390 283579.20 5333823.76 -733.14 315.0 -63.0 555.0 In MRE

6391 283320.07 5333988.63 -766.75 212.0 -48.0 251.5 In MRE

6392 283320.25 5333987.99 -766.82 209.0 -58.0 260.3 In MRE

6393 283320.62 5333988.32 -767.32 212.0 -65.0 284.3 In MRE

6394 283319.62 5333989.21 -767.07 225.0 -61.0 279.0 In MRE

6395 283319.69 5333989.25 -767.32 225.0 -70.0 285.0 In MRE

6396 283319.61 5333989.89 -767.12 248.0 -80.0 300.1 In MRE

6397 283482.49 5333856.37 -741.71 15.5 -76.0 452.3 In MRE

6398 283201.82 5334044.03 -769.79 315.0 -35.0 97.4 In MRE

6399 283201.88 5334044.07 -768.58 315.0 3.0 122.0 In MRE

6400 283201.63 5334044.26 -767.04 315.0 29.0 265.9 In MRE

6401 283578.54 5333823.53 -733.81 315.0 -59.0 460.0 In MRE

6402 283201.88 5334044.00 -770.08 315.0 -67.0 117.2 In MRE

6403 283201.85 5334044.07 -769.94 317.0 -52.5 140.0 In MRE

6404 283201.86 5334044.06 -769.27 318.0 -16.0 185.0 In MRE

6405 283202.00 5334044.02 -768.90 315.0 -5.0 188.3 In MRE

6406 283202.15 5334044.31 -770.13 333.0 -60.0 136.3 In MRE






6407 283578.50 5333823.58 -733.82 315.0 -55.0 239.4 In MRE

6408 283479.76 5333857.03 -741.87 315.0 -77.0 454.0 In MRE

6409 283578.56 5333823.22 -733.83 315.0 -72.0 659.3 In MRE

6410 283479.89 5333856.89 -741.87 313.0 -80.0 533.0 In MRE

6411 283578.59 5333823.15 -733.85 317.0 -76.0 673.7 In MRE

6412 283319.74 5333989.85 -767.13 254.0 -71.0 302.0 In MRE

6413 283357.35 5334016.62 -761.50 311.0 18.0 59.0 In MRE

6414 283357.14 5334016.93 -761.39 317.0 18.0 86.7 In MRE

6415 283136.64 5334114.38 -373.14 300.0 26.0 315.0 In MRE

6416 283136.76 5334114.39 -373.14 303.0 21.0 325.0 In MRE

6417 283648.68 5333859.62 -731.61 310.0 -76.0 40.5 In MRE

6418 283480.24 5333857.39 -742.15 315.0 -72.0 27.7 In MRE

6419 283202.14 5334044.39 -769.80 335.0 -50.0 89.0 In MRE

6420 283202.69 5334044.75 -769.58 348.0 -45.0 49.8 In MRE

6421 283202.71 5334044.75 -770.21 353.0 -61.0 58.0 In MRE

6422 283201.66 5334043.75 -769.58 325.0 -30.0 68.5 In MRE

6423 283200.96 5334043.00 -769.88 295.0 -46.0 98.3 In MRE

6424 283200.90 5334042.77 -770.22 286.0 -60.0 148.4 In MRE

6425 283324.59 5333992.27 -765.54 45.0 2.0 82.0 In MRE

6426 283323.58 5333993.41 -765.62 15.0 2.0 77.8 In MRE

6427 283578.45 5333823.06 -733.85 315.0 -74.0 699.4 In MRE

6428 283320.44 5333988.24 -767.15 225.0 -69.0 66.0 In MRE

6429 283320.43 5333988.19 -767.00 225.0 -65.0 255.0 In MRE

6430 283320.56 5333988.05 -766.77 225.0 -55.0 221.0 In MRE

6431 283320.50 5333988.66 -767.21 225.0 -77.0 273.7 In MRE

6432 283318.34 5333990.53 -767.14 160.0 -83.0 302.1 In MRE

6433 283200.96 5334043.29 -768.64 299.0 3.0 200.4 In MRE






6434 283201.03 5334043.45 -767.62 302.0 24.0 152.0 In MRE

6435 283200.94 5334043.02 -769.67 290.0 -27.0 99.5 In MRE

6436 283201.00 5334043.15 -768.69 288.0 3.0 117.7 In MRE

6437 283139.59 5334113.91 -375.49 46.0 -75.0 377.2 In MRE

6438 283480.13 5333857.57 -742.15 315.0 -69.0 25.7 In MRE

6439 283479.90 5333857.76 -742.15 315.0 -65.0 427.7 In MRE

6440 283479.97 5333857.63 -742.06 315.0 -61.0 346.0 In MRE

6441 283139.85 5334113.51 -375.51 63.0 -74.0 392.3 In MRE

6442 283137.04 5334114.47 -374.64 302.0 -1.0 326.7 In MRE

6443 283136.83 5334114.19 -374.88 302.0 -30.0 335.3 In MRE

6444 283136.89 5334114.14 -375.30 298.0 -50.0 319.6 In MRE

6445 283137.12 5334113.74 -375.54 285.0 -72.0 349.7 In MRE

6446 283171.67 5334034.94 -769.54 340.0 -25.0 152.7 In MRE

6447 283171.63 5334034.79 -769.54 340.0 -8.0 170.3 In MRE

6448 283172.24 5334035.38 -769.01 338.0 7.0 151.0 In MRE

6449 283172.07 5334035.04 -770.37 345.0 -38.0 112.0 In MRE

6450 283172.35 5334034.92 -770.75 358.0 -54.0 128.2 In MRE

6451 283647.91 5333859.07 -731.73 336.0 -85.0 58.0 In MRE

6452 283137.22 5334113.91 -375.31 304.0 -76.0 395.2 In MRE

6453 283137.07 5334114.10 -375.31 313.0 -60.0 377.3 In MRE

6454 283136.95 5334114.25 -374.96 313.0 -43.0 350.4 In MRE

6455 283319.82 5333989.97 -767.14 268.0 -81.0 232.0 In MRE

6456 283319.70 5333992.10 -767.22 315.0 -85.0 234.6 In MRE

6457 283321.43 5333992.99 -766.88 19.4.0 -85.0 165.5 In MRE

6458 283321.83 5333989.61 -767.22 138.1 -81.0 263.9 In MRE

6459 283319.22 5333989.70 -767.01 240.6 -65.5 273.1 In MRE

6460 283480.30 5333857.75 -742.10 322.0 -73.0 392.0 In MRE






6461 283480.41 5333857.72 -742.06 322.0 -75.0 452.0 In MRE

6462 283480.33 5333858.25 -742.14 322.3 -75.0 222.2 In MRE

6463 283480.33 5333858.25 -742.14 322.4 -81.3 485.0 In MRE

6464 283174.63 5334032.78 -771.00 45.0 -78.0 94.0 In MRE

6465 283174.86 5334033.06 -770.88 45.0 -61.0 139.1 In MRE

6466 283174.95 5334033.09 -770.46 45.0 -45.0 85.0 In MRE

6467 283174.93 5334033.10 -770.09 45.0 -33.0 86.0 In MRE

6468 283174.96 5334033.12 -769.84 45.0 -19.0 83.7 In MRE

6469 283174.90 5334033.01 -769.48 45.0 -4.0 95.0 In MRE

6470 283174.81 5334032.89 -768.90 45.0 10.0 101.5 In MRE

6471 283174.48 5334032.64 -770.25 45.0 -87.0 140.3 In MRE

6472 283651.98 5333854.32 -729.98 135.0 2.0 54.6 In MRE

6473 283651.80 5333854.52 -730.49 135.0 -30.0 30.5 In MRE

6474 283651.71 5333854.60 -731.28 135.0 -60.0 44.0 In MRE

6475 283651.54 5333854.83 -731.40 135.0 -80.0 32.2 In MRE

6476 283650.46 5333852.85 -729.45 180.0 2.0 32.0 In MRE

6477 283650.49 5333853.49 -731.57 180.0 -60.0 28.8 In MRE

6478 283653.38 5333858.17 -729.65 45.0 2.0 148.9 In MRE

6479 283579.20 5333823.66 -734.00 325.0 -72.0 567.2 In MRE

6480 283136.99 5334113.98 -375.45 297.5 -64.9 377.4 In MRE

6481 283136.80 5334113.77 -375.38 283.8 -75.4 377.4 In MRE

6482 283139.99 5334113.62 -375.53 42.4 -77.3 385.0 In MRE

6483 283139.68 5334114.10 -375.52 29.7 -77.0 389.4 In MRE

6484 283139.42 5334113.79 -375.50 20.1 -79.4 415.0 In MRE

6485 283173.13 5334034.29 -767.19 5.0 33.0 287.9 In MRE

6493 283173.24 5334034.35 -767.44 5.0 25.0 248.9 In MRE

6494 283173.13 5334034.13 -766.70 5.0 40.0 251.0 In MRE






6495 283319.18 5333989.76 -766.82 239.9 -59.4 255.3 In MRE

6496 283319.25 5333989.54 -766.62 233.6 -47.9 225.7 In MRE

6497 283319.13 5333989.57 -766.46 234.5 -38.3 227.1 In MRE

6500 283320.75 5333987.94 -767.14 234.1 -64.1 242.1 In MRE

6501 283319.19 5333989.76 -766.94 241.1 -72.2 23.7 In MRE

6510 283579.32 5333823.95 -734.15 315.4 -71.4 539.8 In MRE

6511 283137.97 5334114.70 -374.50 334.5 -19.3 275.7 In MRE

6512 283137.98 5334114.70 -374.64 334.6 -25.3 300.1 In MRE

6513 283137.96 5334114.70 -374.71 328.1 -29.1 410.7 In MRE

6514 283138.46 5334114.88 -374.35 345.9 -9.0 326.8 In MRE

6515 283138.42 5334114.93 -374.47 345.6 -14.7 326.6 In MRE

6516 283138.41 5334114.94 -374.63 343.9 -21.2 350.5 In MRE

6517 283173.55 5334034.1 -767.32 17.3 29.9 251.3 In MRE

6518 283173.55 5334034.21 -767.66 15.6 23.9 251.3 In MRE

6519 283173.55 5334034.09 -766.78 15.3 38.7 251.4 In MRE

6521 283137.84 5334114.65 -374.84 329.0 -28.4 18.3 In MRE

6522 283137.84 5334114.65 -374.84 329.6 -33.4 500.6 In MRE

6523 283137.84 5334114.65 -374.84 326.3 -34.2 500.6 In MRE

6524 283172.89 5334034.37 -767.19 354.8 32.3 243.5 In MRE

6527 283174.02 5334033.56 -767.26 30.8 31.6 251.8 In MRE

6529 283174.07 5334033.71 -767.01 31.4 36.0 251.0 In MRE

6531 283137.91 5334114.54 -375.17 325.9 -50.2 542.3 In MRE

6532 283137.93 5334114.53 -375.26 334.1 -62.0 377.3 In MRE

6533 283138.17 5334114.52 -375.37 342.1 -71.7 403.0 In MRE

6384A 283501.77 5333909.11 -1,041.96 13.7 -79.3 336.9 In MRE

6418A 283479.90 5333857.18 -742.06 315.0 -72.0 430.0 In MRE

6428A 283320.64 5333988.21 -767.08 215.0 -70.0 262.7 In MRE






6438A 283480.10 5333857.55 -742.15 315.0 -69.0 389.4 In MRE

6462A 283480.33 5333858.25 -742.14 322.4 -77.8 449.0 In MRE

6479A 283546.17 5333880.97 -961.04 320.1 -75.2 100.0 In MRE

6501A 283319.45 5333990.04 -767.11 240.6 -73.2 23.9 In MRE

6501B 283319.46 5333989.83 -767.80 239.6 -74.5 274.8 In MRE

6510A 283524.03 5333894.75 -1,010.85 324.1 -73.3 287.6 In MRE

6521A 283137.84 5334114.65 -374.84 331.2 -30.2 461.0 In MRE





List of drillholes not included in the MRE (Aug. 6, 2019 to Jan. 15, 2020) on the Kiena Mine Complex Property

(UTM NAD 83 Zone 18N)


