43-101 technical report for pea of the o'brien project

InnovExplo Inc. Consultants–Mines–Exploration 560, 3e Avenue, Val-d’Or (Québec) J9P 1S4 Telephone: 819.874-0447 Facsimile: 819.874-0379 Toll-free: 866.749-8140 Email: [email protected] Web site: www.innovexplo.com

WSP Canada Inc. 1075, 3e Avenue

Val-d’Or (Québec) J9P 0J7 Telephone: 819-8254711

Facsmile 819-825-4715 Web Site: www.wspgroup.com

Lamont Inc. 10, chemin des Conifères

Lac-Beauport (Québec) G3B 2E7 Telephone: 418-928-9028

Web Site: www.lamont-expertconseil.com

TECHNICAL REPORT AND PRELIMINARY ECONOMIC ASSESSMENT FOR THE O’BRIEN PROJECT, ABITIBI, QUÉBEC

(according to National Instrument 43-101 and Form 43-101F1)

Project Location Latitude: 48º 14’ 07’’ North; Longitude: 78º 22’ 54’’ West

Cadillac Township Province of Québec, Canada

Prepared for

Radisson Mining Resources Inc. C.P. 307

Rouyn-Noranda, Québec Canada J9X 5C3

Prepared by:

Sylvie Poirier, Eng. Pierre-Luc Richard, M.Sc., P.Geo. Bruno Turcotte, P.Geo. Laurent Roy, Eng.

Annie Lavoie, Eng. Éric Poirier, Eng. Marie-Claude Dion St-Pierre, Eng. M.A.Sc.

Ann Lamontagne, Eng., Ph.D.

InnovExplo Inc. WSP Canada Inc. Lamont Inc.

Effective Date: November 29, 2015 Signature Date: January 29, 2016

mailto:[email protected]

http://www.innovexplo.com/

http://www.wspgroup.com/

http://www.lamont-expertconseil.com/

www.innovexplo.com

43-101 Technical Report – O’Brien Project 2

TABLE OF CONTENTS SIGNATURE PAGE – INNOVEXPLO ........................................................................................................ 10

SIGNATURE PAGE – INNOVEXPLO ........................................................................................................ 11

SIGNATURE PAGE – WSP CANADA INC. ............................................................................................... 12



SIGNATURE PAGE – LAMONT INC. ........................................................................................................ 15

CERTIFICATE OF AUTHOR – SYLVIE POIRIER...................................................................................... 16

CERTIFICATE OF AUTHOR – PIERRE-LUC RICHARD .......................................................................... 17

CERTIFICATE OF AUTHOR – BRUNO TURCOTTE ................................................................................ 18

CERTIFICATE OF AUTHOR – LAURENT ROY ........................................................................................ 19

CERTIFICATE OF AUTHOR – ANNIE LAVOIE......................................................................................... 20

CERTIFICATE OF AUTHOR – ÉRIC POIRIER .......................................................................................... 21

CERTIFICATE OF AUTHOR – MARIE-CLAUDE DION ST-PIERRE ........................................................ 22

CERTIFICATE OF AUTHOR – ANN LAMONTAGNE ............................................................................... 23

1. SUMMARY .......................................................................................................................................... 24

Introduction ................................................................................................................................ 24 Property Description and Location ......................................................................................... 24 Geological Setting and Mineralization .................................................................................... 25 Data Verification ........................................................................................................................ 25 Mineral Resource Estimates..................................................................................................... 25 Metallurgy and Milling ............................................................................................................... 28 Environment ............................................................................................................................... 29 Mining Plan................................................................................................................................. 30 Capital and operating cost ....................................................................................................... 31

Financial analysis ...................................................................................................................... 33 Risks and Opportunities ........................................................................................................... 35 Recommendations .................................................................................................................... 37

2. INTRODUCTION ................................................................................................................................. 41

Principal sources of information ............................................................................................. 41 Qualified persons and inspection of the Project.................................................................... 42 Note regarding the 2015 Preliminary Economic Assessment .............................................. 43 Units and Currencies ................................................................................................................ 44

3. RELIANCE ON OTHER EXPERTS .................................................................................................... 45

4. PROPERTY DESCRIPTIONS AND LOCATIONS .............................................................................. 46

Location ...................................................................................................................................... 46 Mining Rights in the Province of Québec ............................................................................... 47 Current Property Description ................................................................................................... 47 Historical Property Description................................................................................................ 47 Urban Perimeter ......................................................................................................................... 50 Territory Akin to an Area for Vacationing ............................................................................... 50 Permits ........................................................................................................................................ 50 Environmental Liabilities .......................................................................................................... 51 Comments on Item 4 ................................................................................................................. 51

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5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY ........................................................................................................................................ 52

Accessibility ............................................................................................................................... 52 Climate ........................................................................................................................................ 54 Local Resources and Infrastructure ........................................................................................ 54 Physiography ............................................................................................................................. 54

6. HISTORY ............................................................................................................................................. 56

O’Brien Property ........................................................................................................................ 56 O’Brien Gold Mines Ltd ....................................................................................................... 56 Darius Gold Mines Inc. ........................................................................................................ 60 Sulpetro Minerals / Novamin Resources / Breakwater Resources ..................................... 62 Radisson Mining Resources ................................................................................................ 63

Kewagama Property .................................................................................................................. 69 Kewagama Gold Mines Ltd ................................................................................................. 69 Sulpetro Minerals / Novamin Resources / Breakwater Resources ..................................... 71 Radisson Mining Resources ................................................................................................ 71

7. GEOLOGICAL SETTING AND MINERALIZATION ........................................................................... 74

Abitibi Terrane (Abitibi Subprovince) ...................................................................................... 74 Cadillac Area .............................................................................................................................. 76 Property Geology ...................................................................................................................... 79

Cadillac Group ..................................................................................................................... 79 Piché Group ......................................................................................................................... 79

7.3.2.1 Porphyritic andesite ....................................................................................................................... 79 7.3.2.2 Conglomerate ................................................................................................................................ 80 7.3.2.3 Volcanic rocks ................................................................................................................................ 80 7.3.2.4 Graphitic schist .............................................................................................................................. 80

Pontiac Group ...................................................................................................................... 80 Mineralization ............................................................................................................................. 80

O’Brien mine ........................................................................................................................ 80 7.4.1.1 No. 1 Vein ...................................................................................................................................... 81 7.4.1.2 No. 4 Vein ...................................................................................................................................... 81 7.4.1.3 No. 9 Vein ...................................................................................................................................... 81

Zone 36E area ..................................................................................................................... 81 Kewagama area................................................................................................................... 82

Hydrothermal Alteration ........................................................................................................... 83

8. DEPOSIT TYPES ................................................................................................................................ 84

9. EXPLORATION ................................................................................................................................... 87

Type 1 targets ............................................................................................................................ 87 Type 2 targets ............................................................................................................................ 87 Type 3 targets ............................................................................................................................ 87

10. DRILLING ........................................................................................................................................ 89

11. SAMPLE PREPARATION, ANALYSIS, AND SECURITY ............................................................. 90

12. DATA VERIFICATION .................................................................................................................... 91

Historical Work .......................................................................................................................... 91 Radisson Database ................................................................................................................... 91 Radisson Diamond Drilling ....................................................................................................... 91 Radisson Logging, Sampling and Assaying Procedures ..................................................... 92 Mined-out Voids ......................................................................................................................... 94 Conclusion ................................................................................................................................. 94

13. MINERAL PROCESSING AND METALLURGICAL TESTING ..................................................... 95

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Historical Data ........................................................................................................................... 95 Darius mill ............................................................................................................................ 95 Review of historical testwork ............................................................................................... 97

Zone 36E AREA Testwork ......................................................................................................... 98 Gravity separation................................................................................................................ 99 Flotation grind size............................................................................................................... 99 Combination of gravity and flotation .................................................................................. 100 Cyclic flotation tests ........................................................................................................... 101 Combination of gravity and cyanidation ............................................................................ 102

14. MINERAL RESOURCE ESTIMATES ........................................................................................... 103

Drill Hole Database .................................................................................................................. 103 Interpretation of Mineralized Zones ....................................................................................... 105 Underground Workings .......................................................................................................... 106 High Grade Capping ................................................................................................................ 108 Compositing ............................................................................................................................. 113 Bulk Density ............................................................................................................................. 115 Block Model.............................................................................................................................. 115 Variography and Search Ellipsoids ....................................................................................... 118 Grade Interpolation ................................................................................................................. 118

Resource Categories ........................................................................................................... 120 Mineral resource classification definition ....................................................................... 120 Mineral resource classification ...................................................................................... 120

Cut-off Grade ........................................................................................................................ 125 Mineral Resource Estimate ................................................................................................. 126

15. MINERAL RESERVE ESTIMATES .............................................................................................. 128

16. MINING METHODS ....................................................................................................................... 129

Cautionary Statement ............................................................................................................. 129 Introduction .............................................................................................................................. 129 Mineral Resources Considered in the Mining Plan .............................................................. 129 Potentially Mineable Mineral Resources ............................................................................... 129

Cut-off grade ...................................................................................................................... 130 Geotechnical Evaluation ......................................................................................................... 130

Typical ground support patterns ........................................................................................ 130 Mining Methods ....................................................................................................................... 131

Modified Avoca mining method ......................................................................................... 131 Long-hole method with captive sublevels .......................................................................... 132

16.6.2.1 Mining dilution and recoveries ...................................................................................................... 132 Kewagama shaft dewatering and shaft rehabilitation ......................................................... 132 Underground mine design ...................................................................................................... 133 Primary development .............................................................................................................. 133

Secondary development .................................................................................................... 133 Stope development ............................................................................................................ 133 Stope ground support ........................................................................................................ 134

Mine Sequence..................................................................................................................... 134 Mining Rate .......................................................................................................................... 134 Mine plan schedule criteria ................................................................................................ 134 Development and Production Schedule ........................................................................... 135 Equipment Selection and Requirements .......................................................................... 137 Manpower Requirements .................................................................................................... 137 Mining Services ................................................................................................................... 138

Ventilation ...................................................................................................................... 138 Dewatering ..................................................................................................................... 138 Compressed air ............................................................................................................. 138 Underground power distribution .................................................................................... 138

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17. RECOVERY METHODS ............................................................................................................... 139

Mineral processing description and recovery ...................................................................... 140 Process description ........................................................................................................... 141 Expected recovery ............................................................................................................. 142

18. PROJECT INFRASTRUCTURE ................................................................................................... 144

Surface Water Management ................................................................................................... 144 Overburden, waste and ore pads ...................................................................................... 144 Mine dewatering water ...................................................................................................... 144 Water treatment plant ........................................................................................................ 144

Tailings Storage Facility ......................................................................................................... 144 Access Road ............................................................................................................................ 145

Main access road............................................................................................................... 145 Site access roads .............................................................................................................. 145

Garage ...................................................................................................................................... 145 Portal and Underground Mine Surface Equipment .............................................................. 145 Explosives Storage ................................................................................................................. 146 Administrative Building and Dry Complex ........................................................................... 146 Electrical Distribution ............................................................................................................. 147 Existing Infrastructure in the O'Brien Area........................................................................... 147

19. MARKET STUDIES AND CONTRACTS ...................................................................................... 148

Market Studies ......................................................................................................................... 148 Metal Pricing ............................................................................................................................ 148

20. ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT ......... 149

Previous Work on the Property .............................................................................................. 149 Former O’Brien mine ......................................................................................................... 149 Former Kewagama mine ................................................................................................... 150 Liability of Radisson regarding former mine sites .............................................................. 150

Environmental Site Description and Characterization ........................................................ 150 Physical environment ........................................................................................................ 151 Biological environment ...................................................................................................... 151

Management of Waste Rock, Tailings, Ore and Water ........................................................ 151 Chemical characteristics of waste rock and ore ................................................................ 152 Tailings characteristics ...................................................................................................... 153 Run-off water management ............................................................................................... 153

Permitting Requirements ........................................................................................................ 153 Social or Community Impact .................................................................................................. 154 Mine Closure and Rehabilitation ............................................................................................ 155

21. CAPITAL AND OPERATING COSTS .......................................................................................... 156

Capital Costs ............................................................................................................................ 156 Capitalized operating costs ............................................................................................... 157 Capitalized revenue ........................................................................................................... 157 Royalties ............................................................................................................................ 157 Development costs ............................................................................................................ 157 Mobile equipment .............................................................................................................. 158 Surface infrastructure ........................................................................................................ 158

21.1.6.1 Site preparation and installation ................................................................................................... 158 21.1.6.2 Buildings ...................................................................................................................................... 159 21.1.6.3 Water management and distribution ............................................................................................ 160

Mine service infrastructure ................................................................................................ 160 EPCM cost ......................................................................................................................... 161 Closure Costs ................................................................................................................. 161

Salvage value ................................................................................................................ 161

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Operating Costs ....................................................................................................................... 161 Definition drilling ................................................................................................................ 162 Stope development ............................................................................................................ 162 Contractor indirect costs .................................................................................................... 162 Mining costs ....................................................................................................................... 162 O’Brien staff and general costs ......................................................................................... 163 Energy ............................................................................................................................... 164 Milling and transportation .................................................................................................. 164 Environment ....................................................................................................................... 165 Capitalized operating costs ............................................................................................... 165

Taxes and royalties ....................................................................................................... 165

22. ECONOMIC ANALYSIS ................................................................................................................ 166

Financial Analysis ................................................................................................................... 166 Sensitivity Analysis ................................................................................................................. 170

23. ADJACENT PROPERTIES ........................................................................................................... 175

Agnico-Eagle Mines Ltd Property .......................................................................................... 175 New Alger Property ................................................................................................................. 177 Ironwood Project ..................................................................................................................... 177 Comments on Item 23 ............................................................................................................. 178

24. OTHER RELEVANT DATA AND INFORMATION ....................................................................... 179

25. CONCLUSIONS ............................................................................................................................ 180

Mineral Resource Estimate ..................................................................................................... 180 Metallurgy and Milling ............................................................................................................. 181 Environment ............................................................................................................................. 183 Capital and operating cost ..................................................................................................... 183 Mining Plan............................................................................................................................... 185 Financial analysis .................................................................................................................... 186 Risks and Opportunities ......................................................................................................... 188

26. RECOMMENDATIONS ................................................................................................................. 191

27. REFERENCES .............................................................................................................................. 194

LIST OF FIGURES Figure 4.1 – Location of the O’Brien Project in the Province of Québec ..................................................... 46 Figure 4.2 – Location map showing mining titles constituting the O’Brien Project ...................................... 48 Figure 4.3 – Location map showing historical mining titles constituting the O’Brien Project ...................... 49 Figure 5.1 – Topography and accessibility of the O’Brien Project .............................................................. 53 Figure 7.1 – Stratigraphic map of the Abitibi Greenstone Belt. The geology of the southern Abitibi

Greenstone Belt is based on Ayer et al. (2005) and the Québec portion on Goutier and Melançon (2007). Figure modified from Thurston et al. (2008). ................................. 75

Figure 7.2 – Geological syntheses of the Cadillac mining camp with location of active and closed mines, ore deposits and showings. Modified from Lafrance et al. (2003a, 2003b) ........... 78

Figure 8.1 – Inferred crustal levels of gold deposition showing the different types of lode gold deposits and the inferred deposit clan (from Dubé et al., 2001; Poulsen et al., 2000) .................... 84

Figure 8.2 – Schematic diagram illustrating the setting of greenstone-hosted quartz-carbonate vein deposits (from Poulsen et al., 2000) .................................................................................. 85

Figure 9.1 – 3D view looking NNE, showing the different areas defined by Richard and Fallara (2015). .. 88 Figure 12.1 – Photo of the logging facility building taken during a site visit in January 2015 ..................... 92 Figure 12.2 – Photo of the indoor core storage facilities taken during a site visit in January 2015 ............ 93

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Figure 12.3 – Photo of the outdoor core storage facilities taken during an earlier site visit by InnovExplo in 2014 ............................................................................................................ 93

Figure 12.4 – Photo of the sample preparation facility taken during a site visit in January 2015 ............... 94 Figure 13.1 – Darius mill flowsheet ............................................................................................................. 96 Figure 14.1 – Surface plan view of the O’Brien drill hole database. Top: All drill holes in the database

(n = 2,125); Bottom: validated holes in the 36E and Kewagama areas used for the 2015 resource estimate (n = 620) .................................................................................... 104

Figure 14.2 – 3D view looking northeast of the 55 mineralized solids. ..................................................... 105 Figure 14.3 – 3D view looking northeast of the underground workings in the 36E and Kewagama

areas in relation to resource blocks (red). Note that the compilation of the underground workings to the west (old O’Brien mine) is incomplete. ............................. 107

Figure 14.4 – Different graphs supporting a capping of 30 g/t Au for the mineralized zones east of the fault. ................................................................................................................................. 109

Figure 14.5 – Different graphs supporting a capping of 30 g/t Au for the mineralized zones west of the fault. ................................................................................................................................. 110

Figure 14.6 – Different graphs supporting a capping of 4 g/t Au for the dilution envelope east of the fault. ................................................................................................................................. 111

Figure 14.7 – Different graphs supporting a capping of 3.5 g/t Au for the dilution envelope west of the fault. ................................................................................................................................. 112

Figure 14.8 – 3D view looking north-northeast showing Zone 101, all drill holes and the ellipsoid of Pass 1 (50m x 25m x 12.5m). .......................................................................................... 119

Figure 14.9 – 3D view looking north-northeast showing Zone 101, all drill holes and the ellipsoid of Pass 2 (100m x 50m x 25m). ........................................................................................... 119

Figure 14.10 – Longitudinal view looking north showing all interpolated blocks of Zone 101 with respective categorization. ................................................................................................ 122

Figure 14.11 – Longitudinal view looking north showing all interpolated blocks of Zone 222 with respective categorization. ................................................................................................ 123

Figure 14.12 – 3D view looking northeast showing all indicated blocks above the cut-off grade of 3.50 g/t Au. ....................................................................................................................... 124

Figure 14.13 – 3D view looking northeast showing all indicated blocks above the cut-off grade of 3.50 g/t Au among drill holes and historical underground workings. ............................... 124

Figure 14.14 – Graph showing variations of gold prices in $US, the CAD: USD exchange rate, and the resultant gold price in $C. The dashed line presents the values used to determine the cut-off grade for the resource estimate presented in this report (roughly averages of the previous six months). ................................................................................................. 126

Figure 16.1 – Longitudinal view of the modified Avoca mining method: drilling, blasting and mucking activities. .......................................................................................................................... 132

Figure 16.2 – O’Brien mine development and stopes ............................................................................... 136 Figure 17.1 – Typical gravity / CIP flowsheet ............................................................................................ 142 Figure 22.1 – Sensitivity analysis of economic parameters on after-tax NPV at 5% (millions $) .............. 171 Figure 22.2 – Sensitivity analysis of grade on after-tax NPV at 5% (millions $) ....................................... 172 Figure 22.3 – Sensitivity analysis of economic parameters on after-tax IRR ............................................ 173 Figure 22.4 – Sensitivity analysis of grade on after-tax IRR ..................................................................... 174 Figure 23.1 – Adjacent properties of the O’Brien Project, showing past and current producers. ............. 176 LIST OF TABLES Table 6.1 – Total mine workings at the O’Brien mine from 1926 to 1957 ................................................... 59 Table 6.2 – Total gold production of the O’Brien mine from 1926 to 1957 ................................................. 60 Table 6.3 ─ Total gold production from the O’Brien mine from 1974 to 1981 ............................................. 62 Table 6.4 ─ Holes drilled by Radisson between 1995-2013 ....................................................................... 69 Table 6.5 ─ Total of holes drilled by Radisson from 2003 to 2011 ............................................................. 73 Table 6.6 ─ Best results obtained from Radisson’s drilling campaigns ...................................................... 73 Table 13.1 – Ore processed between 1979 and 1982 ................................................................................ 97

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Table 13.2 – Summary of gold recoveries based on laboratory results for each extraction method tested ................................................................................................................................. 98

Table 13.3 – Calculated gold head grade ................................................................................................... 99 Table 13.4 – Gravity recoveries .................................................................................................................. 99 Table 13.5 – Flotation recovery and grind size ......................................................................................... 100 Table 13.6 – Summary of gravity and flotation gold recoveries ................................................................ 100 Table 13.7 – Arsenic mass balance .......................................................................................................... 101 Table 13.8 – Summary of cyclic tests ........................................................................................................ 101 Table 13.9 – Summary of gravity and cyanidation test results .................................................................. 102 Table 14.1 – Summary statistics for the raw assays by dataset ............................................................... 108 Table 14.2 – Summary statistics for the composites ................................................................................. 114 Table 14.3 – Summary statistics for the composites ................................................................................. 115 Table 14.4 – Block model properties ......................................................................................................... 115 Table 14.5 – Block model .......................................................................................................................... 117 Table 14.6 – Input parameters used for the underground cut-off grade estimation .................................. 125 Table 14.7 – O’Brien Project Mineral Resource Estimate at a 3.50 g/t Au cut-off (O’Brien and

Kewagama claim blocks) and sensitivity at other cut-off scenarios ................................ 127 Table 16.1 - Resources considered in the mining plan (cut-off 3.5 g/t) .................................................... 129 Table 16.2 – Cut-off grade parameters (CAD) .......................................................................................... 130 Table 16.3 – Bolt length as a function of span .......................................................................................... 131 Table 16.4 – Mine plan tonnage distribution ............................................................................................. 134 Table 16.5 – O’Brien mine development quantities .................................................................................. 135 Table 16.6 – O’Brien mine production rates .............................................................................................. 135 Table 16.7 – Radisson mining staff ........................................................................................................... 137 Table 17.1 - Potential plants for custom milling......................................................................................... 139 Table 17.2 – Trade-off study ..................................................................................................................... 140 Table 17.3 – Recoveries obtained in laboratory ........................................................................................ 142 Table 17.4 – Expected gold recovery ........................................................................................................ 143 Table 21.1 – Capital cost estimate ............................................................................................................ 156 Table 21.2 – Capitalized operating costs .................................................................................................. 157 Table 21.3 - Development costs ................................................................................................................ 158 Table 21.4 – Surface infrastructure costs .................................................................................................. 158 Table 21.5 – Site preparation and installation ........................................................................................... 159 Table 21.6 – Buildings ............................................................................................................................... 160 Table 21.7 – Water management and distribution .................................................................................... 160 Table 21.8 – Mine service infrastructure costs .......................................................................................... 161 Table 21.9 – Summary of operating costs ................................................................................................. 162 Table 21.10 – Mining costs........................................................................................................................ 163 Table 21.11 – O’Brien staff salaries .......................................................................................................... 163 Table 21.12 – Annual energy cost (average for Years 3-6) ...................................................................... 164 Table 21.13 – Annual environmental cost ................................................................................................. 165 Table 22.1 – Cash flow analysis summary ................................................................................................ 167 Table 22.2 – Economic analysis for the O’Brien Project (figures in Canadian dollars) ............................. 169 Table 22.3 – Sensitivity analysis of economic parameters on after-tax NPV at 5% (millions $) ............... 171 Table 22.4 – Sensitivity analysis of grade and Gold Price on after-tax NPV at 5% (millions $) ............... 172 Table 22.5 – Sensitivity analysis of economic parameters on after-tax IRR ............................................. 173 Table 22.6 – Sensitivity analysis of grade and Gold Price on after-tax IRR ............................................. 174 Table 25.1 – Risks of the O’Brien Project ................................................................................................. 189 Table 25.2 – Opportunities of the O’Brien Project ..................................................................................... 190 Table 26.1 – Estimated costs for the recommended work program ......................................................... 193

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LIST OF APPENDICES APPENDIX I – UNITS, CONVERSION FACTOR, ABBREVIATION......................................................... 202 APPENDIX II – MINING RIGHTS IN THE PROVINCE OF QUÉBEC....................................................... 204 APPENDIX III – DETAILED LIST OF MINING TITLES ............................................................................. 208 APPENDIX IV – DETAILED LIST OF HISTORICAL MINING TITLES...................................................... 210 APPENDIX V – SURFACE PLANS ........................................................................................................... 212 APPENDIX VI – ENVIRONMENTAL CHARACTERIZATION ................................................................... 215

SIGNATURE PAGE – INNOVEXPLO

TECHNICAL REPORT FOR THE O’BRIEN PROJECT, ABITIBI, QUÉBEC


Prepared for



Original signed and sealed (“Sylvie Poirier”)

Signed at Longueuil on January 29, 2016

Sylvie Poirier, Eng. InnovExplo – Consulting Firm Longueuil (Québec)

SIGNATURE PAGE – INNOVEXPLO



Prepared for



Original signed and sealed (“Pierre-Luc Richard”)

Signed at Val-d’Or on January 29, 2016

Pierre-Luc Richard, M.Sc., P.Geo.. InnovExplo – Consulting Firm Val-d’Or (Québec)

Original signed and sealed (“Bruno Turcotte”)


Bruno Turcotte, P.Geo. InnovExplo – Consulting Firm Val-d’Or (Québec)

Original signed and sealed (“Laurent Roy”)


Laurent Roy, Eng. InnovExplo – Consulting Firm Val-d’Or (Québec)

SIGNATURE PAGE – WSP CANADA INC.



Prepared for



Original signed and sealed (“Annie Lavoie”)

Signed at Montreal on January 29, 2016

Annie Lavoie. Eng. WSP Canada inc. Montréal (Québec)




Prepared for



Original signed and sealed (“Éric Poirier”)


Éric Poirier, Eng. WSP Canada inc. Val-d’Or (Québec)




Prepared for



Original signed and sealed (“Marie-Claude Dion St-Pierre”)

Signed at Québec on January 29, 2016

Marie-Claude Dion St-Pierre, Eng WSP Canada inc. Quebec (Québec)

SIGNATURE PAGE – LAMONT INC.



Prepared for



Original signed and sealed (“Ann Lamontagne”)

Signed at Lac-Beauport on January 29, 2016

Ann Lamontagne, Eng., Ph.D. Lamont inc. Lac-Beauport (Québec)

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CERTIFICATE OF AUTHOR – SYLVIE POIRIER I, Sylvie Poirier, Eng (OIQ no.112196; PEO no.100156918) do hereby certify that:

1. I am a Consulting Engineer of: InnovExplo, 560, 3e Avenue, Val-d’Or, Québec, Canada, J9P 1S4.

2. I graduated with a Bachelor’s degree in mining Engineering from École Polytechnique (Montréal, Québec) in 1993.

3. I am a member of the Ordre des Ingénieurs du Québec (OIQ, no. 112196), the Professional Engineers of Ontario (PEO no. 100156918), and the Canadian Institute of Mines (145365).

4. I have worked as an engineer for a total of twenty (20) years since graduating from university. My mining expertise was acquired while working for Lafarge Canada and for Placer Dome and McWatters at the Sigma mine, as well as for Natural Resources Canada on a special research initiative program on narrow-vein mining. I have been a consulting engineer for InnovExplo Inc. since September 2008.

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

6. I am co-author of and also shares responsibility for sections 1, 2, 3, 21, 22, 24, and 25 to 27 of the report titled “Technical Report and Preliminary Economic Assessment for the O’Brien Project, Abitibi, Québec (according to Regulation 43-101 and Form 43-101F1) (the “Technical Report”), effective as of November 29, 2015 and dated January 29, 2016, prepared for Radisson Mining Resources Inc. I supervised the assembly of the report.

7. I have not visited the O’Brien project.

8. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

9. I am independent of the issuer applying all of the tests in Section 1.5 of Regulation 43-101 (National Instrument 43-101).

10. I have read Regulation 43-101 respecting standards of disclosure for mineral projects and Form 43-101F1, and the sections of the Technical Report for which I was responsible have been prepared in accordance with that regulation and form.

Signed on this 29th day of January 2016 (Original signed and sealed)

Sylvie Poirier, Eng. InnovExplo Inc.

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CERTIFICATE OF AUTHOR – PIERRE-LUC RICHARD I, Pierre-Luc Richard, M.Sc., P.Geo. (OGQ licence No. 1119, APGO licence No. 1714), do hereby certify that:

1. I am employed as a geologist by and carried out this assignment for InnovExplo Inc. – Consulting Firm in Mines and Exploration, 560, 3e Avenue, Val-d’Or, Québec, Canada, J9P 1S4.

2. I graduated with a Bachelor’s degree in geology from the Université du Québec à Montreal (Montreal, Québec) in 2004. In addition, I obtained an M.Sc. from the Université du Québec à Chicoutimi (Chicoutimi, Québec) in 2012.

3. I am a member in good standing of the Ordre des Géologues du Québec (OGQ licence No. 1119) and of the Association of Professional Geoscientists of Ontario (APGO licence No. 1714).

4. I have worked in the mining industry for more than 10 years. My exploration expertise has been acquired with Richmont Mines Inc., the Ministry of Natural Resources of Québec (Geology Branch), and numerous exploration companies through InnovExplo. My mining expertise was acquired at the Beaufor mine and several other producers through InnovExplo. I managed numerous technical reports, resource estimates and audits. I have been a geological consultant for InnovExplo Inc. since February 2007.

5. I have read the definition of "qualified person" set out in Regulation 43-101 / National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI 43-101.

6. I am responsible for the mineral resource estimate and responsible for and author of sections 12 and 14 and I am co-author of and also shares responsibility for sections 1, 7, 25 to 27 of the report titled “Technical Report and Preliminary Economic Assessment for the O’Brien Project, Abitibi, Québec (according to Regulation 43-101 and Form 43-101F1) (the “Technical Report”), effective as of November 29, 2015 and dated January 29, 2016, prepared for Radisson Mining Resources Inc. I visited the property on January 19 and January 27, 2015.

7. I have not had any other prior involvement with the project that is the subject of the Technical Report other than being an independent author of a previous 43-101 Report on the project (2015) and working on a target generation mandate (2015).

8. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Report, the omission of which would make the Technical Report misleading.

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

10. I have read NI 43-101 Respecting Standards of Disclosure for Mineral projects and Form 43-101F1, and the items for which I am a qualified person in this Technical Report have been prepared in accordance with that regulation and form.

Dated this 29th day of January, 2016. (Original signed and sealed)

Pierre-Luc Richard, PGeo, MSc InnovExplo Inc

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CERTIFICATE OF AUTHOR – BRUNO TURCOTTE I, Bruno Turcotte, P.Geo. (APGO licence No. 2136, OGQ licence No. 453), do hereby certify that:

1. I am employed as a geologist by and carried out this assignment for InnovExplo Inc. – Consulting Firm in Mines and Exploration, 560, 3e Avenue, Val-d’Or, Québec, Canada, J9P 1S4.

2. I graduated with a Bachelor of Geology degree from Université Laval in the city of Québec in 1995. In addition, I obtained a Master’s degree in Earth Sciences from Université Laval in the city of Québec in 1999.

3. I am a member of the Ordre des Géologues du Québec (OGQ licence No. 453) and of the Association of Professional Geoscientists of Ontario (APGO licence No. 2136).

4. I have worked as a geologist for a total of 20 years since graduating from university. I acquired my exploration expertise with Noranda Exploration Inc., Breakwater Resources Ltd, South-Malartic Exploration Inc. and Richmont Mines Inc. I acquired my mining expertise on the Croinor Preproduction Project and at the Beaufor mine. I have been a geological consultant for InnovExplo Inc. since March 2007.

5. I have read the definition of "qualified person" set out in Regulation 43-101 / National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI 43-101.

6. I am responsible for sections 4 to 6, 8 to 11, 23 and 25 to 27 and I am co-author of and also shares responsibility for sections 1, 7, 25 to 27 of the technical report entitled “Technical Report and Preliminary Economic Assessment for the O’Brien Project, Abitibi, Québec (according to Regulation 43-101 and Form 43-101F1) (the “Technical Report”), effective as of November 29, 2015 and dated January 29, 2016, prepared for Radisson Mining Resources Inc.

7. I have not had any prior involvement with the project that is the subject of the Technical Report. I have not visited the O’Brien Project.

8. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission of which would make the Technical Report misleading.

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

10. I have read NI 43-101 Respecting Standards of Disclosure for Mineral projects and Form 43-101F1, and the Technical Report has been prepared in accordance with that instrument and form.

Signed on this 29 day of January 2016 (Original signed and sealed) Bruno Turcotte, PGeo, MSc InnovExplo Inc

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CERTIFICATE OF AUTHOR – LAURENT ROY I, Laurent Roy, Eng. (OIQ no.109779) do hereby certify that:

1. I am a Consulting Engineer of: InnovExplo Inc., 560 3e Avenue, Val-d’Or, Québec, Canada, J9P 1S4.

2. I graduated with a Bachelor’s degree in mining Engineering from École Polytechnique (Montréal, Québec) in 1992.

3. I am a member of the Ordre des Ingénieurs du Québec (OIQ, no. 109779).

4. I have worked as an engineer for a total of twenty-two (22) years since graduating from university. My mining expertise was acquired while working for Talpa Mining Contractor, Richmont Mines at Francoeur and Beaufor mines, Doyon-Westwood and CasaBerardi mines. I have been a consulting engineer for InnovExplo Inc. since September 2012.

5. I have read the definition of “qualified person” set out in Regulation 43-101/NI43-101 and certify that by reason of my education, affiliation with a professional association (as defined in Regulation 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of Regulation 43-101.

6. I am responsible for and author of Section 16 and I am co-author of and also shares responsibility for sections 21, 22, and 25 to 27 of the report titled “Technical Report and Preliminary Economic Assessment for the O’Brien Project, Abitibi, Québec (according to Regulation 43-101 and Form 43-101F1) (the “Technical Report”), effective as of November 29, 2015 and dated January 29, 2016, prepared for Radisson Mining Resources Inc.

7. I had prior involvement with the property that is the subject of the Technical Report.

8. I visited the property for the purpose of this report however, I visited the O’Brien Project site on September 9, 2014, accompanied by Yolande Bisson of O’Brien Project and Éric Caron of InnovExplo.

9. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which would make the Technical Report misleading.

10. I am independent of the issuer applying all of the tests in Section 1.5 of Regulation 43-101 or National Instrument 43-101.

11. I have read Regulation 43-101 respecting standards of disclosure for mineral projects and Form 43-101F1, and the sections of the Technical Report, for which I was responsible, have been prepared in accordance with that regulation and form.

Signed on this 29th day of January, 2016 (Original signed and sealed)

Laurent Roy, Eng. InnovExplo Inc.

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CERTIFICATE OF AUTHOR – ANNIE LAVOIE I, Annie Lavoie, Eng. (OIQ #124421), do hereby certify that:

1. I am employed as a consulting metallurgical engineer, and carried out this assignment for, WSP Canada Inc., 1600, boul. René Levesque Ouest, Montréal, Québec, Canada, H3H 1P9

2. I graduated with a Bachelor’s degree in Material and Metallurgical Engineering (B.Eng.) from Université Laval (Sainte-Foy, Québec) in 2000.

3. I am a member of the Ordre des Ingénieurs du Québec (OIQ #124421). 4. I have over ten (10) years of experience as a metallurgical engineer in the metallurgical and

mineralogical industry. My expertise has been acquired with Noranda Copper, Falconbridge, Xstrata and Osisko. I have been a consulting metallurgical engineer for WSP Canada since January 2012.

5. I have read the definition of “qualified person” set out in Regulation 43-101 (“R 43-101”) standards for disclosure for mineral projects and certify that by reason of my education, affiliation with a professional association (as defined in R 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of R 43-101.

6. I am responsible for the preparation of Mineral and Metallurgical section (sections 13 and 17) and I am co-author of and also shares responsibility for sections 1 and 25 to 27 of the Technical Report entitled “Technical Report and Preliminary Economic Assessment for the O’Brien Project, Abitibi, Québec (according to Regulation 43-101 and Form 43-101F1) (the “Technical Report”), effective as of November 29, 2015 and dated January 29, 2016, prepared for Radisson Mining Resources Inc. I supervised the assembly of the report.






Annie Lavoie, Eng. WSP

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CERTIFICATE OF AUTHOR – ÉRIC POIRIER I, Eric Poirier, Eng (OIQ no.120063) do hereby certify that:

1. I am a consulting engineer with WSP Canada inc., 1075, 3rd Avenue, Val-d’Or, Quebec, Canada, J9P 0J7.

2. I graduated with Bachelor’s degrees in Electrical Engineering and Computer Science Engineering from Université du Québec à Chicoutimi (Chicoutimi, Québec) in 1996 and 1997.

3. I am a member of the Ordre des Ingénieurs du Québec (OIQ, no. 120063), the Professional Engineers of Ontario (PEO no. 100112909), the Association of Professional Engineers and Geoscientists of the Province of Manitoba (APEGM no. 33233) and the Northwest Territories and Nunavut Association of Professional Engineers and Geoscientists (NAPEG no.L2229).

4. I have worked as an electrical engineer and project manager for a total of eighteen (18) years since graduating from university. My technical expertise includes electrical distribution, cost estimation, automation and instrumentation. I have been involved in many scoping studies and feasibility studies. I have participated in worldwide projects as electrical designer or as multidisciplinary project manager. I have been a consulting engineer for WSP Canada Inc. since January 1998.


6. I am responsible for sections 18.3 to 18.8 and I am co-author of and also shares responsibility for sections 1, 21, 25 to 27 of the report titled “Technical Report and Preliminary Economic Assessment for the O’Brien Project, Abitibi, Québec (according to Regulation 43-101 and Form 43-101F1) (the “Technical Report”), effective as of November 29, 2015 and dated January 29, 2016, prepared for Radisson Mining Resources Inc.

7. I have visited the O’Brien and Kewagama sites on July 22, 2014 and October 5, 2015.





Éric Poirier, Eng. WSP Canada inc.

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CERTIFICATE OF AUTHOR – MARIE-CLAUDE DION ST-PIERRE I, Marie-Claude Dion St-Pierre, Eng. M.A.Sc. (OIQ no. 140947) do hereby certify that:

1. I am a Project Manager with WSP Canada inc. with a business address at 5355 boul des Gradins, Québec, Québec, Canada, G2J 1C8.

2. I am a graduate of Sherbrooke University, (Bachelor’s degree in Chemical Engineering, 2004 and Master’s degree in Chemical Engineering, 2007).

3. I am a member of the Ordre des Ingénieurs du Québec (OIQ, no. 140947).

4. I have worked as an engineer for a total of nine (9) years since obtaining my Bachelor’s degree. My mining expertise was acquired while working for GENIVAR and Les mines Agnico-Eagle Ltée. I have been a consulting engineer for WSP Canada Inc. since January 2014.


6. I am responsible for the section 18.1 and 18.2 and I am co-author of and also shares responsibility for sections 1, 21 and 25 to 27 of the report titled “Technical Report and Preliminary Economic Assessment for the O’Brien Project, Abitibi, Québec (according to Regulation 43-101 and Form 43-101F1) (the “Technical Report”), effective as of November 29, 2015 and dated January 29, 2016, prepared for Radisson Mining Resources Inc.






Marie-Claude Dion St-Pierre, Eng. M.A.Sc. WSP Canada inc.