6486 283172.7159 5334034.942 -769.9642 356.9 -46.7 100.5 Post-MRE

6487 283172.95 5334034.657 -769.663 5 -37.9 90.5 Post-MRE

6488 283171.6612 5334034.933 -769.5372 349.9 -23.7 101.4 Post-MRE

6489 283171.8108 5334034.917 -769.639 323.7 -28.4 80.1 Post-MRE

6490 283171.7591 5334034.977 -769.3522 324.4 -15.9 122.6 Post-MRE

6491 283171.9962 5334034.975 -769.8555 327.7 -36.7 88.3 Post-MRE

6492 283172.1498 5334034.91 -770.1275 331.1 -55.6 77.5 Post-MRE

6498 283319.3955 5333989.338 -766.8327 222.5 -48.8 227.3 Post-MRE

6499 283319.4152 5333989.318 -766.7193 225.8 -41.3 227.6 Post-MRE

6502 283320.4115 5333989.001 -767.1729 197.2 -78.1 272.8 Post-MRE

6503 283320.4785 5333989.173 -767.2574 208.1 -72.8 240.6 Post-MRE

6504 283320.2345 5333989.343 -766.8924 211 -68.1 250.8 Post-MRE

6505 283480.7014 5333856.722 -741.764 337.7 -81.6 518.5 Post-MRE

6506 283481.4557 5333857.427 -739.6255 8.1 -83.1 595.1 Post-MRE

6507 283480.7015 5333856.612 -741.9547 329.8 -80.5 529.2 Post-MRE

6508 283480.8742 5333856.868 -741.7493 337.1 -76.9 450.4 Post-MRE

6509 283481.1141 5333856.868 -741.7929 335.5 -76.5 455 Post-MRE

6520 283518.8113 5333719.116 -735.2447 351.5 -65.6 619.9 Post-MRE

6520A 283510.0215 5333850.477 -1027.62 354 -64.3 361.7 Post-MRE

6525 283172.808 5334034.72 -767.6844 354.9 23.4 238.8 Post-MRE

6526 283172.901 5334034.297 -766.8776 353.9 37.1 247.6 Post-MRE

6528 283174.1052 5334033.628 -767.7822 34.4 22.4 186 Post-MRE

6530 283172.8775 5334034.617 -767.9594 4.3 16.7 231 Post-MRE

6534 283172.2216 5334034.94 -766.6022 341.4 38.6 233 Post-MRE

6535 283172.205 5334034.977 -767.2895 343.8 30.5 245.4 Post-MRE






6536 283172.143 5334035.124 -767.5521 343.8 24.5 278 Post-MRE

6537 283172.1641 5334035.163 -767.9182 344.1 17.5 302.5 Post-MRE

6538 283331.3592 5333919.807 -761.5355 359.7 -84.8 321.8 Post-MRE

6539 283331.822 5333919.948 -761.5231 359.1 -73.7 269.3 Post-MRE

6540 283331.5755 5333920.183 -761.6331 358.5 -75.1 301.3 Post-MRE

6541 283331.7212 5333920.046 -761.6331 19.3 -82.9 287.3 Post-MRE

6542 283331.94 5333919.642 -761.428 21.5 -75 266 Post-MRE

6543 283332.0969 5333919.406 -761.384 28 -75.5 311 Post-MRE

6544 283332.3867 5333919.222 -761.389 41.9 -81.6 310.6 Post-MRE

6545 283332.6359 5333919.281 -761.442 40.9 -74.6 355 Post-MRE

6546 283137.5425 5334114.502 -374.8425 325.2 -41.1 502.5 Post-MRE

6547 283137.4698 5334114.479 -374.9339 334.8 -40.1 527.7 Post-MRE

6548 283137.0546 5334114.303 -374.7588 322.1 -36.4 512.5 Post-MRE

6549 283462.0357 5333840.183 -741.989 314.5 -50.8 331 Post-MRE

6550 283462.1837 5333840.049 -742.2329 313.8 -62.6 339.6 Post-MRE

6551 283462.2545 5333839.872 -742.3566 317.4 -65.6 356 Post-MRE

6552 283462.2694 5333839.641 -742.171 314.7 -68.6 460.7 Post-MRE

6553 283462.4319 5333839.508 -742.187 319.7 -72.8 383.9 Post-MRE

6554 283462.5955 5333839.635 -742.419 320.5 -81 454 Post-MRE

6555 283519.2613 5333718.889 -735.0641 2.9 -67.2 719 Post-MRE

6555A 283535.3701 5333840.451 -1036.26 7.7 -68.2 270.2 Post-MRE

6556 283137.0546 5334114.303 -374.7588 320.5 -29 517 Post-MRE

6557 283137.4036 5334114.462 -374.7813 319.5 -34 503.7 Post-MRE

6558 283178.8626 5334027.44 -770.682 39.5 -79.1 108.4 Post-MRE

6559 283179.2385 5334027.734 -770.6858 40.6 -64 135.2 Post-MRE

6560 283179.4351 5334027.858 -770.3616 43.9 -51.9 101.4 Post-MRE

6561 283179.3763 5334027.883 -769.9057 44.4 -38.6 89.5 Post-MRE






6562 283179.3505 5334027.824 -769.5739 43 -24.8 81.7 Post-MRE

6563 283179.3966 5334027.834 -769.2194 44.5 -11.2 87 Post-MRE

6564 283179.3583 5334027.802 -768.4078 42.9 11.7 100.1 Post-MRE

6565 283179.3645 5334027.714 -767.3993 43.2 29.2 104.5 Post-MRE

6566 283190.8604 5334014.956 -771.0826 43.3 -73.6 135 Post-MRE

6567 283191.1283 5334014.71 -770.854 40.6 -67.6 121.2 Post-MRE

6568 283190.9959 5334015.046 -770.916 40.1 -59.2 153.9 Post-MRE

6569 283191.1736 5334014.975 -770.877 37.9 -54.7 24 Post-MRE

6569A 283191.1666 5334014.92 -770.911 41.8 -52.5 108.8 Post-MRE

6570 283191.2572 5334014.991 -770.726 42.2 -44.1 90.5 Post-MRE

6571 283191.1795 5334014.915 -770.226 40.1 -32.8 86.4 Post-MRE

6572 283191.0978 5334014.925 -769.895 42.4 -20.6 90.5 Post-MRE

6573 283191.1826 5334015.004 -769.46 43.6 -2.9 82.5 Post-MRE

6574 283191.1074 5334014.942 -768.729 44.6 12.3 92.2 Post-MRE

6575 283191.1141 5334014.905 -768.271 46.3 23.1 101.4 Post-MRE

6576 283191.0918 5334014.899 -770.43 44 32.7 116.5 Post-MRE

6577 283137.3156 5334114.344 -375.0302 317.3 -45.8 417.8 Post-MRE

6578 283138.0374 5334114.436 -375.13 316.5 -39.6 403.1 Post-MRE

6579 283519.4375 5333718.97 -735.1669 8.3 -66.1 555.6 Post-MRE

6580 283520.0145 5333718.581 -735.2502 3.7 -69 781.8 Post-MRE

6580A 283548.2103 5333875.278 -1156.11 11.3 -69.7 310.4 Post-MRE

6580B 283546.1653 5333866.903 -1132.64 11.7 -69.8 347.1 Post-MRE

6581 283520.474 5333718.727 -735.13 351.8 -69.8 655.6 Post-MRE

6585 283519.3859 5333718.315 -733.5752 13.5 -76.4 667 Post-MRE

6586 283138.0374 5334114.436 -375.13 317.3 -22.9 491 Post-MRE

6587 283138.0374 5334114.436 -375.13 315.8 -18 560.6 Post-MRE

6588 283138.0374 5334114.436 -375.13 316.3 -13.4 526 Post-MRE






6589 283138.0374 5334114.436 -375.13 316.5 -7.9 526.1 Post-MRE

6590 283522.5751 5333717.941 -733.483 66.1 2.5 197.8 Post-MRE

6591 283522.4936 5333716.787 -733.401 84.6 3.9 255.8 Post-MRE

6592 283399.803 5333881.957 -751.559 288.1 -68.2 295.2 Post-MRE

6593 283397.8436 5333880.45 -751.63 280.8 -64.2 290.4 Post-MRE

6594 283397.8436 5333880.45 -751.63 274.6 -59.5 275 Post-MRE

6595 283397.8436 5333880.45 -751.63 267 -56.4 250 Post-MRE

6596 283397.8436 5333880.45 -751.63 273 -71.1 337.7 Post-MRE

6597 283397.8436 5333880.45 -751.63 269.4 -65.7 301 Post-MRE

6598 283397.8436 5333880.45 -751.63 274.1 -76 351.7 Post-MRE

6601 283461.6724 5333839.882 -742.13 353.6 -79.2 507.4 Post-MRE

6602 283462.0357 5333840.183 -741.989 354.9 -79.1 20.9 Post-MRE

6602A 283462.0331 5333840.18 -741.99 350.5 -79 511.4 Post-MRE

6603 283462.0357 5333840.183 -741.989 359.8 -74.6 369 Post-MRE

6604 283462.0357 5333840.183 -741.989 359.9 -79.4 516.9 Post-MRE

6605 283462.0357 5333840.183 -741.989 2.4 -77.5 23.9 Post-MRE

6606 283201.2227 5334004.136 -771.13 44.5 -68.2 162.7 Post-MRE

6607 283201.2227 5334004.136 -771.13 41.4 -44.7 86 Post-MRE

6608 283201.2227 5334004.136 -771.13 42.1 13.5 87 Post-MRE

6609 283201.2227 5334004.136 -771.13 43.3 25.8 126 Post-MRE

6610 283201.2227 5334004.136 -771.13 46.5 32.5 140 Post-MRE

6611 283224.4916 5333991.31 -771.73 44.5 -84.9 145.8 Post-MRE

6612 283224.4916 5333991.31 -771.73 41.9 -77.3 141.8 Post-MRE

6613 283224.4916 5333991.31 -771.73 41.6 -64 138.3 Post-MRE

6614 283224.4916 5333991.31 -771.73 44.6 6.1 89 Post-MRE

6615 283224.4916 5333991.31 -771.73 45.8 13.6 89.8 Post-MRE

6618 283138.0374 5334114.436 -375.13 312.2 -24 458.7 Post-MRE






6619 283138.0374 5334114.436 -375.13 313.3 -28.2 468.7 Post-MRE

6620 283138.0374 5334114.436 -375.13 312.8 -33.6 282.6 Post-MRE

6621 283244.3608 5333958.588 -772.93 44.4 -83.6 190 Post-MRE

6627 283118.9262 5333997.513 -522.43 287.5 1.9 527.3 Post-MRE

6628 283118.9262 5333997.513 -522.43 262.2 2.1 350.3 Post-MRE

6635 283118.9262 5333997.513 -522.43 235.8 2.6 500 Post-MRE




JUNE 2020

Appendix C: Polygonal resources summary




JUNE 2020 Appendix C 1

Martin Zone Indicated Resources

ZONE GROUP ZONE Polygone Name TRUE WIDTH (m) Cut Au (g/t) TONNES OUNCES (CUT) CLASSIFICATION 2019