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CERTIFICATE OF AUTHOR – ANN LAMONTAGNE I, Ann Lamontagne, Eng., Ph.D. (OIQ no.104345) do hereby certify that:

1. I am a president of Lamont with an office at 10 chemin des Conifères, Lac-Beauport, Québec, G3B 2E7.

2. I graduated with a Bachelor’s degree in civil Engineering from Laval University (Québec, Québec) in 1990.

3. I am a registered member of the Order of Engineers of Quebec (OEQ, no. 104345).

4. I have worked as a civil engineer continuously since my graduation from university.


6. I am responsible for the section 20 and share responsibility for sections 1, and 25 to 27 of the report entitled “Technical Report and Preliminary Economic Assessment for the O’Brien Project, Abitibi, Québec (according to Regulation 43-101 and Form 43-101F1) (the “Technical Report”), effective as of November 29, 2015 and dated January 29, 2016, prepared for Radisson Mining Resources Inc.

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

8. I have not visited the site.

9. I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this report.

10. Neither I, nor any affiliated entity of mine, is at present under an agreement, arrangement or understanding or expects to become an insider, associate, affiliated entity or employee of Radisson Mining Resources, or any associated or affiliated entities.

Signed on this 29 day of January 2016 (Original signed and sealed)

Ann Lamontagne, Eng., Ph.D. Lamont inc.

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1. SUMMARY

Introduction At the request of Radisson Mining Resources Inc. (“Radisson” or the “issuer”), InnovExplo Inc. (“InnovExplo”) was retained to prepare a Preliminary Economic Assessment (the “PEA”) and Technical Report (the “report”) for the O’Brien Project (the ‘’Project’’), in accordance with National Instrument 43-101 Respecting Standards of Disclosure for Mineral Projects (“NI 43-101”) and its related form 43-101F1. The PEA was prepared with contributions from WSP Canada Inc. (“WSP”) and Lamont inc (“Lamont”) InnovExplo is an independent mining and exploration consulting firm based in Val-d’Or, Québec. WSP is a professional services firm that operates in different market sectors: property and buildings, transport and infrastructure, environment, industry, mining, oil and gas, and power and energy. Lamont is a professional services firm covering all aspects concerning the design of surface infrastructure (water management, tailings management, access, etc.) and the environment (site characterization, soil and water samplings). Lamont also covers the process associated with the project authorizations from both provincial and federal governments. This Technical Report supports the disclosure of the PEA results covering the Kewagama and 36E areas of the O’Brien Project, near the town of Cadillac in the Province of Québec.

Property Description and Location The O’Brien Project is located in the province of Québec, Canada, just north of the municipality of Cadillac, within the new limits of the city of Rouyn-Noranda. Cadillac lies approximately 45 km east of downtown Rouyn-Noranda, and 45 km west of downtown Val-d’Or. The current O’Brien Project represents the amalgamation of the O’Brien and Kewagama properties. Both properties are 100% owned by Radisson. The O’Brien Project consists of a contiguous block comprising 36 mining claims covering an aggregate area of 636.6 hectares. On July 30, 2015, all mining titles constituting the O’Brien Project were converted into “map-designated claims”. Consequently, the O’Brien Project now consists of one contiguous block comprising 21 mining claims staked by electronic map designation (map-designated claims) covering an aggregate area of 637.09 hectares The O’Brien property included a mining lease that expired in 2008 and was subsequently converted back into claims. A payment of $1,000,000 must be made to Breakwater Resources Ltd (now Nyrstar) upon commencement of commercial production on either one of the O’Brien or Kewagama properties, against which shall be deducted any costs required to restore the O’Brien tailing ponds. Additionally, there is a 2% NSR royalty payable to KWG Resources Inc. in the event of commercial production on the Kewagama property.

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Geological Setting and Mineralization The property is underlain by rocks of the Southern Volcanic Zone of the Abitibi Subprovince, intruded by Proterozoic diabase dykes. The Cadillac–Larder Lake Fault Zone (CLLFZ) runs along an E-W axis and separates the metasedimentary Pontiac Subprovince to the south from the volcano-sedimentary Abitibi Subprovince to the north. In Québec, about forty or so gold deposits, which have produced over 60 million ounces of gold since the early 20th century, are associated with this major structure and its subsidiary faults. The O’Brien Project straddles the Piché Group volcanic rocks that separate the Pontiac Group metasedimentary rocks to the south from the Cadillac Group metasedimentary rocks to the north. In the property area, all lithologies strike E-W and dip steeply south at approximately 85°. The CLLFZ is a major regional crustal break that consists mainly of chlorite-talc-carbonate ultramafic schist and ranges in thickness from 30 to 100 m (100 to 300 ft) in the mine area, and narrows significantly to about 12 m (40 ft) wide to the east of Zone 36 East. Across the property, the fault is subparallel and passes near the Piché Group-Cadillac Group contact, but is generally enveloped by Cadillac Group sedimentary rocks comprising argillites, greywackes and, to a lesser extent, chert. Gold production from the former O’Brien mine came from a few quartz veins running almost parallel to the formations. The mine’s productive sector was generally limited to a narrow strip that included the O’Brien conglomerate and the northern porphyritic andesite. Approximately 95% of the O’Brien ore came from four veins (No. 1, No. 4, No. 9 or “F”, and No. 14) in the eastern part of the mine. The veins contained high-grade shoots that occasionally yielded considerable amounts of visible gold. The main veins generally strike from 083° to 098°, and dip steeply to the south (-84° to -90°). The stopes averaged 0.75 to 0.90 m (2.5 to 3 ft) wide. Gold mineralization extends vertically down to at least the 3450' level.

Data Verification InnovExplo’s data verification included visits to the project’s office, as well as to the logging and core storage facilities. It also included a review of selected core intervals, drill hole collar locations, assays, the QA/QC program, downhole surveys, information on mined-out areas, and the descriptions of lithologies, alterations and structures. Site visits were completed by Pierre-Luc Richard of InnovExplo on January 19 and January 27, 2015. Laurent Roy of InnovExplo visited the property on April 9 and 10, 2015, accompanied by Jean Garant of InnovExplo. In mid-December 2015, Radisson began a surface diamond drilling program at the O’Brien Project. The drilling program is still ongoing. To date, no gold results have been reported by Radisson from this drilling program.

Mineral Resource Estimates The 2015 O’Brien Mineral Resource Estimate herein was prepared by Pierre-Luc Richard, P.Geo., with contributions from Alain Carrier, M.Sc., P.Geo., using all available information. The main objective of the mandate assigned by Radisson was

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to update the 2013 Mineral Resource Estimate prepared by RPA, which was published as an NI 43 101 compliant report titled “Technical Report on the O’Brien Project Mineral Resource Estimate, Québec, Canada” (de l’Étoile and Salmon, 2013). The 2013 resource estimate focused solely on the 36E area while the herein presented resource estimate also includes the Kewagama area. The Kewagama area was mined in the past as the Kewagama mine, and the 36E area was partially mined from extensions of the Kewagama and O’Brien mines. The 2015 resource area measures 2.1 km along strike, 0.6 km wide and 0.7 km deep. The resource estimate is based on a compilation of historical and recent diamond drill holes and a litho-structural model constructed by InnovExplo. The GEMS diamond drill holes database contains 310 surface diamond drill holes and 1,815 underground drill holes. From these, a subset of 620 holes (279 from surface and 341 from underground) located inside the limits of the resource estimate area were used for the 2015 resource estimate, representing all the holes that had been compiled and validated at the time this estimate was being initiated. In order to conduct accurate resource modelling of the deposit, InnovExplo based its mineralized-zone wireframe model on the drill hole database and the authors’ knowledge of the O’Brien mine. A total of 4,215 construction lines (1,372 3D rings and 2,843 tie lines) were created in order to produce valid solids. A total of 55 mineralized solids (coded 101 to 230) that honour the drill hole database were also created. Given the density of the processed data, the search ellipse criteria, the drill hole density, and the specific interpolation parameters, InnovExplo is of the opinion that the current internal mineral resource estimate can be classified as Indicated and Inferred resources. The estimate is compliant with CIM standards and guidelines for reporting mineral resources and reserves. The following table displays the results of the In Situ Mineral Resource Estimate for the O’Brien Project (55 mineralized zones and 2 dilution envelopes) at the official 3.50 g/t Au cut-off grade (O’Brien and Kewagama claim blocks) and sensitivity at other cut-off scenarios. The reader should be cautioned that the figures listed in the following table, apart from the official scenario at 3.50 g/t Au, should not be misinterpreted as a mineral resource statement. The reported quantities and grade estimates at different cut-off grades are only presented to demonstrate the sensitivity of the resource model to the selection of a reporting cut-off grade.

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2015 O’Brien Project Mineral Resource Estimate at a 3.50 g/t Au cut-off (O’Brien and Kewagama claim blocks) and sensitivity at other cut-off scenarios (Table 14.7)

The Independent and Qualified Persons for the Mineral Resource Estimate, as defined by NI 43-101, are Pierre-Luc Richard, P.Geo., M.Sc. and

Alain Carrier. P.Geo., M.Sc., of InnovExplo Inc., and the effective date of the estimate is April 10, 2015. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. The resource model includes the previously named 36E Zone and former Kewagama mine. The historical O’Brien mine area is not included in this

resource as it had not been compiled or validated at the time this estimate is being prepared. The model includes 55 gold-bearing zones, not all of which include resources at the official cut-off grade. A dilution envelope was also modelled, but no resource at the official cut-off grade is being reported for the envelope.

Results are presented in situ and undiluted. Sensitivity was assessed using cut-off grades of 2.00, 2.50, 3.00, 3.50, 4.00 and 5.00 g/t Au. The official resource is reported at a cut-off of

3.50 g/t Au. The reader is cautioned that the figures presented herein, apart from the official scenario at 3.50 g/t Au, should not be misinterpreted as a mineral resource statement. The reported quantities and grade estimates at different cut-off grades are only presented to demonstrate the sensitivity of the resource model to the selection of a reporting cut-off grade.

Cut-off grades must be re-evaluated in light of prevailing market conditions (gold price, exchange rate and mining cost). A fixed density of 2.67g/cm3 was used for all zones. A minimum true thickness of 1.5 m was applied, using the grade of the adjacent material when assayed, or a value of zero when not assayed. High grade capping (Au) was done on raw assay data and established on a sector basis (Western zones: 65g/t, Eastern zones: 30g/t, Western

dilution zone: 3.5 g/t Eastern dilution zone: 4.0g/t). Compositing was done on drill hole intercepts falling within the mineralized zones (composite = 0.80 m). Resources were evaluated from drill holes using a 2-pass ID2 interpolation method in a block model (block size = 3 m x 3 m x 3 m). The inferred category is only defined within the areas where blocks were interpolated during pass 1 or pass 2. The indicated category is only defined in areas where the maximum distance to the closest drill hole composite is less than 20m for blocks

interpolated in pass 1. Ounce (troy) = metric tons x grade / 31.10348. Calculations used metric units (metres, tonnes and g/t). The number of metric tons was rounded to the nearest hundred. Any discrepancies in the totals are due to rounding effects. Rounding followed

the recommendations in NI 43-101. InnovExplo is not aware of any known environmental, permitting, legal, title-related, taxation, socio-political, marketing or other relevant issue that

could materially affect the Mineral Resource Estimate.

Zone Cut-off Tonnage Grade Ounces Zone Cut-off Tonnage Grade Ounces

2.00 1,384,700 4.22 188,049 2.00 3,388,500 3.64 396,601

2.50 991,200 5.01 159,770 2.50 2,254,100 4.36 315,725

3.00 748,800 5.75 138,456 3.00 1,525,300 5.12 251,293

3.50 570,800 6.53 119,819 3.50 918,300 6.38 188,466

4.00 444,300 7.33 104,676 4.00 663,500 7.42 158,273

5.00 320,800 8.43 86,939 5.00 486,200 8.52 133,245

Indicated

All

Zones

Inferred

All

Zones

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Metallurgy and Milling There are many historical documents relating to the O’Brien Project area. Several test programs have been carried out since the 1970s. These were executed by various laboratories. The relationship between historical results and the area that is being studied is complex. Most of the time, samples were identified under the name of the zone. However, these names have changed over time, depending on which company owned the deposit. Nevertheless, these data provide an overview of the mineralogy, treatment methods and gold recoveries that may be obtained for samples taken from this area. The O'Brien Project, as currently defined, covers the 36E and Kewagama areas. The 36E area is divided into four zones: Upper West, West Central, West and Lower Central. The Kewagama area covers the eastern sector. In 2014, new laboratory testwork was undertaken on samples from the 36E area by the URSTM. In view of potential mining activities, custom milling will be the preferred option. The recent metallurgical testwork has demonstrated the amenability of O’Brien mineralized material to the gravity, leaching and flotation processes. The O’Brien Project is planned for a five-year period at a production rate of approximately 500 tpd. Five gold concentrators located within a 75-km radius were then identified as being able to potentially process the O’Brien material: the Kiena Mill, the Sigma-Lamaque Complex, the Camflo Mill, the Westwood Mill and the Aurbel Mill. The following Table summarizes the main features of these milling options. Potential plants for custom milling (Table 17.1)

Mill Company Process Capacity Distance Mill status (operating or closed)

Interest for custom milling

Kiena Mill Wesdome Leaching/CIP 1,000 to 2,200 tpd

48 km Closed NA

Sigma-Lamaque Complex

Integra Gold Gravity Concentration & Leaching/CIP

1,200 to 2,400 tpd

67 km Closed No interest

Camflo Mill Richmont Mines

Leaching/Merrill-Crowe

800 to 1,200 tpd

35 km Operating No interest

Westwood Mill

Iamgold Gravity Concentration & Leaching/CIP

Or Gravity Concentration

& Flotation

2,400 tpd

800 tpd

19 km Operating Yes

Aurbel Mill QMX Gravity Concentration & Flotation & Flotation Concentrate Leaching

500 to 800 tpd

75 km Operating Yes with environmental

conditions

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The companies were contacted to find out their interest in performing custom milling. The Westwood and Aurbel mills have shown interest. This PEA is based on the use of the Westwood mill. The proximity of the O'Brien Project helps reduce transportation costs. In addition, the plant gives no restriction for environmental treatment. A trade-off study was conducted to compare treatment costs and potential recoveries for the two flowsheets available at the Westwood mill, see the following table. Trade-off study (Table 17.2)

Gravity/Flotation Gravity/Cyanidation

Gold value

Ore grade1 g/t 6.46 6.46

Recovery % 94.52 91.53

Total3 C$/t 289 280

Milling cost

Preparation and trucking $/t 5.78 5.78

Custom milling $/t 31 31

Smelting $/t 45 NA

Total $/t 81.78 36.78 1 Based on mining plan 2 URSTM test KN-F-3 3 Assumption section 17.1.2

BASED ON Gold PRICE at C$1475 /oz

The smelting cost was estimated based on information from similar projects. Preparation and trucking quotations were obtained from suppliers. The budgetary custom milling cost was estimated by the mill based on current knowledge of the ore. However, prices may be adjusted when additional information becomes available. Westwood's gravity and CIP circuit appears to be a good compromise based on the URSTM metallurgical results and the above considerations. The recovery will be lower but the treatment costs are significantly less. However, further work is required to validate the amount of free gold and the recovery by leaching process and then, determine a specific flowsheet that will optimize themetallurgical performance.

Environment The area where the future mining activities will take place has already been impacted by previous mining activity. The area that is planned for development is adjacent to the previous (removed) infrastructure that was present on the Kewagama mine site. The Project activities should be constrained to an area that is less than 15 hectares. To obtain permits for the project, an environmental baseline study is required. This study will define the receiving environment before project development including the physical, biological and social environmental aspects. For this type of project, the

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study area should cover the location of the infrastructure within the project area that covers at least 15 hectares. For this project, no Environmental Impact Assessment will be required, as the Project remains lower than 2000 tpd (EQA Q-2, r.23) and no Physical Activities (SOR/2012-147) could trigger the Federal Process. Permits will be mainly issued by the “Ministère de l’Énergie et des Ressources Naturelles” and by the “Ministère du Développment durable, de l’Environnement et de la Lutte contre les Changements Climatiques”. From 2012 to 2014, Radisson conducted a geochemical characterization study of ore and waste rock samples. The majority of waste rock samples show no potential for acid generation but results indicate that all ore samples show a potential for acid generation. Samples of waste rock and ore have also been tested for their metal leaching (ML) potential. According with definition of Quebec’s Directive 019 and TCLP results, both waste rock and ore are leachable for some metals. The management of waste rock pile, ore stockpile as well as surface run-off were deisgned accordingly. Mine closure and rehabilitation cost have been estimated at $ 3.6 M. The closure cost estimate is based on capping the waste rock pile with an impermeable cover to limit infiltration and on the re-vegetation of the overburden layer that will cover the waste rock pile.

Mining Plan

The proposed mining plan for the O’Brien Project was prepared using the inferred and indicated resources estimated by InnovExplo. Due to the narrow vein nature of the orebody, two (2) underground mining methods were considered in the study, modified Avoca and long-hole mining with captive sublevels. The mining plan for the O’Brien Project comprises a combination of conventional and mechanized mining. The approach in this study has been to prioritize the modified Avoca mining method when possible. When this approach was not convenient, long-hole mining with captive sublevels was selected. The mineralized material will be transported to surface using a combination of 3.5-cubic-yard to 6-cubic-yard scoop trams and 30-tonne trucks. Waste material will be used to backfill mined out stopes as much as possible or will be brought to surface and stored on a dedicated waste pad. The current PEA is based on an underground mine with access by decline to a vertical depth of 550 metres in the 36E area and 250 metres in the Kewagama area. The production drifts will be accessed via crosscuts connecting to the ramp. A portion of the resources will be mined using captive methods, however haulage will always be mechanized. The mineral resource block model prepared by InnovExplo was used for the PEA. First, the resources available for mining were defined by creating the stope geometry in the block model at a cut-off grade of 3.5 g/t. Then a second triage was done using a diluted cut-off grade of 4.01 g/t. The guideline used in the stope design was a minimum mining width of 1.8 metres for subvertical stopes. The subvertical structures were cut at 18-metre vertical intervals corresponding to access level elevations.

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The conversion of mineral resources to potential mineral reserves takes into account dilution and losses during mining operations. The mineral resources are already diluted to a minimum width of 1.8 metres. Mining recovery was established at 85%, to take into account pillar requirements. A 30% dilution was also taken into account for stope excavation. Finally, a 95% recovery was applied to account for mining operating losses. For stopes with a diluted grade of less than 5.0 g/t, an evaluation was made to determine the economic viability of each stope, considering the development required to access the stope. If the economic viability could not be justified, the stope was discarded. Following this exercise, that included mine dilution and mine recovery a total of 712,521 tonnes at 6.46 g/t (147,986 oz) was included in the mine plan. Mine development will be accelerated in the first two years of the project to provide a degree of flexibility in terms of access, which should facilitate scheduling during the production period. The development sequence will ensure that many stopes are available for mining at a number of different locations at any given time. However, some of the stopes can only be mined at the end of the mine life since they are located directly over or under the level, therefore preventing any further access on that level when mined. The expected average daily production rate during the production period is estimated in this PEA between 450 and 500 t/day. The overall project mine life is expected to be approximately 6 years, including a two-year pre-production period. In the opinion of author Laurent Roy, Eng., the mine plan should be achievable given the flexibility and number of available working places. The following table summarizes the annual tonnage distribution according to the mine plan. Mine plan tonnage distribution (Table 16.4)

Pre-production Production Total

Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Production (t) 33,194 126,494 129,593 134,524 127,259 551,064

Grade (g/t) 7.20 7.05 7.39 5.66 6.53 6.68 Development (t) 3,196 33,474 32,080 40,298 52,409 161,457

Grade (g/t) 7.05 5.74 6.19 5.95 5.11 5.70 Total tonnage milled (t) 3,196 66,668 158,574 169,891 186,933 127,259 712,521

Grade (g/t) 7.05 6.47 6.87 7.04 5.50 6.53 6.46

Capital and operating cost The PEA is based on capital pricing as of the third quarter of 2015. The PES assumes that the development and mining of the mine will be done by contractors and that they will supply the mobile equipment.

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The capitals costs were estimated using the following sources of information:

Quotes from equipment suppliers Comparable installations at other mining projects Contractor costs InnovExplo’s internal database

The capital cost estimates are accurate within ±20%. The preproduction costs are estimated at $36,76M, net of production revenue received during the second year of the preproduction period ($19,11M). Preproduction capital costs are minimal given that there is no need to build processing and tailings facilities. Preproduction is anticipated to take 2 years with the majority of proceeds used for ramp construction and for sufficient development of mineralized zones, or working faces, to conduct mining at the proposed mining rate and mill throughput. Sustaining capital is estimated at $21.35 million, including $3.7 million for final closure costs and considering a salvage value to $1,46M. Capital cost estimate (Table 21.1)

Description Pre-production Sustaining capital Total cost

Capitalized operating costs $21.33 M $21.33 M Capitalized revenue -$19.30 M -$19.30 M Royalty payment $1.00 M $1.00 M Development $20.01 M $17.13 M $37.14 M Mobile equipment $0.21 M $0.18 M $0.39 M Surface infrastructure $6.45 M $0.02 M $6.48 M Mine service infrastructure $7.29 M $0.78 M $8.06 M Closure costs $3.70 M $3.70 M Salvage value -$1.46 M -$1.46 M EPCM $0.77 M $0.77 M Total $36.76 M $21.35 M $58.12 M

Operating costs are estimated in 2015 Canadian dollars with no allowance for escalation. The total operating cost and average unit operating costs are summarized in the following table. The overall unit operating cost is $177.10 per tonne.

Operating costs are summarized below for the production period.

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Summary of total operating costs (Table 21.9)

Description Total cost Unit cost ($/t) ($/oz)

Definition drilling and sampling $2.47 M 3.85 20.29 Stope development $22.09 M 34.38 181.16 Contractor indirect costs $18.06 M 28.11 148.11 Mining costs $27.30 M 42.48 223.84 O'Brien staff and general $12.38 M 19.27 101.53 Energy costs $5.89 M 9.17 48.32 Milling and transportation $23.64 M 36.78 193.80 Environment $1.97 M 3.06 16.14 Total $113.81 M 177.10 933.18

Financial analysis An after-tax model was developed for the O’Brien Project. All costs are in 2015 Canadian dollars with no allowance for inflation or escalation. Income taxes are calculated in accordance with the federal and provincial tax legislations relating to mining companies. The calculations were made by Lucie Chouinard of Raymond Chabot. The federal income tax rate is 15% and the combined provincial income tax rate is 11.9%. Québec mining duties are calculated in accordance with Bill 55, which contains amendments to Québec’s Mining Tax Act and received its first reading in the Québec legislature on November 12, 2013. The Kewagama property consisted of a contiguous block comprising three (3) mining claims covering an aggregate area of 112.07 hectares. Radisson owned a 100% interest in the Kewagama property, with a 2% NSR royalty payable to KWG Resources Inc. in the event of commercial production. In addition, a $1,000,000 payment must be made to Breakwater Resources Ltd (now Nyrstar) upon commencement of commercial production on either one of the O’Brien or Kewagama properties, against which shall be deducted any costs required to restore the O’Brien tailing ponds. In the cash flow analysis, this royalty was considered on all ounces produced from the Kewagama property. The economic evaluation was performed using the Internal Rate of Return (IRR) and the Net Present Value (NPV) methods. This Preliminary Economic Assessment (PEA) is preliminary in nature as it includes Inferred Mineral Resources that are too speculative geologically to have economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the PEA will be realized.

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The following parameters were considered in the financial analysis.

An average gold price of US$1,180 per ounce and an exchange rate of 1.25 CAD/1 USD.

Milling recovery of 91.5%. Refining cost of $3/oz. Royalty payment of 2% NSR payable to KWG Resources Inc. on all ounces

produced from the Kewagama property. A residual fiscal base of $ 5.8M was considered in the tax estimation

regarding previous expenses by Radisson on the O’Brien Project. Resources as presented in Section 14. Future annual cash flow estimates based on grade, gold recoveries and cost

estimates as previously discussed in this Report. 69,864 tonnes of mineralized material to be processed during the pre-

production period, deemed as capital production and not included in production nor the revenue derived from it.

The main parameters and cash flow analysis results for the entire project are presented in the following table. Cash flow analysis summary (Table 22.1)

Parameters Results Current mineral resources included (indicated and inferred) 712,521 tonnes @ 6.46 g/t Au

Mill recovery 91.5% Life of mine ("LOM") (including 24 months of pre-production) 6 years

Daily mine production 440 tpd

Gold recovered over LOM 135,308 oz

Gold price (USD) $1,180

Exchange rate (CAD/USD) 1.25

Gold price (CAD) $1,475

Total gross revenue $199.5M

Pre-production capital cost $36.8M

Average operating cost per tonne $178/tonne

Average operating cost per ounce in US$ US$752/ounce

PRE-TAX

LOM NPV at 5% discount rate (C$) $0.2M

Internal Rate of Return (IRR) 5.18%

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Parameters Results

Payback period (years) 5.6

AFTER-TAX

LOM NPV at 5% discount rate (C$) $(1.9)M

IRR (%) 3.15%


Risks and Opportunities The following table identifies the significant internal risks, potential impacts and possible risk mitigation measures that could affect the economic outcome of the project. The list does not include the external risks that apply to all mining projects (e.g., changes in metal prices, exchange rates, availability of investment capital, change in government regulations, etc.). Significant opportunities that could improve the economics, timing and permitting of the project are identified in the following table. Further information and study is required before these opportunities can be included in the project economics. Risks of the O’Brien Project (Table 25.1) RISK Potential Impact Possible Risk Mitigation

Proximity of the historical O’Brien mine where

environmental, economic, and/or technical potential

issues could arise from the presence of 8,938 barrels of

arsenic trioxide stored underground at level 1500'

This underground storage site

is classified as a class 1 dangerous waste material site

by the GERLED group, a government entity with the mandate to catalogue and

monitor all known dangerous waste material sites in the

Province of Québec.

Although the current resources are located away from the storage facility, pumping water (which

would be necessary to bring the O’Brien Project to production) could potentially disturb the groundwater

and therefore affect the current situation, which is believed to be stable.

Historical precautions may have failed to contain the arsenic trioxide within the containment area over the

last 30 years.

In 1985, the Québec Ministry of Environment authorized the installation of new waterproof and

reinforced concrete plugs (2.3 m wide) at the entrance of each drift containing the barrels, and the

subsequent flooding of the mine;

Drilling from either surface or underground locations could breach the confinement facility.

A buffer zone around the drifts where the barrels are stored should be modelled in

3D, and this buffer zone should be excluded from any future drilling

program.

A hydrogeological study could be initiated to establish whether this area poses a risk and to characterize said

risk. Groundwater should be characterized in order to understand the impact that bringing the current resource

to production would have on the area.

Social acceptability Possibility that portions or the entirety of the O’Brien Project could not be explored or exploited.

Develop a pro-active and transparent strategy to identify all stakeholders and

develop a communication plan. Organize information sessions, publish information on the mining project, and meet with host

communities.

Metallurgical recoveries are based on limited testwork

Recovery might differ from what is currently being assumed.

Further variability testing of the deposit to confirm metallurgical conditions and

efficiencies.

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RISK Potential Impact Possible Risk Mitigation

The custom milling scenario is based on the fact that the

Westwood mill has expressed interest. The plant has

availability in the near future for custom feed. This scenario

could change.

Operating cost used in the PEA could be higher or lower depending on custom milling option available

at the time of operating the projet.

Free gold recovery The content of free gold recoverable by gravity has a

significant impact on the overall gold recovery. Historical data show that the free gold content varies

from one zone to another

Further metallurgical testwork must be conducted to confirm the gold recoveries for a gravity/CIP flowsheet. Only 2 tests

were done in the recent laboratory program. Most of the historical tests

gave lower recoveries for various cyanidation scenarios.

Limited testwork to determine whether waste rock would be

potentially acid generating (PAG)

Additional capital may be required to prepare a storage site for PAG waste.

Further testing to confirm whether the waste is PAG or non-acid generating

(NAG).

Surface and/or underground geotechnical evaluations not

available

The minimum mining width used for the resource estimate might need to be adjusted if assumptions

differ from reality.

The waste pile design is based on common geotechnical data, therefore footprint & pad

construction requirements might be reduced or enlarged, according to the surface geotechnical

evaluation results.

Geotechnical assessments at a larger scale to confirm rock quality

(underground and at surface) to validate assumptions.

Opportunities of the O’Brien Project (Table 25.2)

OPPORTUNITIES Explanation Potential benefit

Aditional geochemical tests on waste rock

Kinetic leaching tests could be done to confirm the ML potential of waste rock

If waste rock is not leachable, an impervious liner nor an impervious cover will be required.

Conduct specific gravity tests from core samples

Potential to increase the 2.67 g/cm3 specific gravity value currently used for

the resource estimate.

An increase in specific gravity increases the tonnage and therefore the ounces of gold.

Compilation of the old O’Brien mine workings

Potential to locate historical underground stopes, channel samples and drill holes with enough precision to

allow this area to be added to the geological model.

An entirely new area could be added that is not considered in the current resource estimate

presented in this report.

Compilation and validation of all remaining historical drill holes

Potential to upgrade the geological model and identify additional resources.

Adding resources increases the economic value of the mining project.

Compilation and validation of all historical underground channel

samples Potential to upgrade some indicated resources to the measured category.

Adding measured resources increases the economic value of the mining project.

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OPPORTUNITIES Explanation Potential benefit

Regarding specifically at Westwood mill opportunities:

A regrind mill could be

refurbished to reduce the grind size before cyanidation.

If the Westwood mill can provide a retention time of 72 hours, higher

recoveries could be achieved.

A bulk sample test should be performed in the Westwood mill.

Surface definition diamond drilling

Potential to upgrade some inferred resources to the indicated category.

Adding indicated resources increases the economic value of the mining project.

Surface exploration diamond drilling on Target 1

Extension of the mineralization within the drilling gap between the historical Kewagama mine

and the 36E area

Potential to identify additional inferred resources.

Adding inferred resources increases the economic value of the mining project.


Extention at depth of the ore

shoot originating in the Kewagama area




Subparallel mineralized zones

north and south of the currently identified zones



Identification of remaining mineralization in the old O’Brine mine area through compilation

and drilling



Recommendations Based on the PEA results, InnovExplo recommends a two-phase work program with the objective, in Phase 1, of increasing the continuity and tonnage of the resources to potentially improve the economics of the project and update the mineral resource estimate and the PEA. Contingent upon the success of Phase 1, InnovExplo recommends initiating a surface exploration and/or conversion drilling program and updating the resources accordingly. Supported by the new resource estimate, InnovExplo also recommends an underground development program. Phase 1 The property-scale compilation should be updated. As part of this compilation, the Company should complete a 3D compilation of the remaining historical openings of the old O’Brien mine, which would have a positive impact on locating all remaining historical underground drill holes and channel samples. The remaining historical data (drill holes, channel samples, etc.) should also be compiled, and the results used to upgrade the current model and resource estimate.

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Exploration drilling should target the currently identified areas of interest described in this report, but also target the discovery of additional zones over the entire project. If additional work proves to have a positive impact on the project, the current resource estimate should be updated to include compiled and validated historical drill holes, future drill holes, underground channel samples and updated 3D models of voids and mineralized zones. Based on the results of the updated resource estimate, the PEA should be updated. Regarding environmental matters, WSP recommends that additional site investigations, data collection, surveys and analyses be initiated as the project progresses to subsequent levels of design, to confirm or revise the current assumptions used for this study. Here is a non-exhaustive list of studies that are recommended:

Geochemical characterization of the waste rock, the ore and the tailings; Characterization of the mine water (groundwater); A baseline study of the receiving environment will be required for the

permitting application process; On-site evaluation of the current water management infrastructure (ponds,

ditches, liners, etc.); Geotechnical and hydrogeological studies for waste rock, ore, and

overburden pads; In an effort to potentially improve mill recovery, WSP recommends:

To conduct a metallurgical study to confirm and improve gold recoveries with a gravity/CIP flowsheet for 36E and Kewagama mineralized material: o Sample the entire mineralized area to evaluate the free gold content per

area/level; o Measure ball mill and abrasion work indexes to better estimate power

and grinding media consumption; o Conduct metallurgical tests in line with the Westwood mill flowsheet

(gravity concentration followed by cyanidation of gravity tails) to optimize reagent consumption;

o Conduct metallurgical tests with a longer retention time; o Conduct further diagnostic testing (via QEMSCAN or other) to determine

the nature of the unleached gold; o Conduct a trade-off study to evaluate whether refurbishing the regrind

mill to obtain a finer grind and thus improve recoveries would be economically advantageous;

o Conduct corresponding metallurgical tests to determine the expected recoveries.

Phase 2 Contingent upon the success of Phase 1, InnovExplo recommends a Phase 2 that includes conversion drilling, which should be devoted to upgrading part of the inferred resources to the indicated category.

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It is recommended to update the mineral resource estimate to include all drilling results. Provision for an underground development program, namely including a bulk sampling campaign aimed at confirming the metallurgy and the continuity of mineralized zones, is considered in the recommended budget. It is recommended to obtain more detailed information about the Westwood process to better evaluate the gold recovery. Additional metallurgical testing should be initiated to improve knowledge through targeted laboratory tests on the cyanidation and gravity circuit conditions and to analyze the mineralogy of gold in discharges. There is a significant amount of data on flotation recovery. However, results for the two cyanidation tests conducted by URSTM are higher than reported historical data. These values should be confirmed to increase the level of confidence in the recovery rate. In addition, the two zones (36E and Kewagama) should be tested individually. The presence of free gold is crucial to recovery. Several historical tests indicate that recovery varies according to the mineralized zone. InnovExplo and WSP have prepared a cost estimate for the recommended two-phase work program to serve as a guideline for the project. The budget for the proposed program is presented in the following table. Expenditures for Phase 1 are estimated at C$3,772,000 (including 15% for contingencies). Expenditures for Phase 2 are estimated at C$19,280,000 (including 15% for contingencies). The grand total is C$23,050,000 (including 15% for contingencies). Phase 2 is contingent upon the success of Phase 1. Estimated costs for the recommended work program (Table 26.1)

Phase 1 - Work Program

Budget

Description

Cost

1a Property-scale compilation including 3D compilation of all remaining historical openings and historical data $100,000

1b Surface exploration drilling (all inclusive) 25,000 m $2,500,000

1c Stakeholder mapping, communication plan $50,000

1d Environmental studies $300,000

1e 3D model and resource estimate update $80,000

1f PEA update $250,000 Contingencies (~ 15%) $490,000 Phase 1 subtotal $3,770,000

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Budget

Description

Cost 2a Surface exploration and/or conversion drilling (all inclusive) 25,000 m $2,500,000

2b 3D model and resource estimate update $80,000

2c Provision for an underground development program $13,500,000

2d Provision for environmental and hydrogeological characterization studies

$600,000

2e Metallurgical testing $100,000 Contingencies (~ 15%) $2,5000,000 Phase 2 subtotal $19,280,000

TOTAL (Phase 1 and Phase 2) C$ 23,050,000

InnovExplo is of the opinion that the recommended two-phase work program and proposed expenditures are appropriate and well thought out, and that the character of the O’Brien Project is of sufficient merit to justify the recommended program. InnovExplo believes that the proposed budget reasonably reflects the type and amount of the contemplated activities.

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2. INTRODUCTION

At the request of Radisson Mining Resources Inc. (“Radisson” or the “issuer”), InnovExplo Inc. (“InnovExplo”) was retained to prepare a Preliminary Economic Assessment (the “PEA”) and Technical Report (the “report”) for the O’Brien Project (the ‘’Project’’), in accordance with National Instrument 43-101 Respecting Standards of Disclosure for Mineral Projects (“NI 43-101”) and its related form 43-101F1. The PEA was prepared with contributions from WSP Canada Inc. (“WSP”) and Lamont inc. (“Lamont”). InnovExplo is an independent mining and exploration consulting firm based in Val-d’Or, Québec. WSP is a professional services firm that operates in different market sectors: property and buildings, transport and infrastructure, environment, industry, mining, oil and gas, and power and energy. Lamont is a professional services firm covering all aspects concerning the design of surface infrastructure (water management, tailings management, access, etc.) and the environment (site characterization, soil and water samplings). Lamont also covers the process associated with the project authorizations from both provincial and federal governments. This report is addressed to Radisson Mining Resources. Radisson is a junior mining company publicly traded under the symbol RDS on the TSX Venture Exchange (TSXV) in Toronto. Radisson is involved in the acquisition, exploration and development of mining properties. Its properties are located in the Province of Québec in the Abitibi-Témiscamingue and Saguenay─Lac-St-Jean regions. This Technical Report supports the disclosure of the PEA results covering the Kewagama and 36E areas of the O’Brien Project, near the town of Cadillac in the Province of Québec.

Examine the potential economic viability of mining the O’Brien deposit; Propose a strategy and preliminary timetable to further develop the project.

The PEA evaluates and/or provides the following items:

The best project design determined from multiple options; The most appropriate mining method determined according to the geometry

and grade of the O’Brien deposit; The basic design for most of the facilities, and the infrastructure needed to

access, develop and mine the mineralized zones; The estimated capital and operating costs; A preliminary cash flow model and an analysis of the financial aspects; Recommendations for additional work to be done in order to advance the

project; A technical report compliant with Form 43-101F1.

Principal sources of information

The PEA prepared by InnovExplo and its collaborators is based on published material as well as data, professional opinions and unpublished material submitted by Radisson or requested by InnovExplo or other participating consultants to complete

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the study. Cost estimation data were also obtained from service providers, suppliers, distributors and manufacturers. Authors also consulted other information sources, such as GESTIM, the Québec government's online claim management system for the status of mining titles, and the SIGEOM online warehouse for assessment work, both available via the website of the Ministry of Energy and Natural Resources (MERN: Ministère de l'Énergie et des Ressources Naturelles), in addition to technical reports, annual information forms, annual reports, management’s discussion and analysis reports, and press releases published by Radisson on SEDAR. InnovExplo and the other participating consultants conducted a review and appraisal of the information used to prepare this PEA, including the conclusions and recommendations, and they are of the opinion that such information is valid and appropriate considering the nature and level of the study (PEA) and the purpose for which the report is prepared. The authors have fully researched and documented the conclusions and recommendations made in the report. Other sources of information used in this report are listed in the references or elsewhere in the text of the report. The consultants do not have nor have they previously had any material interest in Radisson or related entities. The relationship with Radisson is solely a professional association between the client and the independent consultants. This report was prepared in return for fees based upon agreed commercial rates, and the payment of these fees is in no way contingent on the results of this report.

Qualified persons and inspection of the Project The qualified persons (QPs) responsible for the preparation of this Technical Report are:

Sylvie Poirier, Eng. (OIQ #112196, PEO #100156918) of InnovExplo; Pierre-Luc Richard, M.Sc., P.Geo. (OGQ #1119, APGO #1714) of

InnovExplo; Bruno Turcotte, P.Geo. (OGQ #453) of InnovExplo; Laurent Roy, Eng. (OIQ #109779) of InnovExplo Annie Lavoie, Eng. (OIQ #124421) of WSP ; Eric Poirier, Eng. (OIQ #120063) of WSP; Marie-Claude Dion St-Pierre, Eng. M.A.Sc. (OIQ #14097) of WSP. Ann Lamontagne, Eng., Ph.D. (OIQ #104345) of Lamont.