Martin Martin A' (up) 3.00 6.53 1,050.00 220.53 Indicated

Martin Martin A (up) 3.00 3.71 9,872.02 1,178.06 Indicated

Martin Martin B (up) 3.00 4.67 8,290.54 1,245.54 Indicated

Martin Martin D (up) 3.00 3.73 12,627.33 1,515.05 Indicated

Martin Martin A' (down) 3.00 6.53 1,050.00 220.53 Indicated

Martin Martin A (down) 3.00 3.71 9,872.02 1,178.06 Indicated

Martin Martin B (down) 3.00 4.67 8,290.54 1,245.54 Indicated

Martin Martin D (down) 3.00 3.73 12,627.33 1,515.05 Indicated

Martin Martin S698 2.81 2.20 1,574.31 111.35 Indicated

Martin Martin U-1838 2.85 4.64 756.01 112.78 Indicated

Martin Martin U-1884 3.04 7.99 1,791.59 460.23 Indicated






Martin Martin U-4041 9.69 4.87 9,548.62 1,495.07 Indicated



















































































TOTAL Indicated 4.86 236,375.68 36,957.28

TOTAL Inferred - - -





Dubuisson Zone Indicated Resources


Dubuisson Dubuisson S531 7.60 4.25 48,228.57 6,589.98 Indicated


Dubuisson Dubuisson S541 3.32 0.87 15,793.26 441.76 Indicated

Dubuisson Dubuisson S551 7.18 9.40 2,402.95 726.21 Indicated


Dubuisson Dubuisson S557A 4.78 1.81 20,815.59 1,211.32 Indicated








TOTAL Indicated 5.46 281,520.56 49,377.35






Dubuisson North1 Indicated Resources

ZONE GROUP ZONE Polygone

Name HORIZONTAL

THICKNESS (m) Cut Au (g/t) TONNES OUNCES (CUT) CLASSIFICATION 2019

Dubuisson North1 Dubuisson North1 U5941 9.62 11.97 52,201.77 20,089.56 Indicated

Dubuisson North1 Dubuisson North1 U5947 7.13 2.22 43,728.32 3,121.09 Indicated

Dubuisson North1 Dubuisson North1 S755 2.90 22.58 22,457.35 16,303.22 Indicated

Dubuisson North1 Dubuisson North1 S756 2.50 3.36 8,317.89 898.55 Indicated





TOTAL Indicated 7.67 193,732.94 47,778.39


Dubuisson North2 Indicated Resources

ZONE GROUP ZONE Polygone Name HORIZONTAL THICKNESS (m) Cut Au (g/t) TONNES OUNCES (CUT) CLASSIFICATION 2019

Dubuisson Dubuisson North2 U5941 3.66 10.68 16,603.92 5,701.29 Indicated

Dubuisson Dubuisson North2 U5947 2.97 7.26 17,471.21 4,078.03 Indicated

Dubuisson Dubuisson North2 S763 4.58 4.81 24,484.25 3,786.37 Indicated

Dubuisson Dubuisson North2 S765A 3.59 5.41 22,044.08 3,834.25 Indicated

TOTAL Indicated 6.71 80,603.45 17,399.94






Northwest Indicated Resources


NorthWest NorthWest S194 3.90 3.50 20,567.57 2,314.42 Indicated






NorthWest NorthWest S281A 3.00 5.56 15,256.59 2,727.24 Indicated







NorthWest NorthWest U1633 4.07 4.20 19,083.42 2,576.89 Indicated

NorthWest NorthWest U1681 6.41 3.28 33,421.53 3,524.45 Indicated

TOTAL Indicated 3.79 467,358.33 57,010.26






Presqu'ile Zone 1 Indicated Resources


Presqu'ile Zone1 S-739 2.50 2.51 14,131.73 1,140.41 Indicated





Presqu'ile Zone1 S-748 2.50 0.04 3,367.64 4.33 Indicated





TOTAL Indicated 6.64 91,804.11 19,589.22


Presqu'ile Zone 2 Indicated Resources





TOTAL Indicated 8.67 51,164.85 14,263.11






Wesdome Zone A Indicated and Inferred Resources

ZONE GROUP ZONE Polygone Name HORIZONTAL THICKNESS (m) Au (g/t) TONNES OUNCES (CUT) CLASSIFICATION 2019

Wesdome A 82-054 1.67 11.40 4,444.00 1,628.81 Indicated


Wesdome A 82-059 2.34 5.40 2,220.00 385.42 Indicated


Wesdome A 82-064 1.50 4.60 621.00 91.84 Indicated










Wesdome A DS-037 1.66 5.50 3,130.00 553.48 Indicated

Wesdome A WD-281 1.50 6.00 1,639.00 316.17 Indicated

Wesdome A WD-281A 3.06 6.40 3,579.00 736.43 Indicated


Wesdome A WD-286 1.93 6.20 5,053.00 1,007.24 Indicated




Wesdome A WD-321 1.50 4.50 644.00 93.17 Indicated

Wesdome A WD-322 5.47 6.40 732.00 150.62 Indicated

Wesdome A WQ-76 4.51 15.30 5,617.00 2,763.04 Indicated

Wesdome A WQ-83 4.85 7.70 11,638.00 2,881.11 Indicated

Wesdome A WQ-87 1.50 6.90 2,973.00 659.53 Indicated







Wesdome A 82-050 2.01 7.80 13,946.00 3,497.32 Inferred



Wesdome A 82-064 1.50 4.60 1,109.00 164.01 Inferred

Wesdome A 82-066A 2.40 7.30 24,446.00 5,737.49 Inferred

Wesdome A 82-067A 1.50 4.10 5,117.00 674.51 Inferred



Wesdome A 82-079 4.63 6.10 173.00 33.93 Inferred









Wesdome A WD-248 2.68 4.00 23,380.00 3,006.74 Inferred






Wesdome A WD-281 1.50 6.00 168.00 32.41 Inferred

Wesdome A WD-281 1.50 6.00 153.00 29.51 Inferred

Wesdome A WD-281A 3.06 6.40 11,115.00 2,287.08 Inferred
















TOTAL Indicated 7.29 126,251.00 29,606.08

TOTAL Inferred 6.46 646,009.00 134,270.26





Wesdome Zone AF Indicated and Inferred Resources


Wesdome AF 82-088 1.50 6.30 2,686.00 544.05 Indicated

Wesdome AF 82-100 1.50 4.00 3,364.00 432.62 Indicated

Wesdome AF WD-302 2.52 11.80 5,923.00 2,247.06 Indicated

Wesdome AF WQ-83 3.71 7.50 7,339.00 1,769.66 Indicated

Wesdome AF 82-059 1.87 4.50 4,916.00 711.24 Inferred

Wesdome AF 82-088 1.50 6.30 743.00 150.49 Inferred

Wesdome AF 82-100 1.50 4.00 6,316.00 812.26 Inferred

Wesdome AF WD-302 2.52 11.80 7,098.00 2,692.83 Inferred

Wesdome AF WQ-83 3.71 7.50 19,823.00 4,779.93 Inferred

Wesdome AF WQ-89 1.50 4.80 5,389.00 831.65 Inferred

TOTAL Indicated 8.04 19,312.00 4,993.39

TOTAL Inferred 7.01 44,285.00 9,978.40

Wesdome Zone AH Indicated and Inferred Resources


Wesdome AH WD-286 1.50 10.00 7,089.00 2,279.17 Inferred

Wesdome AH 82-065 2.33 10.90 7,558.00 2,648.65 Indicated

Wesdome AH 82-079 1.50 13.90 3,056.00 1,365.71 Indicated

Wesdome AH WD-263 1.50 10.80 3,375.00 1,171.89 Indicated

Wesdome AH WD-286 1.50 10.00 3,774.00 1,213.37 Indicated

TOTAL Indicated 11.21 17,763.00 6,399.63

TOTAL Inferred 10.00 7,089.00 2,279.17





Wesdome Zone AH2 Inferred Resources


Wesdome AH2 82-139 2.95 4.40 5,180.00 732.78 Inferred

Wesdome AH2 82-154 1.50 5.40 6,110.00 1,060.78 Inferred

Wesdome AH2 82-183 1.50 13.10 22,913.00 9,650.38 Inferred

TOTAL Indicated - - -

TOTAL Inferred 10.41 34,203.00 11,443.94

Wesdome Zone A1 Indicated and Inferred Resources


Wesdome A1 82-071 1.50 4.60 6,159.00 910.88 Indicated

Wesdome A1 WD-314 1.50 6.20 6,121.00 1,220.13 Indicated

Wesdome A1 WD-261 3.88 24.90 11,659.00 9,333.65 Inferred

Wesdome A1 WD-314 1.50 6.20 3,947.00 786.77 Inferred

TOTAL Indicated 5.40 12,280.00 2,131.00

TOTAL Inferred 20.17 15,606.00 10,120.43





Wesdome Zone B Indicated and Inferred Resources


Wesdome B 82-064 2.13 6.10 10,990.00 2,155.35 Indicated

Wesdome B 82-067A 1.50 5.50 5,307.00 938.43 Indicated

Wesdome B WD-302 1.50 4.30 11,066.00 1,529.85 Indicated

Wesdome B WQ-76 1.50 6.10 8,408.00 1,648.97 Indicated

Wesdome B WQ-83 1.50 11.20 7,907.00 2,847.22 Indicated

Wesdome B 82-063 1.66 6.00 2,179.00 420.34 Inferred

Wesdome B 82-064 2.13 6.10 1,480.00 290.26 Inferred

Wesdome B 82-067A 1.50 5.50 11,835.00 2,092.77 Inferred

Wesdome B WD-274 1.50 10.80 16,735.00 5,810.86 Inferred

Wesdome B WD-302 1.50 4.30 12.00 1.66 Inferred

Wesdome B WD-311 1.50 8.70 25,076.00 7,014.05 Inferred

TOTAL Indicated 6.49 43,678.00 9,119.83

TOTAL Inferred 8.48 57,317.00 15,629.93

Wesdome Zone D Indicated and Inferred Resources


Wesdome D 82-070 1.50 4.90 8,104.00 1,276.69 Indicated

Wesdome D WD-314 3.05 4.80 13,211.00 2,038.77 Indicated

Wesdome D 82-073 1.50 4.60 16,052.00 2,373.99 Inferred

Wesdome D CB-055 1.50 4.90 4,280.00 674.27 Inferred

Wesdome D WD-283 1.50 4.00 16,540.00 2,127.09 Inferred

TOTAL Indicated 4.84 21,315.00 3,315.46

TOTAL Inferred 4.37 36,872.00 5,175.34





Wesdome Zone E Indicated and Inferred Resources


Wesdome E 82-213 1.50 4.90 6,520.00 1,027.15 Indicated



Wesdome E 82-227 1.96 5.70 4,340.00 795.35 Indicated

Wesdome E WD-258 2.17 4.50 18,657.00 2,699.26 Indicated



Wesdome E 82-126 1.50 4.50 18,882.00 2,731.82 Inferred











Wesdome E 82-221 1.50 4.40 1,535.00 217.15 Inferred




Wesdome E WD-258 2.17 4.50 14,814.00 2,143.27 Inferred



TOTAL Indicated 5.44 96,584.00 16,900.00

TOTAL Inferred 6.86 434,950.00 95,922.07





Wesdome Zone E0 Indicated and Inferred Resources


Wesdome E0 82-208 1.50 10.50 5,451.00 1,840.16 Indicated

Wesdome E0 82-180 1.50 12.50 30,753.00 12,359.15 Inferred


Wesdome E0 WD-270 10.83 5.10 160,762.00 26,359.95 Inferred

TOTAL Indicated 10.50 5,451.00 1,840.16

TOTAL Inferred 6.52 202,703.00 42,495.98





Wesdome E1 82-229 1.50 5.30 5,231.00 891.36 Indicated

Wesdome E1 WD-265 1.50 4.70 2,325.00 351.33 Indicated


Wesdome E1 82-213 1.50 4.90 186.00 29.30 Inferred






Wesdome E1 WD-265 1.50 4.70 4,485.00 677.72 Inferred

TOTAL Indicated 5.26 60,060.00 10,155.65

TOTAL Inferred 5.62 94,406.00 17,046.06








Wesdome E3 WD-258 1.50 4.20 10,185.00 1,375.31 Indicated






Wesdome E3 WD-260 1.50 4.40 6,995.00 989.54 Inferred


TOTAL Indicated 5.48 39,115.00 6,891.92

TOTAL Inferred 6.63 65,695.00 14,012.04

Wesdome Zone E4 Inferred Resources








TOTAL Inferred 8.80 86,593.00 24,501.72





Wesdome Zone F Inferred Resources


Wesdome F 82-208 1.50 5.80 19,289.00 3,596.90 Inferred

Wesdome F WD-268 1.50 6.40 16,555.00 3,406.44 Inferred


TOTAL Inferred 6.08 35,844.00 7,003.34

Wesdome Zone F2 Indicated and Inferred Resources


Wesdome F2 CB-071 2.17 12.80 3,015.00 1,240.76 Indicated

Wesdome F2 WD-262 2.60 4.20 5,742.00 775.36 Indicated

Wesdome F2 82-228 1.50 10.10 20,715.00 6,726.63 Inferred

TOTAL Indicated 7.16 8,757.00 2,016.12

TOTAL Inferred 10.10 20,715.00 6,726.63

Wesdome Zone F4 Indicated and Inferred Resources


Wesdome F4 82-212 1.50 11.20 3,693.00 1,329.81 Indicated

Wesdome F4 WD-266 1.50 8.00 2,445.00 628.87 Indicated

Wesdome F4 82-229 1.50 7.90 5,420.00 1,376.63 Inferred

Wesdome F4 WD-264 3.68 10.10 9,150.00 2,971.21 Inferred

TOTAL Indicated 9.93 6,138.00 1,958.67

TOTAL Inferred 9.28 14,570.00 4,347.84




JUNE 2020

Appendix D: Tailings Storage Facility (TSF)

Dry stack concept

▪ Location of mining infrastructure

▪ Plan view

▪ Profiles and details

TSF

305

310

315

319

DITCH A

DITCH B

LIMITS FOR THE NEW ACCUMULATION AREA(STANTEC, 2019)

BERM(CREST ELEVATION: 304)

CORESHACK TO BE RELOCATED

TAILINGS LINE

TAILINGS FILTERINGPLANT

PUMPINGSTATION

PROPOSED RELOCATION OFCORESHACK

POLISHING POND

SOUTH CELL

NORTH CELL

DRAINAGE PUMPING LINE

RAILWAY

ROAD 117

EXISTING TSF

MINING SITE

LAKE MONTIGNY

FINAL EFFLUENT

RECLAIM WATER LINE

12 11 10 9 8 7 6 5 4 3

B

C

D

E

F

G

H

11 10 9 8 7 6 5 4 3 2 1

A

B

C

D

E

F

G

H

PROJECT:

TITLE:

SEAL:

CLIENT: DESIGNED BY:

APPROVED BY:VERIFIED BY:

0 1 2 3 4 5 6 7 8 9 10 cm

DRAWING No.:

SCALE:

DRAFTED BY:

DATE:

SHEET: SIZE: REV.A1

G:\3767\002\40_ING_ENG\41_CIVIL\DESS_DWGS\PROJET_PROJECT\EN _COURS\3767002-00000-41-D20-0001_EN COURS.DWGPR

INTE

D:2

020-

02-2

5 17

:04:

04B

Y:S

AN

AB

RIA

, MA

UR

ICIO

KIENA MINE COMPLEX PEA

TAILINGS STORAGE FACILITYDRY STACK CONCEPT

LOCATION OF MINING INFRASTRACTURE

M. SANABRIA/JA. VIDES M. SANABRIA

JA. VIDES L. PICIACCHIA

1:4000 2020-02-25

3767002-000001-41-D20-0001 01 AA

3767002-000001-41-D20-0001

DRAWING No.