In addition to the principal authors and QPs, the other people involved in the preparation of this report were:

Marie-Claire Dagenais, Eng. (InnovExplo); Éric Caron, Sr. Tech. (InnovExplo).

The list below presents the sections for which each qualified person (as set out in NI 43-101) was mainly responsible:

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Sylvie Poirier, Eng., Director of Engineering, Engineer for InnovExplo Inc., supervised the assembly of the report. She is co-author of and also shares responsibility for sections 1, 2, 3, 21, 22, 24, and 25 to 27. Pierre-Luc Richard, M.Sc., P.Geo., Deputy Director (Resource Estimates), Senior Geologist for InnovExplo Inc., is responsible for the mineral resource estimate and responsible for and author of sections 12 and 14 of the report. He is co-author of and also shares responsibility for sections 1, 7, 25 to 27. Bruno Turcotte, P.Geo., Senior Geologist for InnovExplo Inc., is author of and responsible for sections 4 to 6, 8 to 11, and 23 of the report. He is co-author of and also shares responsibility for sections 1, 7, and 25 to 27. Laurent Roy, Eng., Engineer for InnovExplo Inc., is responsible for and author of Section 16. He is co-author of and also shares responsibility for sections 1, 21, 22, and 25 to 27. Annie Lavoie, Eng., Metallurgical Engineer for WSP, is responsible for and author of sections 13 and 17. She is co-author of and also shares responsibility for sections 1, 21, and 25 to 27. Éric Poirier, Eng., consulting engineer, is responsible for and author of sections 18.3 to 18.8 and shares responsibility for sections 1, 21, 25 to 27. Marie-Claude Dion, Eng. M.A.Sc., Project Manager, is author of and responsible for section 18.1 and 18.2. She is co-author of and also shares responsibility for sections 1, 21, and 25 to 27. Ann Lamontagne, Eng., Ph.D., consultant, is author of and responsible for Section 20. She is co-author of and also shares responsibility for sections 1 and 25 to 27. The following QPs visited the O’Brien property for the purposes of the PEA:

For the purpose of this report I visited O’Brien and Kewagama properties on September 9, 2014, accompanied by Yolande Bisson of O’Brien Project and Éric Caron of InnovExplo.

Eric Poirier of WSP visited O’Brien and Kewagama properties on July 22, 2014 and October 5, 2015.

Note regarding the 2015 Preliminary Economic Assessment

A Preliminary Economic Assessment (PEA) means a study, other than a Preliminary Feasibility Study or a Feasibility Study, which includes an economic analysis of the potential viability of mineral resources. A PEA is defined only in NI 43-101, not in the CIM definition standards. A PEA is preliminary in nature; it includes 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 a preliminary economic assessment will be realized. A PEA uses the concept of “mineral resources within a conceptual mining plan” or “mineral resources within a PEA design plan”. If an issuer can qualify mineral reserves or make a production decision based on a PEA, it may be misleading to call it a PEA. A mineral reserve is the economically mineable part of a measured or

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indicated mineral resource demonstrated by at least a Preliminary Feasibility Study. The results of a Feasibility Study may reasonably serve as the basis for a final decision by a proponent or financial institution to proceed with, or finance, the development of a project.

Units and Currencies All currency amounts are stated in Canadian dollars ($, C$) or US dollars (US$). Quantities are stated in metric units, as per standard Canadian and international practice, including metric tonnes (tonnes, t) and kilograms (kg) for weight, kilometres (km) or metres (m) for distance, hectares (ha) for area, and grams (g) or grams per metric tonne (g/t) for gold grades. Wherever applicable, imperial units have been converted to the International System of Units (SI units) for consistency. A list of abbreviations used in this report is provided in Appendix I.

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3. RELIANCE ON OTHER EXPERTS

The authors, qualified and independent persons as defined by NI 43-101, were contracted by the issuer to study technical documentation relevant to the report, to prepare a PEA on the O’Brien Project and to recommend a work program if warranted. InnovExplo has reviewed the mining titles and their status, as well as any agreements and technical data supplied by the issuer (or its agents) and any available public sources of relevant technical information. Some of the geological and/or technical reports for projects in the vicinity of the O’Brien Project were prepared before the implementation of NI 43-101 in 2001. The authors of such reports appear to have been qualified and the information prepared according to standards that were acceptable to the exploration community at the time. In some cases, however, the data are incomplete and do not fully meet the current requirements of NI 43-101. InnovExplo has no reason to believe that any of the information used to prepare this report is invalid or contains misrepresentations. InnovExplo relied on the following reports and opinions for information that was not within the authors’ fields of expertise:

The issuer supplied information about mining titles, option agreements, royalty agreements, environmental liabilities, permits and details of negotiations with First Nations. InnovExplo is not qualified to express any legal opinion with respect to property titles, current ownership or possible litigation.

Lucie Chouinard (M.Fisc., CPA, CA) of Samson Bélair/Deloitte & Touche completed the after-tax cash flow estimation.

Michèle Mainville, M.Sc.A., of Vee Geoservices provided the linguistic editing for a draft version of the present report.

InnovExplo believes the information used to prepare this report and to formulate its conclusions and recommendations is valid and appropriate considering the status of the project and the purpose for which the report is prepared. The authors, by virtue of their technical review of the project’s exploration potential, affirm that the work program and recommendations presented herein are in accordance with NI 43-101 and CIM technical standards. The authors have sourced the information for this report from the collection of reports listed in Section 27 (References).

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4. PROPERTY DESCRIPTIONS AND LOCATIONS

Location The O’Brien Project is located in the Province of Québec, Canada, just north of the municipality of Cadillac, within the new limits of the city of Rouyn-Noranda (Fig. 4.1).

Figure 4.1 – Location of the O’Brien Project in the Province of Québec

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Cadillac lies approximately 45 km east of downtown Rouyn-Noranda and 45 km west of downtown Val-d’Or. A small part of the urban perimeter of the town of Cadillac touches the southern limit of the Project (Fig. 4.2). The O’Brien Project is located in NTS map sheet 32 D/01 in Cadillac Township. The approximate centre of the Project is at Latitude 48º 14’ 07’’ N and Longitude 78º 22’ 54’’ W, and the approximate UTM coordinates are 694330E and 5345765N, NAD 83, Zone 17.

Mining Rights in the Province of Québec A brief overview of the most common mining rights in the Province of Québec for mineral substances in the domain of the State is provided in Appendix II.

Current Property Description On July 30, 2015, all mining titles constituting the O’Brien Project were converted into “map-designated claims”. Consequently, the O’Brien Project now consists of one contiguous block comprising 21 mining claims staked by electronic map designation (map-designated claims) covering an aggregate area of 637.09 hectares (Fig. 4.2). The map-designated claims are subject to terms under a number of agreements (see Section 4.4). In GESTIM, all titles are in good standing and registered 100% to Radisson Mining Resources Inc. A detailed list of current mining titles, ownership and expiration dates is provided in Appendix III.

Historical Property Description Before July 30, 2015, the O’Brien Project (Fig. 4.3) was represented by the amalgamation of the O’Brien and Kewagama properties, and consisted of 36 mining claims covering an aggregate area of 636.6 hectares. The staked mining claims and map-designated claims were held 100% by Radisson. The staked mining claims and map-designated claims were and continue to be subject to terms under a number of agreements. A detailed list of historical mining titles, ownership, royalties is provided in Appendix IV. The O’Brien property consisted of a contiguous block comprising 33 mining claims covering an aggregate area of 524.53 hectares. Radisson owned a 100% interest in the O’Brien property. The O’Brien property included a mining lease that expired in 2008. The mining lease was not renewed and was converted back into claims. The Kewagama property consisted of a contiguous block comprising three (3) mining claims covering an aggregate area of 112.07 hectares. Radisson owned a 100% interest in the Kewagama property, with a 2% NSR royalty payable to KWG Resources Inc. in the event of commercial production. A $1,000,000 payment must be made to Breakwater Resources Ltd (now Nyrstar) upon commencement of commercial production on either one of the O’Brien or Kewagama properties, against which shall be deducted any costs required to restore the O’Brien tailing ponds.

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Figure 4.2 – Location map showing mining titles constituting the O’Brien Project

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Figure 4.3 – Location map showing historical mining titles constituting the O’Brien Project

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Urban Perimeter As far as exploration and mining activities are concerned, part of the O’Brien Project is affected by regulations regarding the presence of an “urban perimeter” (gray area in Fig. 4.2) or an “area dedicated to vacationing” (dark green area in Fig. 4.2). The restriction, as documented in GESTIM, is “Exploration Prohibited” (see Bill 70, 2013, chapter 32, section 124). According to Bill 70, any mineral substance forming part of the domain of the State and found in an urban perimeter shown on maps kept at the registrar’s office, except mineral substances found in a territory subject to a mining right obtained before December 10, 2013, is withdrawn from prospecting, mining exploration and mining operations as of that date, until the territories provided for in section 304.1.1 of the Mining Act are determined (as of December 10, 2013, the Act to amend the Mining Act: Bill 70). According to section 304.1.1 of the Mining Act, any mineral substance forming part of the domain of the State and found in a parcel of land on which a claim may be obtained and that is included in a mining-incompatible territory delimited in a land use and development plan in accordance with the Act respecting Land Use Planning and Development (chapter A-19.1) is withdrawn from prospecting, mining exploration and mining operations from the time the territory is shown on the maps kept at the office of the registrar. A mining-incompatible territory is a territory in which the viability of activities would be compromised by the impacts of mining. The O’Brien property only includes mining rights obtained before December 10, 2013 and thus exploration is permitted on the mining rights overlapping the urban perimeter and the area dedicated to vacationing until mining-incompatible territories are determined by the regional county municipality (RCM, or MRC in French). In the event that a claim overlaps a mining-incompatible territory, exploration will still be permitted on the overlapping claim, but renewal of such claim will only be permitted if work is performed on the claim during any term occurring after the determination of the mining-incompatible territory (section 61 of the Mining Act).

Territory Akin to an Area for Vacationing According to section 304.1.1 of the Mining Act, mining-incompatible territories will be delimited by RCMs. These mining-incompatible territories will be withdrawn from mining activities. This exercise will be initiated after section 304.1.1 comes into force, once the government has adopted government policies on land use and development, ensuring guidance for RCMs. Meanwhile, territories akin to areas for vacationing will be shown on mining title maps for information purposes only.

Permits Permits are required for any exploration program that involves tree-cutting to create road access for a drill rig, or to carry out drilling and stripping work. Permitting timelines are short, typically on the order of 3 to 4 weeks. The permits are delivered by the MERN.

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Environmental Liabilities Presently, the MERN has exempted Radisson of all liabilities associated with the historical tailings located on site; however, if Radisson should decide to use the same area for tailings in the future, Radisson would acquire all liabilities for the past and present tailings. The presence of a significant amount of arsenic trioxide stored underground at the old O’Brien mine is thus relevant, and is described in Section 20.

Comments on Item 4 InnovExplo is not aware of any other significant factors and risks that may affect access, ownership, or the right or ability to perform the proposed work program on the property.

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

Accessibility The O’Brien Project (O’Brien and Kewagama properties) is located in the northwest part of the Abitibi administrative region, in the western part of Cadillac Township (Fig. 5.1). Highway 117 runs just south of the project’s border. Well-maintained secondary gravel roads provide easy access to the old O’Brien and Kewagama mine sites.

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Figure 5.1 – Topography and accessibility of the O’Brien Project

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Climate The region is under the influence of a continental climate marked by cold, dry winters and hot, humid summers. Weather statistics for the period 1981-2010 show the average temperature for July is 16.7°C, whereas January temperatures hover around -18°C. The record low for this period was -49.5°C in 1994, and the high was 35.5°C in 2005. There are, on average, 75 days without frost. Historical records of annual precipitation indicate a mean rainfall of 985 mm. Snow accumulates from October to May, with a peak from December to March. The nearest permanent weather monitoring station is the Mont Brun station, approximately 28 km northwest of the O’Brien Project. Further climate data can be found at http://climat.meteo.gc.ca/climate_normals. Climate conditions do not seriously hinder exploration or mining activities, and only minor adjustments are needed for seasonal work, such as summer surface mapping and winter drill programs over boggy areas.

Local Resources and Infrastructure About 45 km to the west, Rouyn-Noranda is a town with a population of approximately 39,000 inhabitants and is considered as the regional centre for the western Abitibi region, while Val-d’Or, 45 km to the east, is a renowned gold mining town of 33,250 inhabitants. The area is traditionally a mining area with several operating mines and active exploration companies. Full infrastructure and an experienced mining workforce are also available in a number of nearby well-established mining towns, such as Val-d’Or, Malartic, and Rouyn-Noranda. Both Rouyn-Noranda and Val-d’Or have commercial airports with regularly scheduled direct flights to Montreal. The historical mines on the O’Brien Project saw production from 1925 to 1956. Mining activities on the O’Brien and Kewagama properties resumed from the early 1970s to 1981. Most of the mine surface infrastructure was dismantled in 2012. The mill building and garage still remain. A large power line straddles the south part of the project, and a railway connected to the national network passes through Cadillac, just 2 km south of the project. Radisson has an exploration office and a large, well-equipped core logging and storage facility at the O’Brien mine site. Surface facilities also include large areas for stockpiling ore and waste materials. A tailings facility of 4 hectares and a polishing basin are located directly north of the old mill. A security guard patrols the mine site several times daily, and Radisson has implemented additional measures to maintain security. The closest mill, Agnico-Eagle’s LaRonde Mill, is located about 7 km by road to the west. Other active mills in the area include Doyon-Westwood (Cadillac), Canadian Malartic (Malartic) and Camflo (Dubuisson).

Physiography The topography of the project area is relatively flat to gently rolling, with local relief up to 20 m. The approximate elevation of the Project varies from 305 to 350 masl. There are no distinct prominent topographic features that stand out. Low-lying grounds are characterized by swamps and ponds, and overall drainage is very poor throughout the area. The Blake River flows northeast, running from the southwest corner through the

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O’Brien property to reach Lac Preissac, 3.2 km northeast of the property. The O’Brien Project lies within the boreal forest domain. Predominant tree species include black spruce, balsam fir and tamarack. Local stands of white birch, jack pine and poplar are established on better-drained areas.

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6. HISTORY

O’Brien Property O’Brien Gold Mines Ltd

The following summary of the work conducted by O’Brien Gold Mines Ltd and its predecessors on the O’Brien property is mostly modified and summarized from Bell and MacLean (1929), Cooke et al., (1931), Bell (1937), Gunning (1937), Brown (1948), Dresser and Denis (1949), Paquet (1989) and Bisson (1994). 1924: Claims were staked in the summer of 1924 by two prospectors, Austin Dumont and W. Herweston from M.J. O'Brien Company Ltd. That same year, the No. 1 Vein, the most productive in terms of tonnage, was discovered by Austin Dumont while prospecting. 1925: A two-compartment exploration shaft (No. 1 Shaft) was sunk on the No. 1 vein to a depth of 110 ft, and drifting and crosscutting commenced. The shaft followed the dip of the No. 1 vein (87°N). The rocks near the shaft are Timiskaming sediments, porphyry dykes or sills, and greenstone bands, all striking about east. It was determined that the No. 1 vein occurred in a band of conglomerate 50 to 80 ft wide (15 to 24 m). The conglomerate consisted of pebbles of greenstone, with a few granitic rocks, up to 5 in (13 cm) long, embedded in a medium-grained arkose groundmass. The quartz of the vein was described as dark and glossy, though some sections of white quartz were observed within the dark quartz. Coarse, free gold was found scattered through the quartz at intervals over its entire length. Arsenopyrite was the most common mineral within the quartz, though some pyrrhotite was seen and a little chalcopyrite was also reported. 1926-1929: During the following winter and in the early summer of 1926, a diamond drilling campaign was carried out, comprising twelve (12) holes for a total of 6,000 ft (GM 07451-A). A total of five (5) principal veins (No. 1 to No. 5) were disclosed by surface and underground work, and by diamond drilling. All underground work was confined to the 100-ft horizon of the mine. A crosscut was driven 340 ft to the north of the shaft to intersect the No. 4 and No. 5 veins, where drifts had been cut to the east. The No. 1 vein was opened up by a drift for approximately 900 ft, and 60 ft north of this drift, a second drift was carried on No. 4 vein for a distance of 800 ft. At 310 ft north of the shaft in the main cross-out, a drift was cut on the No. 5 vein, for a distance of 280 ft. At 320 ft east of the shaft, from the cross-cut driven 40 ft north of the drift on the No. 4 vein, a drift was opened up for 45 ft on what is now known as the No. 3 vein. These main drifts and cross-cuts, together with other lateral work of a minor nature, comprised a total of approximately 3,000 ft of underground workings at the time. The high grade shoots in the No. 4 vein were opened up at intervals of approximately 185, 370 and 800 ft, respectively, to the east of the shaft. The first of these, which was 50 ft long where it was intersected by the drift, was stoped through to the surface as a raise. The No. 4 vein occurs within the porphyry. It was uncovered intermittently over a length of about 1,200 ft, and followed the strike of the porphyry body. The width of the vein varied between 6 and 24 in (15 to 61 cm). In one section about 60 ft long, it

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carried a large amount of coarse free gold in small fissures within the quartz. The adjoining country rock was sheared, and carried some free gold near the vein. Stoping was only commenced in 1929 on the most easterly shoot, which contained very spectacular occurrences of visible gold. By the end of 1929, no stopes had yet been opened up in the No. 1 vein. Several shipments of hand-sorted, high-grade ore were sent to Cobalt where the gold was extracted in the O'Brien Mill. Many specimens were retained for museum and exhibition purposes. The gold produced from the small shipments of high-grade ore milled during 1928 amounted to several hundred ounces. 1930: Diamond drilling followed, and in 1930, the No. 2 Shaft (which became the main shaft) was sunk 300 ft east of the No. 1 Shaft. Levels were established at 100, 200 and 300 ft. 1932-1933: An amalgamation mill, with a capacity of 90 tons per day, was built in 1932 and began operating. As of February 1933, the mill was in continuous operation, processing about 75 tons per day. 1934: The No. 2 Shaft was extended from 300 to 500 ft deep, and the 400' and 500' levels developed. As of July 1934, the mine had produced 38,730 metric tons of ore, averaging 15.43 g/t Au. 1935: An addition for roasting and cyaniding the gold-bearing arsenical concentrates was completed and operating in 1935. As of October 5, 1935, a total of 16,219 ft of drifting, crosscutting and raising had been done. By November 1935, the No. 2 shaft had been deepened to 1,035 ft, and stations were established at 625, 750, 875, and 1,000 ft approximately. Production from September 9, 1934, to October 5, 1935, was given as 26,662 metric tons, with a total gold content of 7,865.481 ounces, or 9.19 g/t Au. Of this 66.12% was recovered as bullion, and 26.12% was saved in concentrates for re-processing by the new addition to the mill. It was estimated that extraction should be 92%, giving an overall extraction of 90.06%. At the end of the year, the reserves were estimated to be 20,585 metric tons at an average grade of 8.13 g/t Au. 1936: Late in April 1936, spectacular high-grade ore was encountered in the new lower levels on the No. 4 vein. On May 11, the 6th, 7th, 8th and 9th levels of the No. 4 vein were developed east from the shaft. On each level, free gold was encountered in the vein about 180 ft east of the shaft. This was reportedly about 200 ft west (on average) from the location where the richer shoots were encountered between the surface and the 5th level. At the end of the year, reserves were estimated to be 108,058 metric tons at an average grade of 25.20 g/t Au. They were mainly related to the new discovery of the No. 4 Vein, a particularly rich vein.