REFERENCE DRAWINGS

DESCRIPTION

REVISIONS

REV VERIFIED BY APPROVED BYDESCRIPTION DATE

- - - - --

-- -

m0050 100 200 300 400

1: 4000

NOTES:

1. COORDINATES ARE SHOWN IN UTM, NAD 83, ZONE18.

2. TERRAIN MODEL DERIVED FROM LIDAR MADE IN2017 BY THE MINISTER OF FORESTS, WILDLIFE ANDPARKS (QUEBEC).

3. LOCATION OF STRUCTURES IS APPROXIMATE. THEEXACT LOCATION WILL BE ADJUSTED ON THE SITE.

53

34

00

0 N

281000 E281500 E

LEGEND

MAYOR CONTOUR

MINOR CONTOUR

SLOPEHIGH EDGE

LOW EDGE

FUTUREPERIMETER DITCH

282000 E

PLAN VIEW

D

A

T

E

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TSF

300297.50

297.50

297.50

297.50

296.50

297.50

297.50

305

310

315

319

A_

B _DITCH A

DITCH B

LIMITS FOR THE NEWACCUMULATION AREA(STANTEC, 2019)

BERM(CREST ELEVATION: 304)

POLISHING POND

B0+000

B0+100

B0+20

0

B0+300B0+400

B0+500B0+562

297.50

297.50

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295

A-01

A-03

A-04

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B-04B-05INTERSECTION POINT

CORESHACK TO BERELOCATED (SEE NOTE 4)

A0+789

A0+000

A0+100

A0+200

A0+300

A0+

400

A0+

500

A0+

600

A0+700

A-02TAILINGS LINE

TAILINGSFILTERING PLANT

PUMPING STATION(SEE NOTE 8)

RECLAIM WATER LINE

304m

2.531 1

4.00

BERM

SURFACE AFTER REMOVING WASTE ROCK (SEE NOTE 5)

WASTE ROCK 1.00

51

(SEE NOTE 6)TSF

(SEE NOTE 7) VARIAB. FROM 4 TO 6.5m

ÉLÉ

V. (m

)

CHAÎNAGE (m)

290

300

310

320

290

300

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320

0+000 0+100 0+200 0+300 0+399

BERM5

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319


BERM1.00EXISTING TERRAIN

TSF

1.00

DITCH A DITCH B

WASTE ROCK PILE

300

(SEE NOTE 6)

CHAÎNAGE (m)

290

300

310

320

290

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320

0+000 0+100 0+200 0+300 0+384

BERM 511.00

51

319

BERM1.00

DITCH B DITCH A

EXISTING TERRAIN


WASTE ROCK PILE

300

TSF(SEE NOTE 6)

ÉLÉ

V. (m

)

CHAÎNAGE (m)

PROFIL Ditch Opt2-2

292

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A0+000 A0+100 A0+200 A0+300 A0+400 A0+500 A0+600 A0+700 A0+789

A-0

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47.5

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leva

tion=

297.

48

A-0

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82.0

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tion=

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28

A-0

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tion=

296.

97

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4E

leva

tion=

296.

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DITCH A

SURFACE AFTERREMOVING WASTE ROCK

CHAÎNAGE (m)

292

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tatio

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tion=

296.

97

B-0

2 S

tatio

n=0+

93.4

3E

leva

tion=

297.

91

B-0

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tatio

n=2+

38.9

9E

leva

tion=

297.

62

B-0

4 S

tatio

n=2+

97.3

9E

leva

tion=

297.

50

B-0

5 S

tatio

n=3+

86.9

6E

leva

tion=

297.

32

B-0

1 S

tatio

n=0+

00.0

0E

leva

tion=

298.

10

INTERSECTION POINTWITH DITCH A

DITCH B

SURFACE AFTERREMOVING WASTE ROCK

SURFACE AFTER REMOVING WASTE ROCK

RIP-RAP

CLAY BACKFILL

1.00 MIN1.50

0.50

31

31

1.50

GEOTEXTILE TEXEL 918 OR EQUIVALENT

1.00

1.00

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PROJECT:

TITLE:

SEAL:

CLIENT: DESIGNED BY:

APPROVED BY:VERIFIED BY:

0 1 2 3 4 5 6 7 8 9 10 cm

DRAWING No.:

SCALE:

DRAFTED BY:

DATE:

SHEET: SIZE: REV.A1

G:\3767\002\40_ING_ENG\41_CIVIL\DESS_DWGS\PROJET_PROJECT\EN _COURS\3767002-00000-41-D20-0002_EN COURS.DWGPR

INTE

D:2

020-

02-2

5 17

:09:

36B

Y:S

AN

AB

RIA

, MA

UR

ICIO

KIENA MINE COMPLEX PEA

TAILINGS STORAGE FACILITYDRY STACK CONCEPT

PLAN VIEW, PROFILES AND DETAILS

M. SANABRIA/JA. VIDES M. SANABRIA

JA. VIDES L. PICIACCHIA

INDICATED 2020-02-25

3767002-000001-41-D20-0002 01 AA

3767002-000001-41-D20-0002

DRAWING No.

REFERENCE DRAWINGS

DESCRIPTION

REVISIONS

REV VERIFIED BY APPROVED BYDESCRIPTION DATE

- - - - --

-- -

00 m2040 40 80 120 160 200

1: 2000

NOTES:

1. COORDINATES ARE SHOWN IN UTM, NAD 83, ZONE 18.

2. TERRAIN MODEL DERIVED FROM LIDAR MADE IN 2017 BY THE MINISTER OFFORESTS, WILDLIFE AND PARKS (QUEBEC).

3. LOCATION OF STRUCTURES IS APPROXIMATE. THE EXACT LOCATION WILL BEADJUSTED ON THE SITE.

4. RELOCATION OF THE CORE SHACK WILL NEED TO BE IN AN AREA WHEREWATER MANAGEMENT IS ALREADY IN PLACE.

5. THE WASTE ROCK PILE HAS AN APPROXIMATIVE WEIGHT OF 300 000 METRICTONS (STANTEC, 2020), WHICH REPRESENTS 175 000M³. SURFACE ELEVATIONAFTER REMOVING WASTE ROCK IS BASED ON THIS VOLUME.

6. DRY STACK TSF COMPACTION: MATERIAL TO BE PLACED IN 300mm LAYERSAND COMPACTED TO 95% MODIFIED PROCTOR DENSITY.

7. WASTE ROCK COMPACTION: MATERIAL TO BE PLACED IN 1000mm LAYERSCOMPACTED TO 95% MODIFIED PROCTOR DENSITY WITH SIX PASSES OF THECOMPACTING MACHINE.

8. IF THE WATER QUALITY EXCEEDS THE ENVIRONMENTAL DESIGN CRITERIA, THEPUMPING STATION WILL TAKE THE DRAINAGE WATER BACK TO THE EXISTINGTSF.

53

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282250 E282500 E

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75

0 N

LEGEND

MAYOR CONTOUR

MINOR CONTOUR

SLOPEHIGH EDGE

LOW EDGE

FUTUREPERIMETER DITCH

282750 E

CROSS SECTION ASCALE (1:1000)

CROSS SECTION BSCALE (1:1000)

BERM TYPICAL CROSS SECTIONSCALE (1:250)

PLAN VIEW

DITCH A -PROFILESCALE (HOR. 1:2500 VERT: 250)

DITCH B -PROFILESCALE (HOR. 1:2500 VERT: 250)

DITCH TYPICAL CROSS SECTIONSCALE (1:100)

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5




JUNE 2020

Appendix E: CPM Independent Gold Price

Projections (March 30, 2020)

168 Seventh St., Suite 310 Telephone: (212) 785-8320 Brooklyn, NY 11215 USA Fax: (585) 625-3672 [email protected] www.cpmgroup.com

CPM Group

CPM Independent Gold Price Projections Produced for: Wesdome Gold Mines Ltd. 30 March 2020

2

FOR INTERNAL CORPORATE USE ONLY. NOT FOR REPRODUCTION, DISTRIBUTION, OR RETRANSMISSION OUTSIDE OF CPM CLIENT COMPANY WITHOUT WRITTEN CPM GROUP CONSENT. CPM Group 168 Seventh Street Suite 310 Brooklyn, NY 11215 USA Telephone: 212-785-8320 Telefax: 212-785-8325 E-mail: [email protected] Website: http://www.cpmgroup.com 30 March 2020 The information contained here has been obtained from sources we believe to be reliable. We believe this information to be reliable, but do not guarantee its accuracy or completeness. Opinions expressed here represent those of CPM Group at the time of publication. This material is for private use of subscribers. CPM Group is not soliciting any action based on it. Information contained here should not be relied on as specific investment or market timing advice. At times the principals and associates of CPM Group may have long or short positions in some of the markets mentioned here.

3

Contents

Price Projections The Basis For The Price Projections Why We Use Projections Instead Of Historical Averages CPM Group Expertise and Past Price Projections Canadian Dollar

4

Prices CPM Group is projecting a nominal average price of $1,644 for gold in 2020. That would be the second highest nominal annual average gold price in history, surpassed only by the $1,670 in 2012. As of 23 March the price had averaged $1,579. This was 21% higher than the average price in the same period in early 2019. The gold price will have to be significantly higher in the final nine months of 2020 in order to achieve that annual average of $1,644. Given the state of the world, such a gold price may be readily achievable. For the past several years CPM had been projecting a more modest rise in gold prices in 2019 – 2021 followed by a much sharper increase to record real and nominal prices in 2023 – 2025. The first quarter of 2020 has delivered a stronger price increase than we had anticipated. It is possible that prices may level off, at presently elevated levels, as CPM projects for the next year or so. CPM still projects a sharp increase in real and nominal prices into 2023, however. As discussed in the next section of this report, that reflects CPM’s projections that economic and financial market conditions will combine with hostile political developments in many parts of the world to stimulate a sharp increase in private investor demand for physical gold around 2023 – 2025. After that period of sharply increased investment demand and gold prices, CPM projects prices will subside, as they always have following such periods of unsettled global economics and politics.