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1937: The milling capacity was increased to 150 tons per day. During the year, the No. 3 Shaft was started on the western section of the property where excellent diamond drilling results had been obtained by the surface drilling campaign of 1935. Stations were established at 125-ft intervals, and a zone of multiple high-grade ore carriers was identified on the upper levels. 1939: The No. 3 Shaft reached a maximum depth of 1,500 ft. Remarkably greater quantities of gold were recovered from 1937 to 1939. They came partly from processing arsenopyrite concentrates that had been stockpiled in the past, but mostly from mining an extremely rich mineralized chimney in the No. 4 Vein. 1940: The first shipment of crude arsenic was made in 1940 to Deloro Smelting & Refining Company in Deloro, Ontario, with production sales continuing until 1950. Crude arsenic, grading 83.0% arsenic trioxide and 8.5 to 12.0 g/t Au, was refined and the sludges returned to the O’Brien mine for gold recovery. 1941: Since 1930, the hoisting from the No. 2 Shaft had been in cars. In 1941, the No. 2 Shaft was converted to skips with an ore transfer system at the 2000' level. That same year, production stoping was by rill shrinkage with changeover to inclined cut and fill in the deeper levels. Stoping width varied from 4 to 20 ft, with the average on the narrow end of that range. The sinking of the No. 4 Shaft began in October 1941. 1942: Production peaked in 1942 at 63,086 metric tons milled averaging 12.79 g/t Au, and reserves were at their highest at 218,648 metric tons averaging 12.14 g/t Au. 1943: In 1943, 43,269 metric tons averaging 11.66 g/t Au were mined from the No. 3 Shaft, representing about 5.6% of the total mine production. A total of 3,400 ft of drifting were completed from the No. 3 Shaft. Apart from a small amount work in 1950, no work was done from the No. 3 Shaft after 1943. Management reports repeatedly cite the labour shortage as the reason. 1949: In January, the sinking of No. 4 Shaft resumed and wascompleted in July 1949. This internal shaft (winze) was sunk from 2,000 to 3,500 ft between 1941 and 1949. Reserves slowly declined between 1942 and1949 and fell off rapidly thereafter. 1952: By 1952, rising costs eroded profits to a break-even point, and ore reserves declined to a 2-year supply. Leads to new high-grade ore were considered to be exhausted on the development levels, and the most favourable prospecting ground was considered to be at the depth. The last commercial crude arsenic shipment was made between 1951 and 1952 to Belgium. 1954: A drilling program was carried out to explore the area between depths of 3,400 and 4,000 ft. Seven holes totalling 4,000 ft were drilled below the 3400' level, and results reported in 1954 indicated continuity of the No. 1 vein, although gold values could not support shaft sinking or a continuing operation. 1956-1957: On July 1956, the operation of the O’Brien mine was closed down. The surface plant, the mill and all equipment where gold might have accumulated were cleaned until 1957. The mine closed because of rising operating costs, lower grades from stopes, and the fixed price of gold at US$35.00.

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In 1956, a stockpile containing an estimated 1,150 metric tons of crude arsenic (arsenic trioxide) was stored in 8,938 barrels west of the No. 3 Shaft on the 1500' level in the 15-G-West and 15-F-West drifts. Drift entries were sealed with concrete plugs about 1.2 m wide. The mine was flooded thereafter. Table 6.1 details the mine workings completed at the O’Brien mine between 1926 and 1957. Table 6.1 – Total mine workings at the O’Brien mine from 1926 to 1957

Between 1926 and 1956, a total 587,120.8 ounces of gold were produced from 1,197,147 metric tons milled with an average grade of 15.25 g/t Au (Table 6.2). Recoveries averaged 96.0% with losses distributed as follows: 2.6% flotation, 0.4% roasting and 1.0% cyaniding. This would indicate a grade of 0.7 to 1.0 g/t Au for the mill tailings. The O’Brien mine also produced 6,313 metric tons of crude arsenic, of which 5,176 metric tons were sold. The ore averaged 0.6% As, and concentrates contained 10% As.

Mine Workings Meters

Drifting 25,588.0

Crosscutting 5,563.5

Raising and Boxholing 2,511.9

Shaft Sinking 1,556.7

Station Cutting 478.2

Underground Drilling 54,282.4

Surface drilling 6,185.9

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Table 6.2 – Total gold production of the O’Brien mine from 1926 to 1957

Darius Gold Mines Inc. The following summary of the work conducted by Darius Gold Mines Inc. on the O’Brien property is mostly modified and summarized from Schaaf (1972; 1976a to 1976f), Scobie (1972), Brethour (1974; 1975a; 1975b; 1976), van de Wall (1980), Lafleur (1980), Rive (1981; 1982), Paquet (1989), Bisson (1994) and Charlton (1994). 1969: Abandoned since its closure in 1956, the O’Brien mine was acquired by A. N. Ferris and the property renamed the Ferris property. The property was re-evaluated, and surface work (mostly scouring) was carried out. That same year, A. N. Ferris created Darius Gold Mines Inc. (“Darius”). 1972: In 1972, Darius began an exploration and reassessment program at the former O’Brien mine. A brief study on the tailings from the former O’Brien mine was carried out to ascertain the form of the contained gold and the amount that might be recoverable by further treatment. Four samples of mill tailings, weighing approximately 100 pounds, were received at Lakefield Laboratories. The head assays from the

YearMetric Tonnes

Mined (Hoisted)

Metric Tonnes

Milled

Au g/t

Milled grade

Ounces of Gold

Recovered

Metric Tonnes

Development

Au g/t

Development

Metric Tonnes

Stopes

Au g/t

Stopes

1926-1932 1,574 94.50 4,782

1933 13,481 10.97 4,755

1934 24,796 9.57 7,626

1935 26,662 6.07 5,200.9

1936 24,497 18.89 14,875.6

1937 33,897 33.84 36,879.5

1938 50,912 50,902 24.61 40,280.2 23,037 12.00 27,875 32.57

1939 52,516 61,286 19.05 37,538.7 22,606 7.89 29,711 34.59

1940 61,286 61,563 14.40 28,494.2 13,808 10.90 45,746 16.77

1941 62,757 62,730 12.52 25257,4 3,468 7.34 53,534 14.40

1942 63,066 63,086 12.79 25,947.0 9,306 11.38 53,760 13.78

1943 62,882 62,701 13.04 26,286.2 3,346 8.64 59,536 13.92

1944 50,552 50,652 16.00 26,049.0 2,875 10.80 47,677 17.11

1945 44,810 44,918 17.98 25,964.2 6,718 14.47 38,092 19.34

1946 45,748 45,784 15.54 22,868.2 4,129 9.60 41,620 16.80

1947 48,053 48,048 14.95 23,092.4 3,200 9.02 44,853 16.05

1948 49,600 49,699 17.09 27,308.5 6,173 7.89 43,427 19.27

1949 52,890 52,702 15.89 26,920.5 3,771 9.02 49,119 17.18

1950 60,550 60,686 14.49 28,266.9 5,197 8.88 55,353 15.77

1951 59,139 59,139 14.66 27,870.9 3,509 8.13 55,630 15.77

1952 61,393 61,393 13.02 25,705.7 2,631 11.69 58,762 13.71

1953 58,088 58,088 12.84 23,973.6 1,420 8.88 56,668 13.44

1954 62,879 62,879 12.74 25,752.5 1,761 10.22 61,118 13.37

1955 63,616 63,616 11.37 23,251.7 1,328 8.23 62,287 11.97

1956 52,012 52,370 11.94 20,099.6 351 7.61 51,661 11.04

1957 2,074.4

TOTAL 1,062,749 1,197,147 15.25 587,120.8 118,635 10.07 936,427 16.17

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sample graded 0.86 g/t Au and 0.047% As, and contained 0.22% sulphur. One flotation test was attempted, but the results were discouraging. The sulphide concentrate was very low grade with low recovery. The cyanidation test was carried out on the tailings sample. The recovery was improved by grinding the sample, yielding recoveries of 78.2% and 81.5% after 24 and 96 hours, respectively. 1973: Darius pumped out the O’Brien mine to the 9th level (1400') and began a sampling program. The headframe and hoist were installed on the No. 2 Shaft. Chip samples were taken at 20-ft intervals on the 250', 375', 500', 625' and 750' levels. 1974: Darius carried out an underground bulk sampling program composed of many samples. The samples were blasted 6 ft high across the width of the vein, and for as long as it was exposed in the drifts. The samples were chosen according to vein width, and varied in length from 20 to 45 ft (6 to 14 m). Once blasted, the samples were mucked and shipped separately by truck to the Malartic Goldfields Mill, a distance of 42 km, where they were sampled and run through the mill circuits. Early in March, a dump ramp was built on the west side of the headframe, and one mucking machine and four 1-ton cars were purchased. Track was installed from the cage, and cars were dumped one at a time directly into the truck. With this method, a complete sample could be mucked and shipped in one day, consisting on an average of about 2 truckloads. Between February and April, 1974, a total of 171 metric tons was extracted from the 375' level in the F and G veins. 1975: At the end of February 1975, a total of 2,500 metric tons averaging 3.14 g/t Au were extracted during the bulk sampling program at the O’Brien mine. A total of 2,406 linear feet of drift backs on the 375', 500', 625', 750' and 875' levels were sampled. A total of 523 ft of drifting and 422 ft of raising (three raises) were completed. A total of eighteen (18) underground holes (74-1 to 74-11, D-16, D-18, D-19, D-21, D-24 and D-25) were drilled for a total of 2,985 ft. 1976: A total of thirty-two (32) underground holes (D-20, D-22, D-23, D-26 to D-29, D-31 to D-51 and D-53 to D-56) were drilled for a total of 4,275 ft. Following the underground drilling campaign, Robert E. Schaaf carried out a mineral inventory compilation on veins No. 1 S, No. 1 N, F9 and H-4-14. 1977: In October 1977, Goldfield Mining Consolidated acquired a 51% interest in the Darius Gold Mines property for US$4,635,000, with a commitment to spend enough money to make the mine operational and explore adjacent properties. The acquisition led to additional restoration work and bulk sampling. Darius built a mill with a capacity of 200 short tons per day, which could be increased to 500 short tons per day. The mill was completed on June 1, 1978, for about C$3,000,000. 1978: A total of 11,018 metric tons grading 1.07 g/t Au were milled in the new mill. The ore essentially came from drifting. 1979: Darius undertook a surface drilling program comprising 24 holes (GF-79-1 to GF-79-24) for a total of 3,979.8 m in order to test the areas that had never been explored.

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A total of 36,106 metric tons grading 3.04 g/t Au were milled in the new mill. The ore was produced from small stopes. 1980: Darius completed a surface drilling program comprising 33 holes (DS-01 to DS-28, DS-30, DS-33 to DS-35, and DS-37) for a total of 4,995.5 m in order to test the area that had never been explored. A total of 33,706 metric tons grading 3.73 g/t Au were milled in the new mill. The ore was produced from small stopes. 1981: The mine was closed at the end of August, and the mill ceased activity in October. An estimated 47,587 metric tons averaging 2.79 g/t Au were milled in the new mill. Between 1974 and 1981, a total 10,852.4 ounces of gold were produced from 128,373 metric tons milled averaging 2.63 g/t Au (Table 6.3). Recoveries averaged 70.0%. During the year, Darius believed it had a buyer for the crude arsenic stored on the 1500' level since 1956. The concrete wall from the level 1500' was screwed. Later, the potential buyer withdrew. Table 6.3 ─ Total gold production from the O’Brien mine from 1974 to 1981

* Estimated data

Sulpetro Minerals / Novamin Resources / Breakwater Resources The following summary of the work conducted by Sulpetro Minerals, Novamin Resources and Breakwater Resources on the O’Brien property is mostly modified and summarized from Vaillant and Hutchinson (1982), Wright (1986), Quan (1987), Glover (1989), Sauvé and Trudel (1991), Trudel et al. (1992), Lelièvre (1994) and Bisson (1994). 1981: In December, Sulpetro Minerals Ltd (“Sulpetro”) bought the property for C$2,800,000 for the purpose of treating ore from its adjoining Kewagama mine to the east. The property was renamed O’Brien Division. Sulpetro tried unsuccessfully to find other buyers for the crude arsenic stored on the 1500' level.

YearMetric Tonnes

Milled

Au g/t

Milled grade

Ounces of Gold

Recovered

1974-1975 2,500 3.14 252.4

1978 11,266 0.78 282.6

1979 36,114 2.48 2,875.7

1980 33,388 3.15 3,381.2

1981 45,105* 2.79* 4,060.4*

TOTAL 128,373 2.63 10,852.4

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1985: Sulpetro completed magnetometric (49.5 line-km) and VLF electromagnetic (49.5 line-km) surveys over the property, including a limited amount of IP (4.9 line-km) surveys. The mine was closed down that same year, although the facilities were kept. All electrical equipment was removed from the No. 2 Shaft. In April, Québec’s Ministry of Environment authorized the installation of new waterproof and reinforced concrete plugs (2.3 m wide) at the entrance of each drift containing crude arsenic. In August, the Ministry of Environment authorized the flooding of the mine. 1986-1987: In January 1986, Sulpetro was reorganized into Novamin Resources Inc. (“Novamin”). In 1986 and 1987, surface drilling was done in the area of the No. 3 Shaft, extending the No. 2 and No. 4 vein structures towards the New Alger property boundary. Eight (8) drill holes totalling 1,999.8 m were drilled (2130-1 to 2130-8). Later, Novamin added eight (8) new holes totalling 2,185 m (2130-9 to 2130-16). These holes led to the discovery of a new gold prospect in the area of line 36+00E (Zone 36 East). It consisted of a series of gold-bearing quartz echelon veins that were similar in nature and character to the mined structures of the O’Brien mine. Control of Novamin was acquired by Breakwater Resources Ltd (“Breakwater”) later in 1987. 1988: At the beginning of the year, Novamin drilled eight (8) additional holes (2130-17 to 2130-24) on the Zone 36 East for a total of 2,198.5 m. 1989: Breakwater completed the acquisition of Novamin and continued drilling the property. A total of 24 holes (2130-25 to 2130-46, incl. 2130-40E and 2130-40W) were drilled on Zone 36 East for a total of 7832.1 m. Surface drilling on the eastern part of the O’Brien mine property had begun to outline a significant gold occurrence. Breakwater outlined an Inferred mineral inventory on Zone 36 East of 249,746 metric tons averaging 8.23 g/t Au using a cut-off grade of 3.4 g/t Au and totalling 66,071 ounces. This inventory was developed using a 7.6-m (25-ft) and 45.7-m (150-ft) vertical maximum zone of influence from each pierce point. The cut-off was 3.4 g/t Au / 1.2 m, with combined individually cut grades diluted to 1.2 m (4 ft) if necessary, and zero values assigned to wing samples. Individually cut assays were established at 34.3 g/t Au. Neither gold price nor exchange rate was mentioned in the Breakwater report. These “resources” are historical in nature and should not be relied upon. It is unlikely they conform to current NI 43-101 criteria or to CIM Standards and Definitions, and they have not been verified to determine their relevance or reliability. They are included in this section for illustrative purposes only and should not be disclosed out of context.

Radisson Mining Resources The following summary of the work conducted by Radisson Mining Resources on the O’Brien property is mostly modified and summarized from Bisson (1994; 1995; 1996;

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2004), Kroon (1996; 1997), Karpoff and Evans (1998), Barrie (2006), Evans (2007), Vincent (2009), David and Gauthier (2012), de l’Étoile and Salmon (2013), and Radisson’s annual reports (1997 to 2013). 1994: On October 24, a deal was signed whereby Radisson Mining Resources Inc. (“Radisson”) could earn a 50% interest in Breakwater’s O'Brien property. Under the deal, Radisson could earn a 50% interest by spending C$3,000,000 on exploration and issuing Breakwater 500,000 class A Radisson shares by Feb. 28, 1999. In addition, the deal gave Breakwater the option to purchase 200,000 Radisson shares at 40 cents each. Breakwater retained ownership of the surface infrastructure, including the mill, but Radisson had the option to purchase a 50% interest in these facilities once it had spent its C$3,000.000. 1995: Radisson compiled data and proceeded with a new geological interpretation on Zone 36 East. Between December 1994 and February 1995, twelve (12) holes (OB-95-47 to OB-95-56, including OB-95-55A and OB-95-56A) totalling 3,998.4 m were drilled on Zone 36 East in order to increase the mineral inventory of the zone. The Indicated mineral inventory of Zone 36 East was estimated at 489,277 metric tons at 7.20 g/t Au using a cut-off grade of 3.4 g/t Au, for a total of 113,260 ounces. This inventory was developed using a 7.6-m (25-ft) and 45.7-m (150-ft) vertical maximum zone of influence from each pierce point. Individually cut assays were established at 34.3 g/t Au. Specific gravity was fixed at 2.67. A 3.4 g/t Au / 1.2 m (true thickness) cut-off was used. Neither gold price nor exchange rate was mentioned in the Radisson report. These “resources” are historical in nature and should not be relied upon. It is unlikely they conform to current NI 43-101 criteria or to CIM Standards and Definitions, and they have not been verified to determine their relevance or reliability. They are included in this section for illustrative purposes only and should not be disclosed out of context. 1996: Between December 1995 and February 1996, Radisson added thirty-one (31) holes (OB96-57 to OB96-75, incl. OB96-57A, OB96-62A and 10 wedged holes) for a total of 11,962.8 m. The purpose of this campaign was to increase the confidence level of the mineral inventory from the surface to 1,200 ft elevation, and to demonstrate the presence of an extension of the veins at a vertical depth below 2000 ft. The total gold resources were 1,270,000 metric tons at an average grade of 6.9 g/t Au (cut) and 8.6 g/t Au (uncut). Of this total, 735,600 metric tons were in Zone 36 East, averaging 7.2 g/t Au (cut) or 10.6 g/t Au (uncut). Kilborn SNC-Lavalin wrote up an independent study supporting the evaluations of Radisson’s geologists. This inventory was developed using a 7.6-m (25-ft) and 45.7-m (150-ft) vertical maximum zone of influence from each pierce point. Assays were cut at 34.3 g/t Au. Specific gravity was fixed at 2.67. A 3.4 g/t Au / 1.2 m (true thickness) cut-off was used. Neither gold price nor exchange rate was mentioned in the Kilborn SNC-Lavalin report. These “resources” are historical in nature and should not be relied upon. It is unlikely they conform to current NI 43-101 criteria or to CIM Standards and Definitions, and they have not been verified to determine their relevance or reliability. They are included in this section for illustrative purposes only and should not be disclosed out of context.