$0

$200

$400

$600

$800

$1,000

$1,200

$1,400

$1,600

$1,800

$2,000

$2,200

$2,400

$0

$200

$400

$600

$800

$1,000

$1,200

$1,400

$1,600

$1,800

$2,000

$2,200

$2,400

1950 1955 1960 1965 1970 1975 1980 1985 1990 1995 2000 2005 2010 2015 2020p 2025p

$/Ounce $/Ounce

Nominal

Real

Base Price = 2018 Actual

Projections

5

Scenario 1: CPM Group's Main ProjectionsReal and Nominal Gold Prices

Prices, Base=2019

U.S. CPIPercent Percent Percent

Year Real Change Nominal Change Change

1950 $353 - $35 - -1955 $316 - $35 - -1960 $285 - $35 - -1965 $268 - $35 - -1970 $268 $41 5.7%

1975 $755 40.4% $161 65.5% 9.1%1976 $542 -28.3% $125 -22.5% 5.8%1977 $606 11.8% $148 18.3% 6.5%1978 $746 23.2% $193 30.8% 7.6%1979 $1,100 47.5% $307 58.7% 11.3%

1980 $2,046 86.0% $612 99.5% 13.5%1981 $1,326 -35.2% $460 -24.9% 10.3%1982 $1,003 -24.4% $376 -18.3% 6.1%1983 $1,099 9.6% $424 12.9% 3.2%1984 $887 -19.4% $360 -15.1% 4.3%

1985 $749 -15.5% $317 -12.0% 3.5%1986 $854 13.9% $368 15.8% 1.9%1987 $1,006 17.8% $446 21.5% 3.7%1988 $943 -6.2% $437 -2.1% 4.1%1989 $780 -17.3% $382 -12.5% 4.8%

1990 $741 -5.0% $384 0.4% 5.4%1991 $668 -9.8% $362 -5.6% 4.2%1992 $613 -8.2% $344 -5.2% 3.0%1993 $624 1.7% $360 4.7% 3.0%1994 $650 4.1% $384 6.7% 2.6%

1995 $632 -2.6% $385 0.2% 2.8%1996 $621 -1.8% $389 1.2% 2.9%1997 $517 -16.9% $333 -14.5% 2.3%1998 $451 -12.8% $295 -11.2% 1.6%1999 $417 -7.4% $280 -5.3% 2.2%

2000 $404 -3.1% $280 0.2% 3.4%2001 $380 -6.0% $272 -3.1% 2.8%2002 $429 12.9% $311 14.5% 1.6%2003 $492 14.8% $364 17.1% 2.3%2004 $541 10.0% $410 12.6% 2.7%

2005 $571 5.5% $446 8.9% 3.4%2006 $758 32.7% $607 35.9% 3.2%2007 $853 12.5% $700 15.4% 2.9%2008 $1,030 20.8% $873 24.7% 3.8%2009 $1,154 12.0% $975 11.7% -0.4%

2010 $1,229 6.5% $1,229 26.1% 1.6%2011 $1,792 45.8% $1,572 27.9% 3.2%2012 $1,867 4.2% $1,670 6.2% 2.1%2013 $1,548 -17.0% $1,410 -15.6% 1.4%2014 $1,365 -11.8% $1,266 -10.2% 1.6%

2015 $1,248 0.1% $1,159 -8.5% 0.1%2016 $1,331 6.7% $1,251 7.9% 1.3%2017 $1,312 -1.5% $1,259 0.7% 2.1%2018 $1,292 -1.5% $1,271 0.9% 2.4%2019 $1,395 8.0% $1,395 9.8% 1.8%2020p $1,616 15.8% $1,644 17.8% 2.0%2021p $1,632 1.0% $1,693 3.0% 2.0%2022p $1,640 0.5% $1,735 2.5% 2.0%2023p $1,837 12.0% $1,978 14.0% 2.0%2024p $2,112 15.0% $2,314 17.0% 2.0%2025p $1,584 -25.0% $1,782 -23.0% 2.0%2026p $1,347 -15.0% $1,550 -13.0% 2.0%2027p $1,212 -10.0% $1,426 -8.0% 2.0%

AveragesActuals

1950 - 1975 $337 4.4% $49 7.7% 3.3% 1950 - 1990 $565 4.2% $161 8.6% 4.4% 1950 - 2013 $650 4.2% $326 7.9% 3.7% 1970 - 2013 $813 6.6% $458 10.9% 4.3% 1977 - 1990 $992 6.2% $372 12.4% 6.2% 1977 - 2019 $937 4.0% $629 7.6% 3.6% 1990 - 2019 $904 2.7% $741 5.1% 2.4% 2000 - 2019 $1,060 7.0% $936 9.2% 2.2%

Projections 2020 - 2024 $1,767 8.9% $1,873 10.9% 2.0% 2020 - 2029 $1,532 -1.0% $1,696 1.0% 2.0%

Source: CPM Group28 March 2020

6

The Basis For The Price Projections

Long-term gold prices are determined primarily by the levels of demand for physical gold by investors and monetary authorities (central banks). Combined these are forms of Stock Demand for gold: Gold buying in which the gold remains in ‘gold’ form, readily fungible and saleable in bars, coins, and other gold products. Of these two, private sector investment demand is the more powerful. Central banks and other monetary authorities tend to be more price elastic than private investors: The monetary authorities prefer to buy gold when the price is low, and tend to buy less when the price rises. Investors tend to chase prices higher, and pull back from buying so much when prices fall. (This is in contrast with Fabrication Demand, in which the gold is used, transformed into jewelry, decorative objects, electronic components, dental alloys, medical devices, and a host of items plated with gold for various reasons.) The simple model is that investment demand, Stock Demand, determines the price of gold, and the price of gold determines fabrication demand, mine production, and secondary recovery of gold from scrap. These other segments of the physical gold market influence gold prices, but the impact of gold prices on fabrication demand, mine production, and secondary supply tends to be more powerful than the effect of these market sectors on gold prices. Stock demand in turn is determined primarily by economic, financial market, and political factors exogenous or external to gold. Some investors buy gold as a commodity, but most investors buy gold based on factors outside the market: Trends in economic growth or recession, inflation, currency market volatility, the strength or weakness of equities and other securities markets, interest rate levels, political stability, banking industry stability, and other factors that affect investors’ sense of personal security. Since 2001 gold has entered a new era of rising gold prices based on a larger group of private investors in more countries buying more gold at higher prices for longer periods of time than ever before, even during World War Two. Around 2000 CPM began writing about this coming “Gold Renaissance” in which private investors first, joined by central banks later, would ‘rediscover’ gold as a financial asset and quasi-currency.

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100

120

66 69 72 75 78 81 84 87 90 93 96 99 02 05 08 11 14 17 20p 23p 26p 29p

Percent Change

Percent Change in Price

Investment Demand and Gold PricesAnnual, Projected Through 2029

Million Ounces

Net Investment Demand (Left Scale)

7

While this has been good for gold prices and gold producers, it is a reflection of the reality that global economic trends, financial market instability and scandals, a deterioration of the ethical standards in governance, decreased international political cooperation, growing economic inequality, extended wars and revolutions in numerous parts of the world, and a host of other economic, financial, and political issues have caused investors to buy more gold. CPM Group’s price projections are modeled on the expectation that these hostile economic and political trends continue during the coming decade, and possibly beyond. This will keep investors interested in buying and holding more gold, as a form of wealth preservation, a portfolio diversifier, an alternative denominator for their wealth and investment portfolio, and for other financial reasons. More than 70% of private financial wealth is denominated in U.S. dollars, a condition many investors would like to reduce. Central banks also will continue to buy gold on a net basis, as a way to diversify away from the U.S. dollar and the euro. Around 62% of monetary reserves held by central banks around the world are in U.S. dollars, with another 20% or so in euro. Central banks want to diversify their monetary reserves, and gold is one asset they are using, along with other currencies, to do so. The key to CPM’s projected higher prices lies with private investment demand, primarily. Lower mine production and increased central bank buying will contribute to higher prices, but ultimately it is the large pool of private financial wealth that is the key gold price determinant. The economic and political environment is expected to remain hostile to overall economic activity in the coming decade. Many of the problems that contributed to the recessions and financial crises in 2000 - 2001 and again in 2007 – 2011 have not been resolved. Many have grown worse since 2011. Other new problems have arisen, compounding these problems. The massive amounts of public and private debt, deficit spending by major governments, the resulting mounting sovereign debt, the deterioration of intergovernmental cooperation, the gathering pace of computerization across industries and services that will compound surplus labor problems, and a host of other issues are expected to keep investors interested in buying and holding gold for at least the next 10 years.