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During fall 1996, eleven (11) outcrop were stripped at a distance of 400 ft east of the No. 2 Shaft in order to evaluate the gold potential of two auriferous structures (2V and Contact Veins) located in sedimentary rocks of the Pontiac Subprovince, near the contact with the Piché Group. Some anomalous gold values were obtained from quartz veins in the sedimentary rocks. 1997: Two drilling programs were conducted in 1997. The first, at the beginning of the year, totalled 1,283 m in seven (7) holes (OB97-76 to OB97-82) and focused on the quartz veins associated with the contact zone between the Pontiac Group and the Piché Group (former mine unit). Drilling was done in the central part, but despite some economic grades, it di not confirm their mining potential. On September 30, 1997, a new drilling program began in Zone 36 East. In all, 4,555 m was drilled in 23 holes (OB97-83 to OB97-103, incl. OB97-87B and OB97-96B) between sections 32E and 44E, from surface tp a vertical depth of 230 m. 1998: Following a letter of agreement signed on December 9, 1998, between Radisson, 3064077 Canada Inc. and Breakwater Resources Ltd, Radisson purchased 100% of the rights to the O’Brien property as well as all the infrastructure, in addition to acquiring the Kewagama property adjacent to the O’Brien property. In June 1998, an independent study signed by Roscoe Postle Associates Inc. (“RPA”) updated the gold resources in Zone 36 East in the O’Brien mine. As at April 30, 1998, the Indicated resources, down to a depth of 610 m below surface and using a cut-off grade of 5.1 g/t Au, totalled 348,365 metric tons grading 9.9 g/t Au cut to 68.5 g/t Au (14.5 g/t uncut) for a total of 111,000 contained ounces (162,000 ounces uncut). As at the same date and to the same depth, Inferred resources at 5.1 g/t Au cut-off grade totalled 15,422 metric tons grading 18.6 g/t Au cut to 68.5 g/t Au (19.8 g/t uncut) for a total of 9,000 contained ounces (10,000 ounces uncut). The specific gravity was fixed at 2.67. The price of gold was US$300/oz with an exchange rate of 1.444. These “resources” are historical in nature and should not be relied upon. It is unlikely they conform to current NI 43-101 criteria or to CIM Standards and Definitions, and they have not been verified to determine their relevance or reliability. They are included in this section for illustrative purposes only and should not be disclosed out of context. RPA’s mandate also included a preliminary prefeasibility study to evaluate the viability of commercial production for the project. The study concluded that the project would not be profitable at the US$300/oz gold price and exchange rate of 1.444. The resources would have to increase, and a better grade than the cut grade of 6.9 g/t Au would have to be confirmed, as well as a metallurgical recovery of at least 90%. Two metallurgical tests were completed in two Canadian laboratories in 1998 on sulphide concentrates originating from Zone 36 East and Zone F. Two different processes were verified: bioleaching at the BC Research Laboratory in Vancouver, and microwaves at the EMR Technology Laboratory in Fredericton, New Brunswick. The objective was to maximize to 90% the recovery of sulfide-related gold at a competitive processing cost. With direct cyanidation, the recovery barely reached 80%.

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In May 1998, two drill holes (OB98-106 and OB-98-107), totalling 546.8 m, were completed on targets identified outside the known zones north of the Cadillac–Larder Lake Fault Zone (CLLFZ). A network of horizontal gold-bearing quartz veins with free gold was discovered. The best grade was 6.9 g/t Au over 2.33 m. In November, another drilling program (1,402.7 m) was completed to locate other gold-bearing veins north of the CLLFZ. The five drill holes (OB98-108 to OB98-112) intersected interesting settings. 2001: On August 24, 2001, Radisson signed an initial agreement with Rocmec concerning preliminary tests and the use of a new extraction technology applied to the gold-bearing quartz veins on the O’Brien property. The two partners decided to drill a series of pilot holes in an easily accessible exposed surface vein near the Radisson installations. Rocmec drilled an initial series of thermal holes supervised by Radisson personnel. This work allowed 1.54 metric tons of gold-bearing quartz vein material to be extracted. The sample thus extracted was processed on a Deister table in the Radisson concentrator, on site in Cadillac. The gold in the batch totalled 35.245 grams, or a grade of 22.83 g/t Au. Recovery reached 77%. This work confirmed a high rate of recovery by gravimetry and an excellent grade for the smokey quartz veins in the former O’Brien mine. 2003: In the summer of 2003, a surface exploration program was carried out for the purposes of verifying the surface extraction potential of gold-bearing quartz veins in the former O’Brien mine area, approximately 900 ft east of the headframe, and of the Zone 36 East veins. The former O’Brien property was stripped to reveal new smokey quartz veins. The samples taken in the stripped zones did not yield economic grades. In the Zone 36 East area, three holes (OB03-02 to OB03-04) were completed for a total of 210.3 m of drilling. Two composite core samples drilled on the same zone, one from a vein and the other from its wall, were constituted and analyzed at Laboratoire LTM in Val-d’Or. The test was intended to determine the content of the vein and the wall, as well as to verify the gold recovery ratio by gravimetric method. A content of 4.80 g/t Au was obtained for the vein with a 63% recovery by gravity. The wall yielded 2.40 g/t Au gold and an equivalent recovery. On their own, these results could not justify a major surface bulk sample test, and it was decided to discontinue efforts to verify this scenario. In July 2003, Radisson decided to abandon its surface exploration efforts on the O’Brien property after carrying out a cursory stripping and short drilling program to verify the possibility of extracting the gold veins reaching the surface. Based on the results, the company concluded it was not worth continuing surface work at this time. 2004: An initial diamond drilling campaign to verify depth potential was completed in 2004 for the purposes of analyzing “Contact Zone”-type gold mineralization on the O’Brien and Kewagama properties. This program studied the favourable horizon to a depth never before explored. The objective was to significantly increase the potential and value of the Radisson lands by discovering more extensive gold structures at depth, along the CLLFZ, compared with the known vein system near the surface. A hole (OB04-01A) was drilled on the O’Brien property under Zone 36 East, reaching a total length of 1,535 m. It confirmed the continuity of the gold-bearing Zone 36 East to double its previously known depth.

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The hole cut Zone 36 East and intersected mineralized alteration zones at depth, also in the Piché Group volcanics. This setting is very similar to that of the Lapa mine Contact Zone, also located within the Piché Group. 2006: A high-resolution aeromagnetic, horizontal gradiometer and XDS-VLF-EM survey was carried out on the O’Brien and Kewagama properties in June 2006. The survey, which was the first phase of the 2006 exploration program, was conducted by Terraquest Ltd with a flight line spacing of 50 m. Data from this survey was used to define drill targets north of the CLLFZ. Radisson also carried out a lithogeochemical sampling program focusing on the talc/chlorite schists in drill core stored at the O’Brien mine site. The program’s objective was to verify the presence of mineralization similar to the D Zone on the Wood/Pandora Project. A diamond drilling program was then carried out on the property. A total of three holes (OB06-17 to OB06-19), totalling 1,198 m were drilled on the No. 2 Vein, Zone 36 East and the North Zone. 2007: Scott Wilson Roscoe Postle Associates Inc. (“RPA”) estimated the mineral resources of Zone 36 East using the historical surface and underground drilling data available in April 2007. The resources provided below were estimated using a conventional 2D longitudinal block resource estimation methodology, a horizontal thickness for indicated resources ranging from 1.2 to 2.7 m with an average of 1.4 m, a gold price of US$575/oz, a US exchange rate of 0.87, a gold recovery of 90%, a specific gravity of 2.67, and a selected capping level of 68.5 g/t Au. At a 5.8 g/t Au gold cut-off grade, RPA estimated that the Indicated resources of Zone 36 East amount to 251,295 metric tons at an average cut grade of 12.3 g/t Au for a total of 97,000 contained ounces. RPA estimated that the Inferred resources totalled 165,110 metric tons at an average cut grade of 9.9 g/t Au for a total of 54,000 contained ounces. The Zone 36 East mineralization was very sensitive to cutting high gold assays, and the cut Indicated average grade was approximately 36% lower than the uncut Indicated average grade. Cutting high gold assays reduced the contained gold in the global resource by approximately 30% from the uncut figure. The 2007 exploration program included 60.8 km of line cutting, 46.1 km of IP, and 2,053.2 m of diamond drilling in 15 holes (OB07-120 to OB07-134); the drilling program continued until March 2008. The purpose of the drilling program was to test the resource blocks identified in the 2007 NI 43-101 report on Zone 36 East resources (Evans, 2007). In late 2007, negotiations were initiated with Aurizon Mines Ltd (“Aurizon”), which was interested in becoming Radisson’s partner on the O’Brien/Kewagama Project. 2008: From January to March 2008, the drilling program totalled 3,738.7 m in 21 holes (OB-08 to OB08-150). On April 14, 2008 Radisson agreed to grant Aurizon an option to acquire an undivided 50% interest in the O’Brien/Kewagama Project. The transaction was subject to a number of conditions, including completion of satisfactory due diligence. By

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September 2008, Aurizon had been conducting a due diligence investigation on the project for almost six (6) months. Subsequently, Aurizon requested that it be entitled to earn a 75% interest in return for conducting the study, a proposal declined by Radisson. In fall 2008, an exploration drilling program was carried out on the O’Brien property totalling 1,920.6 m in 7 holes (OB08-152, OB08-153, OB-85-153A, OB-08153B, OB08-161, OB08-162, and OB08-162A). Three holes, OB08-153B, OB08-161 and OB08-162 (hole OB08-152 was stopped in the CLLFZ), tested the eastern extension of Zone 36 East and, in particular, the high gold values obtained in hole OB08-149. 2011: A total of six (6) holes (RM-11-03, RM-11-04, RM-11-14 and RM-11-16 to RM-11-18) were drilled on the O’Brien property for a total of 1,989.0 m. The program was designed to carry out resource definition drilling on Zone 36 East to categorize the inferred resource and potentially increase total resources. 2012: An exploration drilling program was carried out on the O’Brien property totalling 2,112.5 m in three (3) holes (OB-12-20 to OB-12-22). The holes also returned gold intersections in Pontiac Group sandstone to the south of the formations containing O’Brien-type mineralization. Visible gold was observed in two of the holes. 2013: RPA estimated the mineral resources of Zone 36 East using the historical surface and underground drilling data available up to December 2012. The resources provided below were estimated using a block model in GEMCOM software, a minimum horizontal width of approximately of 1.8 m, a gold price of US$1,600/oz, a US exchange rate of 1.0, a gold recovery of 90%, a specific gravity of 2.67, the selected capping level was 51.9 g/t Au. At the 3.4 g/t Au gold cut-off grade, RPA estimated that the Indicated resources of Zone 36 East amount to 508,032 metric tons at an average cut grade of 6.5 g/t Au for a total of 106,000 contained ounces. RPA estimates that the Inferred resources amount to 287,582 metric tons at an average cut grade of 7.29 g/t Au for a total of 67,000 contained ounces. According to RPA, there some of the Inferred resource of Zone 36 East could potentially be converted to Indicated through additional drilling. RPA also considered the eastern extension of Zone 36 East, up to the Kewagama property, to be open, and that follow-up exploration on the 2011 and 2012 results was warranted. Table 6.4 shows the statistics from the Radisson drilling campaigns carried out on the O’Brien property between 1995 and 2012.

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Table 6.4 ─ Holes drilled by Radisson between 1995-2013

Kewagama Property Kewagama Gold Mines Ltd

The following summary of the work conducted primarily by Kewagama Gold Mines Ltd on the Kewagama property is mostly modified and summarized from Bell (1937), Gunning (1937), Dresser and Denis (1949), Pouliot (1964), Dugas et al. (1967), Brereton (1973), Thompson (1974), Schaaf (1979), Laronde (1980), Vaillant and Hutchinson (1982). 1928: Activity on the property commenced in 1928 with trenching and diamond drilling by Cartier Malartic Gold Mines. 1931: In 1931, eight (8) of the present claims were acquired by Canadian Gold Operators Ltd. 1932-1933: A considerable amount of development was carried out by Canadian Gold Operators Ltd, including diamond drilling (10 holes aggregating about 5,000 ft), the sinking of a two-compartment shaft to a depth of 125 ft, and approximately 1,500 ft of lateral work (drifts and crosscuts) at the 125' level. The shaft is 4,800 ft east of the O'Brien No. 2 Shaft. The work indicated that geological and structural conditions of the Kewagama property to the east, are essentially similar to those of the adjoining O’Brien property. The exploration revealed the presence of several gold-bearing quartz veins. A total of four veins (Nos. 1, 6, 7 and 8) were developed and investigated. Although the limited amount of drifting that was done on these veins did not establish any ore shoots, it did disclose encouraging gold values. The property was closed down in April 1933. 1934-1935: Underground workings were flooded. 1936: Control of Canadian Gold Operators Ltd was acquired by Ventures Ltd, and the property plus an additional claim adjoining the northeast corner was turned over to a new Ontario company, Kewagama Gold Mines Ltd.

Year Number of

Holes

Total Length

(meter)

1995 10 3,726.2

1996 31 14,530.1

1997 37 6,586.1

1998 7 1,949.5

2003 3 210.3

2004 2 1,656.0

2006 3 1.198.0

2007 15 2,053.2

2008 28 5,659.3

2011 6 1,989.0

2012 3 2,113.5

TOTAL 145 41,610.7

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1937-1939: The shaft was deepened to 524 ft, with three compartments, and new levels were established at 250, 375 and 500 ft. At a point 400 ft east of the shaft, a winze was from the 500' level to the 700' level, and new sublevels were established at 550, 600, and 700 ft. Lateral developments were carried out on four levels from the shaft, and three sublevels from the winze. A total of 12,600 ft of drilling were carried out. Although interesting gold assays were obtained from the material encountered, especially on the lower levels, commercial grade ore was not in sufficient quantity to assure a profitable venture, and all operations were suspended in early 1939 due to the restrictions on gold mining with the outbreak of World War II. 1940: A total of 2,470 metric tons of stockpiled development ore, having an average grade of 9.9 g/t Au, was processed at the neighbouring Thompson Cadillac Mill, from which 790.7 ounces of gold were recovered. 1947: A magnetometer survey was completed over the Piché Group (Cadillac Shear Zone) and the Cadillac Formation north of the shear, to determine whether the gold mineralization of the neighbouring Wood-Central and Pandora properties to the east continued onto the Kewagama property. 1964: Falconbridge Nickel Mines, the successor to Ventures Ltd, initiated a surface drilling program in 1964, partially for assessment work. Four (4) holes totalling 981.7 ft were completed (S-46 to S-49), approximately 50 ft apart, to trace the upward extension of the Winze Zone that had been partially developed from the 500' level from 1937 to 1939. 1973-1974: Surface exploration was renewed by Kewagama Gold Mines Ltd under the direction of Derry, Michener & Booth, Geological Consultants. A program of overburden (basal till) sampling for gold was conducted along the 2,800-ft strike length of the favourable Cadillac Belt of rocks extending east of the 1964 Falconbridge drill holes and north of the Cadillac Shear, to explore the iron formation environment that had been productive on the neighbouring Wood-Central and Pandora properties to the east. Diamond drilling followed, consisting of 13 holes (S74-1 to S74-13) for a total of 3,149 ft. Results were considered encouraging and worthy of underground investigation. 1976: Management control of the company was acquired by A. N. Ferris of Cadillac, Québec. 1977: The mine site was cleared of bush and leveled. 1978: A temporary mining plant–service building, a hoist room, a headframe, a mine dry and a machine shop were constructed. 1979-1980: The hoist was operative in early 1979, and the mine was dewatered and secured in May. Inspection of underground workings took place, followed immediately by sampling and planning. The company removed the pentice to form a third compartment, rehabilitated the shaft, sank approximately 200 ft of shaft, cut a station on the 700' level and drove 800 ft of drift.

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On November 12, 1980, an agreement was signed with St-Joseph Exploration Ltd. In light of strong gold prices and the excellent outlook, St-Joseph Explorations decided to continue exploring the Kewagama property.

Sulpetro Minerals / Novamin Resources / Breakwater Resources The following summary of the work carried out by Sulpetro Minerals, Novamin Resources and Breakwater Resources on the Kewagama property is mostly modified and summarized from Vaillant and Hutchinson (1982) and Pelchat (1996). 1981: Sulpetro Minerals Ltd (formerly St-Joseph Exploration Ltd) deepened the shaft to 1,150 ft. Ore and waste passes were driven from the 7th level to the 4th level. Thirty-one (31) surface drill holes (2120-S-1 to 2120-S-31) were drilled for a total of 4,789.8 m. Geophysical surveys (Mag, VLF, IP) were carried out on the Kewagama property. Five (5) of the holes were drilled to test a coincident Mag and IP anomaly between lines 3+20E and 4+00E. The result was the discovery of the West IP Zone. 1982: Development continued on the 6th and 7th levels, and the Winze Zone was mined out, producing 11,340 metric tons averaging 3.03 g/t Au. Production also continued from the Q, Rand S veins until operations were suspended in November 1982. 1988: Four (4) surface diamond drill holes (2120-S-32 to 2120-S-35) totalling 1,005.8 m were drilled by Novamin to test the Piché Group "Mine Horizon" lithologies between the O'Brien and Kewagama property boundaries at the westernmost end of the 500' Ievel in the Kewagama underground workings. These holes intersected favourable lithologies that could host ore-grade gold mineralization laterally and at depth. 1994: On July 25, the wooden Kewagama shaft was struck by lightning and burned down. 1995: Breakwater Resources re-activated the exploration activities on the Kewagama property, and established new surveyed grid lines spaced 100 metres apart, with a cumulative length of 16 km. As a first step, a compilation of historical work was completed to better understand the geological setting and assess the economic potential of the Kewagama property. Consequently, geological mapping was conducted over the recently cut grid lines, which covered the entire property. The purpose of this work was to study the lithological and structural controls that govern the distribution of the gold mineralization, and to build the geological compilation map of the Kewagama property.

Radisson Mining Resources The following summary of the work conducted by Radisson Mining Resources on the Kewagama property is mostly modified and summarized from Kelly (2003), Bisson (2004), Barrie (2006), Vincent (2009), David and Gauthier (2012), and Radisson’s annual reports (1999 to 2013). 1999: Radisson became 100% owner of the Kewagama property adjacent to and east of the O’Brien property in 1999. A compilation of existing data began during that same year with the objective of assessing the potential of the existing gold showings.

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2003: Radisson drilled one hole (KW03-01) for 176 m in March 2003. Drilling took place in the western sector of the property to verify the existence of near-surface quartz veins. 2004: An initial deep-drilling campaign was carried out in 2004 to study “Contact Zone”-type gold mineralization on the O’Brien and Kewagama properties. A total of seven (7) holes (KW04-02 to KW04-05, KW04-06C, KW04-02W and KW04-04W) were drilled on the Kewagama property, ranging in length from 690 to 1,580 m for a total of 4,839.1 m. This program studied the favourable horizon to a depth never before explored. The objective was to significantly increase the potential and value of the Company’s holdings by discovering more extensive gold structures at depth, along the Cadillac–Larder Lake Fault Zone (CLLFZ), compared with the known vein system near the surface. 2005: Radisson drilled five (5) holes (KW05-07 to KW05-11) for a total of 3,030.0 m. The purpose of the 2005 drilling program was to investigate the area between Zone 36 East and the Kewagama shaft, at a depth of 460 to 600 m. 2006: A high-resolution aeromagnetic, horizontal gradiometer and XDS-VLF-EM survey was carried out on the O’Brien and Kewagama properties in June 2006. The survey, which was the first phase of the 2006 exploration program, was conducted by Terraquest Ltd with a flight line spacing of 50 m. Data from this survey was used to define drill targets north of the CLLFZ. A diamond drilling program was then carried out on the property. A total of five (5) holes totalling 2,237.0 m (KW06-12 to KW06-16) were drilled on the No. 2 Vein, Zone 36 East and the North Zone. The 2006 drilling program confirmed the discovery of the North Zone, which now extends for more than 300 m along strike, from section 43E to 53E, confirming the potential for gold mineralization north of the CLLFZ. 2008: In the fall of 2008, an exploration drilling program targeted two priority sectors on the Kewagama property: the area between Zone 36 East and the Kewagama mine, and the down-dip extensions of the gold zones below the old Kewagama mine stopes. A total of eleven (11) holes totalling 4,946.8 m were drilled on the property (KW08-151, KW08-154 to KW08-160, KW08-164, KW-08-155A and KW08-163A). Holes KW08-155A, 157 and 158 were drilled in the area between Zone 36 East and the old Kewagama mine. Hole KW08-157 cut a narrow high-grade zone. In addition, several high-grade quartz veins were intersected in the sedimentary rock of the Cadillac Group in hole KW08-155. To the east, in the stratigraphic extension of the O’Brien mine, hole KW08-155A cut a wide low-grade mineralized zone. Seven (7) holes were also drilled on the old Kewagama mine site (KW08-151, 154, 156, 159, 160, 163A and 164). Hole KW08-164, drilled nearly 160 m below the operating levels of the Kewagama mine, intersected a wide highly altered zone. 2011: A total of 13 holes (RM-11-01, RM-11-02, RM-11-05 to RM-11-13, RM-11-15 and RM-11-19) were drilled on the Kewagama property.