Gold Supply and Demand . Annual, Available Supply Is Net Of Official Transactions

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50 54 58 62 66 70 74 78 82 86 90 94 98 02 06 10 14 18p 22p 26p

Million Ounces Million O

Available Supply

Investment Demand

Million Ounces Million O

Available Supply

Fabrication Demand

8

Scenario 1: CPM Group's Main ProjectionsGold Supply, Demand, and PricesMillion Troy Ounces

Supply

Mine Output Flow fromin Market Percent Transitional Percent Secondary Percent Total Percent

Year Economies Change Economies Change Recovery Change Supply Change

1950 24.3 - - - - 24.3 -1955 26.8 - 2.2 - - - 29.0 -1960 33.5 - 5.7 - - - 39.2 -1965 41.0 - 11.4 - - - 52.4 -1970 40.9 - -0.1 - - - 40.8 -

1975 30.9 -5.7% 4.8 -32.4% - - 35.7 -10.5%1976 31.2 0.8% 13.2 175.0% - - 44.4 24.2%1977 31.0 -0.5% 12.9 -2.3% 4.5 - 48.4 9.1%1978 31.1 0.1% 13.2 2.3% 5.7 26.7% 50.0 3.2%1979 30.7 -1.2% 6.4 -51.5% 10.9 91.2% 48.0 -4.0%

1980 30.5 -0.7% 2.9 -54.7% 21.2 94.5% 54.6 13.7%1981 31.2 2.3% 9.0 210.3% 10.5 -50.5% 50.7 -7.1%1982 32.4 4.0% 6.5 -27.8% 11.1 5.7% 50.0 -1.3%1983 34.2 5.5% 2.7 -58.5% 11.7 5.4% 48.6 -2.8%1984 35.7 4.3% 3.9 44.4% 10.4 -11.1% 50.0 2.8%

1985 37.9 6.2% 7.6 94.9% 9.0 -13.5% 54.5 9.0%1986 40.3 6.3% 13.2 73.7% 13.0 44.4% 66.5 22.0%1987 42.1 4.5% 10.1 -23.5% 12.8 -1.5% 65.0 -2.2%1988 47.5 12.7% 10.1 0.0% 11.1 -13.3% 68.7 5.6%1989 51.6 8.6% 11.0 8.9% 9.7 -12.6% 72.3 5.2%

1990 54.5 5.7% 10.0 -9.1% 10.7 10.3% 75.2 4.1%1991 56.1 2.9% 10.0 0.0% 11.3 5.6% 77.4 2.9%1992 58.6 4.6% 11.0 10.0% 12.0 6.2% 81.6 5.5%1993 59.9 2.3% 11.0 0.0% 12.6 5.0% 83.5 2.4%1994 60.2 0.5% 10.0 -9.1% 13.8 9.5% 84.0 0.6%

1995 60.1 -0.3% 11.0 10.0% 14.5 5.1% 85.6 1.8%1996 61.5 2.4% 11.5 4.5% 17.0 17.2% 90.0 5.2%1997 64.8 5.4% 12.0 4.3% 15.5 -8.8% 92.3 2.6%1998 67.1 3.5% 15.0 25.0% 22.6 45.8% 104.7 13.4%1999 68.4 2.0% 17.0 13.3% 20.0 -11.5% 105.4 0.7%

2000 77.3 13.0% 4.1 -76.0% 22.7 13.3% 104.1 -1.3%2001 77.4 0.1% 4.4 8.8% 23.6 4.3% 105.5 1.3%2002 76.1 -1.6% 4.2 -5.9% 30.9 30.9% 111.2 5.5%2003 76.1 0.0% 4.4 4.3% 33.9 9.5% 114.4 2.8%2004 72.3 -4.9% 4.4 0.8% 32.0 -5.6% 108.7 -4.9%

2005 74.2 2.6% 4.0 -8.1% 33.3 4.0% 111.5 2.6%2006 71.8 -3.2% 3.6 -9.8% 31.6 -4.9% 107.1 -3.9%2007 67.7 -5.8% 3.7 0.5% 38.8 22.5% 110.1 2.8%2008 69.4 2.5% 3.8 3.7% 45.2 16.6% 118.4 7.5%2009 75.3 8.6% 3.9 3.2% 47.1 4.1% 126.3 6.7%

2010 78.1 3.7% 4.3 9.3% 46.3 -1.5% 128.7 1.9%2011 78.3 0.3% 4.8 13.3% 46.8 1.0% 130.0 1.0%2012 80.9 3.2% 3.5 -28.3% 47.2 0.9% 131.5 1.2%2013 85.6 5.9% 2.2 -36.6% 39.3 -16.8% 127.1 -3.4%2014 89.9 5.0% 2.8 27.3% 35.1 -10.7% 127.8 0.6%

2015 90.0 0.1% 2.9 3.6% 30.1 -14.2% 123.0 -3.7%2016 92.3 2.5% 4.1 41.1% 29.8 -1.1% 126.2 2.6%2017 94.3 2.2% 4.9 20.7% 29.8 0.1% 129.1 2.3%2018 93.8 -0.6% 5.3 7.4% 29.2 -1.9% 128.3 -0.6%

2019 92.4 -1.5% 5.8 9.6% 29.9 2.3% 128.1 -0.2%2020p 90.9 -1.6% 4.6 -20.2% 31.0 3.7% 126.6 -1.2%2021p 90.8 -0.1% 3.9 -16.6% 31.3 0.9% 126.0 -0.5%2022p 90.0 -0.9% 4.1 6.2% 31.6 0.9% 125.7 -0.2%2023p 87.6 -2.7% 4.1 0.0% 33.0 4.6% 124.7 -0.8%2024p 85.6 -2.2% 4.1 -0.2% 34.9 5.6% 124.6 -0.1%2025p 83.5 -2.4% 4.0 -1.7% 34.1 -2.3% 121.6 -2.4%2026p 80.8 -3.3% 3.6 -11.6% 31.7 -7.0% 116.0 -4.6%2027p 80.0 -1.0% 3.6 1.9% 30.7 -3.2% 114.3 -1.5%

Compounded Growth RatesActuals

1950 - 1975 1.0% 3.6% NA 1.6% 1950 - 1990 2.0% 4.2% NA 2.9% 1950 - 2013 2.0% 0.0% NA 2.7% 1970 - 2013 1.7% 0.6% NA 2.7% 1977 - 1990 4.4% -1.9% 6.9% 3.4% 1977 - 2019 2.6% -1.9% 4.6% 2.3% 1990 - 2019 1.8% -1.9% 3.6% 1.9% 2000 - 2019 0.9% 1.9% 1.5% 1.1%

Projections 2020 - 2024 -1.5% -3.0% 3.0% -0.4% 2020 - 2029 -1.3% -2.7% -0.3% -1.1%

Notes: * Compounded growth rates for transitional economies are available from 1953; growth rates for secon--dary supply are available from 1977. NA -- data on secondary supply is not available for years prior to 1977. Source: CPM Group28 March 2020

9

Scenario 1: CPM Group's Main ProjectionsGold Supply, Demand, and PricesMillion Troy Ounces

DemandOfficial Private Investment

Transactions Jewelry Coins BullionPurchases (+) Available Percent and Percent and Purchases (+) Total

or Sales (-) Supply Change Industrial Change Medallions or Sales (-) Investment Year

9.2 15.1 - 12.0 - - 3.1 3.1 195019.0 10.0 - 13.5 - - -3.5 -3.5 19558.3 30.9 - 25.0 - - 5.9 5.9 19606.3 46.1 - 36.0 - - 10.1 10.1 19659.4 31.3 - 41.1 - 3.2 -13.0 -9.8 1970

2.6 33.2 -23.6% 22.8 59.4% 8.8 1.6 10.4 1975-0.8 45.1 36.1% 37.2 63.2% 7.5 0.4 7.9 1976-9.8 58.3 29.1% 43.9 18.0% 6.3 8.1 14.4 1977

-11.7 61.7 5.9% 45.2 3.0% 10.8 5.7 16.5 1978-17.5 65.5 6.2% 39.7 -12.2% 10.4 15.4 25.8 1979

8.1 46.4 -29.1% 25.7 -35.2% 8.9 11.8 20.7 19805.1 45.6 -1.9% 33.4 29.7% 8.2 4.0 12.2 19811.6 48.4 6.3% 36.7 10.0% 6.5 5.2 11.7 1982

-2.1 50.8 4.8% 35.6 -3.0% 6.8 8.4 15.2 1983-1.6 51.6 1.7% 41.6 17.0% 5.5 4.5 10.0 1984

1.5 53.0 2.7% 44.1 6.0% 5.2 3.7 8.9 19851.2 65.4 23.4% 44.4 0.6% 11.4 9.6 21.0 1986

-2.9 67.9 3.9% 45.0 1.3% 7.5 15.4 22.9 19875.4 63.3 -6.8% 56.5 25.7% 5.2 1.6 6.8 1988

-6.6 78.8 24.6% 63.0 11.4% 5.7 10.1 15.8 1989

-1.4 76.6 -2.8% 68.5 8.8% 4.6 3.5 8.1 1990-1.8 79.2 3.3% 70.1 2.3% 5.8 3.3 9.1 1991

-13.0 94.6 19.5% 78.0 11.2% 3.5 13.2 16.7 1992-21.4 105.0 10.9% 82.7 6.0% 4.4 17.9 22.2 1993-5.6 89.7 -14.6% 86.6 4.7% 2.8 0.3 3.1 1994

-9.2 94.8 5.7% 94.2 8.8% 3.1 -2.5 0.6 1995-3.8 93.8 -1.0% 95.5 1.5% 2.6 -4.3 -1.7 1996

-31.2 123.6 31.7% 100.9 5.6% 3.7 19.0 22.7 1997-13.2 117.9 -4.6% 104.3 3.4% 4.4 9.1 13.5 1998-14.9 120.3 2.1% 101.8 -2.4% 4.6 13.9 18.5 1999

-10.8 114.9 -4.5% 112.4 10.5% 3.3 -0.9 2.4 2000-9.0 114.4 -0.4% 99.5 -11.5% 3.2 11.8 15.0 2001

-11.8 123.1 7.5% 92.0 -7.5% 4.1 26.9 31.0 2002-17.9 132.2 7.4% 88.9 -3.4% 4.4 38.9 43.3 2003-16.7 125.4 -5.1% 92.4 3.9% 5.0 28.1 33.1 2004

-20.4 131.9 5.2% 95.0 2.9% 5.6 31.4 36.9 2005-8.8 115.9 -12.1% 94.9 -0.1% 7.1 13.9 21.0 2006

-16.2 126.3 9.0% 97.9 3.1% 6.6 21.8 28.4 20070.6 117.8 -6.8% 95.2 -2.8% 10.0 12.6 22.6 20080.2 126.1 7.0% 79.4 -16.6% 11.2 35.5 46.7 2009

11.9 116.9 -7.3% 82.9 4.4% 11.1 22.9 34.0 20109.7 120.3 2.9% 81.2 -2.0% 10.0 29.1 39.1 2011

11.4 120.1 -0.1% 82.6 1.8% 8.3 29.2 37.5 201217.7 109.4 -8.9% 92.0 11.3% 10.6 6.8 17.4 201316.7 111.1 1.5% 93.5 1.7% 8.3 9.2 17.5 2014

12.4 110.6 -0.4% 96.7 3.4% 9.0 4.9 13.9 20157.4 118.8 7.4% 93.6 -3.2% 10.3 14.9 25.2 2016

10.4 118.7 0.0% 97.0 3.7% 7.5 14.2 21.7 201716.2 112.1 -5.6% 98.1 1.1% 5.9 8.1 14.0 201817.5 110.6 -1.3% 94.0 -4.2% 6.5 10.1 16.6 201919.0 107.6 -2.7% 90.4 -3.8% 6.8 10.4 17.2 2020p18.0 108.0 0.4% 90.2 -0.3% 0.0 0.0 17.8 2021p16.0 109.7 1.6% 89.0 -1.3% 0.0 0.0 20.7 2022p8.0 116.7 6.4% 86.9 -2.4% 0.0 0.0 29.8 2023p5.0 119.6 2.5% 83.3 -4.1% 0.0 0.0 36.3 2024p8.0 113.6 -5.0% 84.8 1.8% 0.0 0.0 28.8 2025p

10.0 106.0 -6.7% 85.9 1.3% 0.0 0.0 20.1 2026p10.0 104.3 -1.6% 86.9 1.2% 0.0 0.0 17.4 2027p

Compounded Growth RatesActuals

1950 - 1975 3.2% 2.6% 1950 - 1990 4.1% 4.5% 1950 - 2013 3.2% 3.3% 1970 - 2013 3.0% 1.9% 1977 - 1990 2.1% 3.5% 1977 - 2019 1.5% 1.8% 1990 - 2019 1.3% 1.1% 2000 - 2019 -0.2% -0.9%

Projections 2020 - 2024 2.7% -2.0% 2020 - 2029 -0.3% -0.1%

Demand statistics for years prior to 1977 represent use of gold from mine production, transitional economysales, and official sector sources, and exclude gold from scrap, while demand statistics beginning with 1977 Coins and medallions prior to 1968 are included in the jewelry and industrial column. There may be some discrepancies due to rounding.

10

Why We Use Projections Instead Of Historical Averages Commodity prices are notoriously volatile. This is true across metals as well as other commodities. The average price over the past three years, often used in various valuation programs, typically varies widely from the prices that develop over the subsequent three years. Historical prices are not a good predictor of future prices Projecting future prices is necessary to value mining and mineral properties for a number of reasons, including undertaking Preliminary Economic Assessments. Estimating, guessing what the prices of the products of mines will be in the future is needed for valuing properties, reserves and resources. Different companies employ various methodologies for projecting prices.

Rolling average historical prices, typically 1, 3, 5 years Spot prices Forward prices from futures markets Consensus forecasts by brokers, analysts, and others Commodity pricing expert Internal management generated forecasts

Half of the practitioners surveyed in a KPMG survey in 2013 used analytically derived projected prices. Nearly one quarter of them use futures prices (mislabeled forward prices in this survey). Some used spot prices, which in the survey may have included averages of historical prices. (It was not clear.)

19%

24%

31%

19%

7%

0%

5%

10%

15%

20%

25%

30%

35%

How do you determine the expected commodity prices for valuation purposes?

Spot price Forward prices Consensus of forecast prices by

brokers/economists

Commodity pricing expert

Other

Source: KPMG 2013

% of Participants

11

Historical Prices Do Not Predict The Future Well The chart and table below show the variance between average gold prices over the previous three years and the subsequent three years on a rolling average from 1980 through 2013.

-40%

-20%

0%

20%

40%

60%

80%

100%

120%

140%

-40%

-20%

0%

20%

40%

60%

80%

100%

120%

140%

1980 1983 1986 1989 1992 1995 1998 2001 2004 2007 2010 2013

Percentage Variance Between Past 3 Years and Forward 3 YearsRolling Average Gold Prices

Rolling Average Gold Prices

YearPast 3 Years

Future 3 Years

% Variance

2002 $277 $362 30.4%

2006 $407 $727 78.6%

2010 $849 $1,490 75.5%

2014 $1,551 $1,225 -21.0%

Average 17.8%

12

We are including similar data on copper to illustrate that it is not only gold for which this sharp variance between recent past prices and future prices exists.

-50%

0%

50%

100%

150%

200%

-50%

0%

50%

100%

150%

200%

2000 2001 2002 2003 2004 2005 2006 2007 2008 2009 2010 2011 2012 2013 2014

Copper: Percentage Variance Between Past 3 Years and Forward 3 YearsRolling Average Copper Prices

13

CPM Group Expertise and Past Price Projections CPM Group’s analysts have been studying the gold market, along with other commodities markets, since 1980, when what is now CPM Group was the Research Department at J. Aron and Company. (J. Aron was acquired by Goldman Sachs in 1981. The Commodities Research Group spun out of Goldman in 1986 to set up CPM Group.) We have been in the forefront of the economic and econometric analysis of gold, precious metals, and other commodities since that time. We have been producing 10-year projections of gold supply, demand, and price since the early 1980s. Additionally, we have published short, intermediate, and long term price projections, including buy and sell recommendations. The chart below shows CPM’s published intermediate term, 2 – 3 year, buy and sell recommendations since December 1980. We have a good track record, basing our price expectations and projections on fundamental and macroeconomic analyses. CPM Group Intermediate-Term Gold Recommendations .

CPM has provided independent gold price projections for use in a number of projects for companies around the world. Our gold price projections have been used in basic corporate strategies, focusing on the wisdom of targeting gold in exploration and development programs. They have been used in Preliminary Economic Assessments, Pre-Feasibility Studies, Bankable Full Feasibility Studies, prospectuses for initial public offerings, secondary offerings, rights offerings, and other applications. Our clients in using these projections have included Anglogold, Ashanti Goldfields, Anglogold Ashanti, Barrick, Newmont, Normandy, Placer Dome, Penoles, Luismin, Goldcorp, Lonmin, Norilsk Nickel, Novagold, Wheaton Precious Metals, and many other mining companies over the past four decades. Additionally, CPM Group’s Managing Partner, Jeffrey M. Christian, was the expert on gold pricing and price forecasting for Barrick Gold in Barrick’s 2015 lawsuit against the Western Australian state tax authorities on valuation issues related to the acquisition by Barrick of Placer Dome. In addition to mining companies, CPM’s long-term price projections are used by institutional investors, banks and financial institutions, central banks and monetary authorities, intergovernmental organizations such as the United Nations, World Bank, and International Monetary Fund, and others.

0

500

1,000

1,500

2,000

2,500

0

500

1,000

1,500

2,000

2,500

80 81 82 83 84 85 86 87 88 89 90 91 92 93 94 95 96 97 98 99 00 01 02 03 04 05 06 07 08 09 10 11 12 13 14 15 16 17 18 19

$ / Ounce $ / Ounce

CO VER/BUY

SELL/SHO RT

$595

$315

$420$346

$444$330

$324

$256

$311

$293

$1,600

$1,211

14

Canadian Dollar CPM Group projects that the Canadian dollar is most likely to weaken against the U.S. dollar over the next couple of years, 2020 and 2021, possibly into 2022. This reflects the expected negative impact on the Canadian economy of the oil price collapse and the oil price war that broke out between Saudi Arabia and Russia in the first quarter 2020, along with issues in the Canadian oil sands and broader petroleum industry. Beyond that expectations are that the Canadian economy may fare better than the United States economy through the rest of the decade, benefiting the Canadian dollar, while the continued global tendency to reduce exposure to the U.S. dollar in reserve asset holdings may further weaken the U.S. dollar somewhat against the Canadian dollar and other currencies. The strengthening of the Canadian dollar against the U.S. dollar beyond 2022 is expected to be relatively modest and to remain within the range of recent decades.

0.80

0.90

1.00

1.10

1.20

1.30

1.40

1.50

1.60

1.70

1.80

1971

1973

1975

1977

1979

1981

1983

1985

1987

1989

1991

1993

1995

1997

1999

2001

2003

2005

2007

2009

2011

2013

2015

2017

2019

2020

2022

2024

2026

2028

USD/Cad Annual Average Exchange Rate

15

US$ per C$ Exchange Rates Annual Averages, Projected through 2029

1971 $1.0101972 $0.9911973 $1.0001974 $0.9781975 $1.0171976 $0.9861977 $1.0641978 $1.1411979 $1.1721980 $1.1691981 $1.1991982 $1.2341983 $1.2331984 $1.2951985 $1.3661986 $1.3891987 $1.3261988 $1.2311989 $1.1841990 $1.1671991 $1.1461992 $1.2091993 $1.2901994 $1.3661995 $1.3721996 $1.3641997 $1.3851998 $1.4841999 $1.4862000 $1.4852001 $1.5492002 $1.5702003 $1.4002004 $1.3012005 $1.2112006 $1.1342007 $1.0742008 $1.0672009 $1.1412010 $1.0302011 $0.9892012 $1.0002013 $1.0302014 $1.1052015 $1.2792016 $1.3252017 $1.2982018 $1.2962019 $1.327

projections2020 $1.3602021 $1.4002022 $1.3902023 $1.3502024 $1.3102025 $1.2702026 $1.2402027 $1.2002028 $1.1602029 $1.120

Source: CPM Group30 March 2020

Documents

PRELIMINARY ECONOMIC ASSESSMENT FOR THE KIENA MINE … · I, Jorge Torrealba, P. Eng., employed with BBA Inc., do hereby consent to the filing of the Technical Report prepared for