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The diamond drilling program led to the discovery of new gold mineralization on the property. This discovery lies in area part of the property that had never been drilled, and demonstrates the property’s potential for additional gold discoveries. The discovery, in the eastern portion of the property, is open along strike and at depth, and was made near the surface. Table 6.5 presents the statistics from Radisson’s drilling campaigns carried out on the Kewagama property between 2003 and 2011. Table 6.6 provides the best results from these campaigns. Table 6.5 ─ Total of holes drilled by Radisson from 2003 to 2011

Table 6.6 ─ Best results obtained from Radisson’s drilling campaigns

Year Number of Holes

Total Length (meter)

2003 1 176.02004 7 4,229.32005 5 3,030.02006 5 2,237.02008 11 4,946.82011 13 4,359.8Total 42 18,978.9

HoleFrom

(meter)

to

(meter)

Corelenght

(meter)

Au g/t

(uncut)

KW04-02W 1,228.8 1,229.8 1.0 17.46KW04-03 515.4 523.6 8.2 5.45KW05-11 564.0 565.5 1.5 5.42KW06-13 183.1 184.1 1.0 10.4KW06-15 202.6 203.8 1.2 10.60KW06-16 443.0 443.8 0.8 20.7KW08-151 554.0 557.0 3.0 4.92

KW08-155A 341.8 347.2 5.4 3.75KW08-156 603.6 605.6 2.0 6.64KW08-157 165.7 166.0 0.3 466.48KW08-164 522.8 535.9 13.1 1,83RM-11-11 132.8 137.0 4.2 2,53RM-11-12 200.9 202.6 1.7 10.00RM-11-13 131.3 132.4 1.0 18.90RM-11-15 215.0 216.0 1.0 11.30

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7. GEOLOGICAL SETTING AND MINERALIZATION

Abitibi Terrane (Abitibi Subprovince) Previously, the Abitibi Greenstone Belt has been subdivided into northern and southern parts based on stratigraphic and structural criteria (e.g., Dimroth et al., 1982; Ludden et al., 1986; Chown et al., 1992). Previous publications used an allochthonous model of greenstone belt development that portrayed the belt as a collage of unrelated fragments. Thurston et al. (2008) presented the first geochronologically constrained stratigraphic and/or lithotectonic map (Fig. 7.1) covering the entire breadth of the Abitibi Greenstone Belt from the Kapuskasing Structural Zone eastward to the Grenville Province. According to Thurston et al. (2008), Superior Province greenstone belts consist of mainly volcanic units unconformably overlain by largely sedimentary Timiskaming-style assemblages, and field and geochronological data indicate that the Abitibi Greenstone Belt developed autochthonously. The Abitibi Greenstone Belt is composed of east-trending synclines largely composed of volcanic rocks and intervening domes cored by synvolcanic and/or syntectonic plutonic rocks (gabbro-diorite, tonalite, and granite) alternating with east-trending bands of turbiditic wackes (MERQ-OGS, 1984; Ayer et al., 2002a; Daigneault et al., 2004; Goutier and Melançon, 2007). Most of the volcanic and sedimentary strata dip vertically and are generally separated by abrupt, east-trending faults with variable dip. Some of these faults, such as the Porcupine-Destor Fault, display evidence for overprinting deformation events including early thrusting, later strike-slip and extension events (Goutier, 1997; Benn and Peschler, 2005; Bateman et al., 2008). Two ages of unconformable successor basins occur: early, widely distributed Porcupine-style basins of fine-grained clastic rocks, followed by Timiskaming-style basins of coarser clastic and minor volcanic rocks which are largely proximal to major strike-slip faults, such as the Porcupine-Destor Fault Zone, the Cadillac–Larder Lake Fault Zone and other similar faults in the northern Abitibi Greenstone Belt (Ayer et al., 2002a; Goutier and Melançon, 2007). In addition, the Abitibi Greenstone Belt is cut by numerous late-tectonic plutons from syenite and gabbro to granite with lesser dykes of lamprophyre and carbonatite. The metamorphic grade in the greenstone belt displays greenschist to sub-greenschist facies (Jolly, 1978; Powell et al., 1993; Dimroth et al., 1983; Benn et al., 1994) except around plutons where amphibolite grade prevails (Joly, 1978). The following more detailed description of the new subdivision of the Abitibi Greenstone Belt is mostly modified and summarized from Thurston et al. (2008) and references therein.

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Figure 7.1 – Stratigraphic map of the Abitibi Greenstone Belt. The geology of the southern Abitibi Greenstone Belt is based on Ayer et al. (2005) and the Québec portion on Goutier and Melançon (2007). Figure modified from Thurston et al. (2008).

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The Abitibi Greenstone Belt is now subdivided into seven discrete volcanic stratigraphic episodes on the basis of groupings of numerous U-Pb zircon ages. New U-Pb zircon ages and recent mapping by the Ontario Geological Survey and Géologie Québec clearly show similarity in timing of volcanic episodes and ages of plutonic activity between the northern and southern Abitibi Greenstone Belt as indicated in Figure 7.1. These seven volcanic episodes are listed from oldest to youngest:

1. Pre-2750 Ma volcanic episode; 2. Pacaud Assemblage (2750-2735 Ma); 3. Deloro Assemblage (2734-2724 Ma); 4. Stoughton-Roquemaure Assemblage (2723-2720 Ma); 5. Kidd-Munro Assemblage (2719-2711 Ma); 6. Tisdale Assemblage (2710-2704 Ma); 7. Blake River Assemblage (2704-2695 Ma).

Cadillac Area

Two types of successor basins are present in the Abitibi Greenstone Belt: early turbidite-dominated (Porcupine Assemblage; Ayer et al., 2002a) laterally extensive basins, succeeded by aerially more restricted alluvial-fluvial or Timiskaming-style basins (Thurston and Chivers, 1990). The geographic limit (Fig. 7.1) between the northern and southern parts of the Abitibi Greenstone Belt has no tectonic significance but is herein provided merely for reader convenience and is similar to the limits between the internal and external zones of Dimroth et al. (1982) and that between the Central Granite-Gneiss and Southern Volcanic zones of Ludden et al. (1986). The boundary passes south of the wackes of the Chicobi and Scapa groups with a maximum depositional age of 2698.8 ± 2.4 Ma (Ayer et al., 1998, 2002b).

The following description of the Cadillac area is mostly modified and summarized from Doucet and Lafrance (2005), and references therein. The Cadillac area is underlain by rocks of the Southern Volcanic Zone of the Abitibi Subprovince intruded by Proterozoic diabase dykes. The Cadillac–Larder Lake Fault Zone (CLLFZ) runs along an E-W axis and separates the Pontiac metasedimentary Subprovince to the south from the Abitibi volcano-sedimentary Subprovince to the north. In Québec, about forty or so gold deposits, which have produced over 60 million ounces of gold since the early 20th century, are associated with this major structure and its subsidiary faults. Intrusive rocks in the Cadillac area include mafic sills (gabbro and diorite) occurring in the Blake River and Piché groups, the synvolcanic Mooshla Pluton, composed of gabbro, quartz diorite, tonalite and trondhjemite, as well as N-S and NE-SW-trending Proterozoic diabase dykes. North of the CLLFZ, regional metamorphism ranges from the greenschist facies to the upper greenschist facies, but the metamorphic grade increases south of the fault to reach the amphibolite facies. From north to south, the following six major lithological units (groups) are observed: Malartic, Kewagama, Blake River, Cadillac, Piché and Pontiac (Figure 7.2).

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The Malartic Group is composed of ultramafic volcanic rocks (komatiites) and tholeiitic basalts (Trudel et al., 1992). The Kewagama Group contains wackes and pelitic rocks. The Blake River Group comprises the Hebecourt and Bousquet formations. The Hebecourt Formation is composed of massive and pillowed basalts, gabbro sills and rhyolites of tholeiitic affinity. According to Lafrance et al. (2003c), the Bousquet Formation includes a lower member and an upper member. The lower member is composed of an intermediate scoriaceous tuff; mafic, intermediate and felsic volcanic rocks; and felsic and mafic subvolcanic intrusions. The upper member consists of massive felsic volcanic rocks and volcaniclastic units. Rocks of the lower member are tholeiitic to transitional, whereas those of the upper member show a transitional to calc-alkaline affinity (Lafrance et aI., 2003c). The Cadillac Group is composed of wackes, pelitic schists with bands of polymictic conglomerate and iron formation. In the Cadillac area, the Piché Group is composed of volcanic rocks (tholeiitic basalts, porphyritic andesites and calc-alkaline block tuffs) interbedded with conglomerates, wackes, graphitic schists and pyritic cherts. Most of the orebodies in the southern part of the Cadillac mining camp are hosted in rocks of the Piché Group, which forms a thin band several tens of kilometres long that follows the trace of the CLLFZ (Fig. 7.2). Sedimentary rocks, mainly wackes, of the Pontiac Group lie south of the CLLFZ. Volcanic and sedimentary rocks in the Cadillac area form a series of E-W-trending steeply dipping monoclonal panels. Volcanic and sedimentary sequences are separated by longitudinal faults parallel to lithological contacts such as the CLLFZ and Lac Imau faults (Figure 7.2).

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Figure 7.2 – Geological syntheses of the Cadillac mining camp with location of active and closed mines, ore deposits and showings. Modified from Lafrance et al. (2003a, 2003b)

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Property Geology The following description of property geology is mostly modified and summarized from Doucet and Lafrance (2005) and Evans (2007), and references therein. The property straddles the Piché Group volcanic rocks that separate Pontiac Group metasedimentary rocks to the south from Cadillac Group metasedimentary rocks to the north. In the property area, all lithologies strike east-west and dip steeply south at approximately 85°. The CLLFZ is a major regional crustal break that consists mainly of chlorite-talc-carbonate ultramafic schist, and ranges in thickness from 100 to 300 ft (30 to 100 m) in the mine area, and narrows significantly to about 40 ft (12 m) wide to the east of Zone 36 East. Across the property, the fault is subparallel and close to the Piché Group-Cadillac Group contact, but is generally enveloped by Cadillac Group sedimentary rocks (argillites, greywackes and, to a lesser extent, chert).

Cadillac Group The Cadillac Group metasedimentary rocks are in the footwall of the mineralization and predominantly in the CLLFZ footwall, and hence the majority of the diamond drill holes did not intersect the Cadillac Group rocks. The limited drill hole intersections show the presence of argillite, greywacke, some pebble conglomerate-like units, and some iron formation.

Piché Group The veins of the O’Brien Project were mostly injected into the volcanic and sedimentary rocks of the Piché Group. From south to north, the Piché Group stratigraphy is divided into the following units:

Southern volcanics: tuff and basaltic schists; Southern porphyritic andesite; Central volcanics: tuff and basaltic schists; Sporadically pebbly greywacke and argillite (“Mine Conglomerate”); Northern porphyritic andesite; Northern volcanics: tuff and mafic schists (with small quantities of argillite,

greywacke, chert and massive to variably porphyritic basalt flows). All the above lithologies generally strike east-west with more pronounced flexures locally. The rock varies from slightly to highly schistose and foliation increases progressively towards the CLLFZ.

7.3.2.1 Porphyritic andesite The southern and northern porphyritic andesites are much alike. They are characterized by abundant quartz eyes ranging in size from 0.1 to 0.5 cm, and range in colour from greyish to buff-beige, set in an aphanitic to fine-grained matrix of intermediate composition. In general, the andesites are intensely sheared and show a more or less brownish biotite and chlorite alteration. The strong foliation often produces an augen texture with quartz phenocrysts. The latter units are continuous horizontally and vertically in the 36E area, and are useful stratigraphic marker

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horizons. The north and south porphyritic andesite units are thicker in the vicinity of the O’Brien mine. It is unclear whether these units are duplicated by folding and faulting. The south porphyritic andesite generally hosts the PC and PN veins, and the north porphyritic andesite is spatially associated with the IN Vein.

7.3.2.2 Conglomerate The O’Brien mine conglomerate is represented in the 36E area by well-bedded greywacke and argillite with the sporadic presence (2% to 5%) of greyish granitic pebbles and other components. The pebbles tend to be somewhat flattened, consistent with north-south compression. The IS Vein is located mainly in this relatively competent lithology. The conglomerate unit is another useful marker horizon.

7.3.2.3 Volcanic rocks The volcanic rocks consist mainly of mafic tuffs and flows. The volcanic rocks generally have tholeiitic signatures (Trudel et al., 1992). In general, the flows are fine grained and exhibit greenschist facies mineral assemblages. The tuffs are of mafic composition and are very abundant. The tuffs can be finely bedded to very schistose. Locally present is massive, fine-grained basalt or lesser amounts of gabbro and amphibolites. Schistosity is more developed in the central and northern volcanic units than the southern unit. Greywacke and argillite lenses occur more frequently between the volcanic rocks in the northern units. The southern volcanic rocks contain the PS Vein. The central volcanic rocks are locally mineralized by the PN and IS veins. The north volcanic unit and sediment interlayers host the IX, FS, and FV veins.

7.3.2.4 Graphitic schist Thin layers of graphitic schist and argillite are present. These are highly sheared and deformed, characterized by tight folding, and often display breccias or slickensides with graphite. Pyrite is abundant, finely laminated and deformed.

Pontiac Group The Pontiac Group metasedimentary rocks consist mainly of greywacke and some argillite, which is sometimes graphitic. In general, the sediments are well stratified. Some zones display weak biotitic alteration or chloritization. Small-scale folding is observed in places. Some greyish to smokey quartz veins and veinlets, similar to gold-bearing veins, appear locally, and some of these host gold (OB-95-48, 52, 53, 54 and 56A).

Mineralization The following description of mineralization is mostly modified and summarized from Evans (2007), and references therein.

O’Brien mine Gold production at the O’Brien mine came from a few quartz veins running almost parallel to the formations. The mine’s productive sector was generally limited to a narrow strip that included the O’Brien conglomerate and the northern porphyritic andesite. Approximately 95% of the O’Brien ore came from four veins (No. 1, No. 4,

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No. 9 or “F”, and No. 14) in the eastern part of the mine. The veins contained high-grade shoots that occasionally yielded considerable amounts of visible gold. The main veins generally strike from 083° to 098°, and dip steeply to the south (-84° to -90°). The stopes averaged 0.75 to 0.90m (2.5 ft to 3 ft) wide. Gold mineralization extends vertically down to at least the 3450' level.

7.4.1.1 No. 1 Vein The No. 1 Vein was the most productive in terms of tonnage and occurs mainly in the conglomerate. This vein comprises No. 1 Vein NE-SW (080º to 090º azimuth) and No. 1 Vein NW-SE (090º to 095º azimuth). No. 1 Vein NE-SW extends from surface to at least the 3000' level and is over 500 ft in strike length. The richest and most productive portion of this vein was from a 50 to 200 ft long shoot (15 to 60 m) that plunges about 85º to the east from about the 750' level down to at least the 3000' level, at its intersection with Vein No. 1 NW-SE, at the conglomerate hanging wall contact. A second moderate-grade shoot, about 50 to 150 ft long (15 to 45 m) plunges about 60º to the east from about the 1000' level to the 2500' level. Vein No. 1 NW-SE extends from about the 750' level to at least the 3450' level, and ranges in horizontal length from about 50 to 600 ft (15 to 180 m). Higher grade shoots plunging about 85º to the east seem to be controlled by vein intersections and vein folds. Both of these veins average 30 cm thick (Mills, 1950).

7.4.1.2 No. 4 Vein The No. 4 Vein is spatially associated with the north porphyritic andesite. It extends from surface down to at least the 3450' level, and has a 1,000 ft strike length. It averaged 30 cm thick (Blais, 1954). Approximately 50% of the gold produced came from this vein. This was due to an exceptionally high grade ore shoot, only 30 to 50 ft long (9 to 15 m) horizontally, but which extended for 625 ft (190 m) from the 500' level down to the 1125' level.

7.4.1.3 No. 9 Vein The No. 9 Vein is located in the northern greywacke and volcanic units. This brown vein is rich in biotite and arsenopyrite. It is also wider than the others. The stopes were rarely less than 4 ft (1.2 m) wide, and could reach 20 ft (6 m) in certain folded zones where visible gold was common. It was mined out from the 1250' level down to the 1375' level along a horizontal length of about 160 ft (50 m).

Zone 36E area The main mineralized structures (“veins”) are generally narrow, ranging in true thickness from a few inches to 22 ft (6.7 m), but have good continuity both horizontally and vertically. Gold-bearing veins occur in different lithologies of the Piché Group and the Pontiac Group. The veins cross the stratigraphy at low angles and are occasionally folded, particularly in volcanic and argillic host rocks. Generally, the veins strike east-west (085° to 097°), dip steeply to the south (-80° to -90°), and contain higher grade shoots that plunge steeply to the east.

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After the 1994-95 drilling program, Radisson completed a new geological interpretation that retained eight of the ten veins (structures) defined by Novamin. These veins were, from south to north: PS, PC, PN, IS, IN, IX, FS and FV. They were located in a 250 ft (75 m) wide corridor within the Piché Group metasedimentary rocks and metavolcanic rocks, and were observed to be best developed between 3,200E and 4,400E. For the current report, InnovExplo completed a 3D geological interpretation that allowed more structures to be identified than before. Often, the veins occur as a group of quartz veinlets scattered in a very sheared and altered zone that has no obvious main vein. Only very competent lithologies, like the conglomerate and the porphyritic andesites, host large veins. In some drill core, the quartz veinlets exhibit small tight folds (Bisson, 1995). Gold grades vary considerably. The gold occurs mainly as fine to coarse free grains that are heterogeneously distributed, mainly in the quartz veins, and to a lesser extent, in the wallrock. Higher gold grades occur in short, steeply plunging shoots with a similar style to those mined at the O’Brien mine (Bisson, 1996). The colour of the gold-bearing quartz veins varies from milky to greyish to dark smokey, and sometimes individual veins contain all three colours in varied proportions. The quartz veins are narrow and range from less than 1 in to over 3 ft wide. The quartz is generally very deformed and brecciated. The veins sometimes contain altered mineralized wallrock xenoliths.

Kewagama area The following description of the Kewagama mine is mostly modified and summarized from Dresser and Denis (1949), and references therein. The gold mineralization occurs in rocks of the Piché Group, south of the CLLFZ, which strikes east-west in this area and dips at 80° to 85° to the south. North of the CLLFZ lies a considerable width of tuffs and agglomerates. In the vicinity of the mine workings, the highly sheared rocks of the Piché Group have an aggregate width of 100 to 130 m. The succession from north to the south is as follows: greenstone (15 to 25 m); “North” porphyry (3 to 10 m); conglomerate (12 to 25 m); greenstone and tuffs (3 to 7 m); “South” porphyry (3 to 9 m); and greenstone (about 60 m). The only gold mineralization of particular interest disclosed by extensive underground workings is in the winze, in a 25-ft raise above the winze, and in the sublevels driven from the winze. These workings revealed an ore shoot with a vertical extent of 70 m and an east-west length of 4.5 to 25 m, in which irregular and discontinuous stringers of blue quartz carry free gold. The majority of these veins are parallel and are contained within the “North porphyry near its north margin, but some continue into the greenstone north of the porphyry. Individual veins are rarely more than 10 cm wide and 3 m long; occasionally, two or three are parallel to one another or overlap for part of their length. Some sections of these narrow veins are decidedly high grade, but in any stoping operation there would be considerable dilution.

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The Kewagama ore shoot described above occurs in the same rocks as the high-grade shoot in the historical No. 4 vein mined at the O’Brien mine, and resembles it for its short lateral extent compared to vertical, and for the fact that it contains the same type of blue quartz and associated minerals. It differs from the O’Brien shoot in that it does not follow one definite fracture, instead consisting of a series of irregular overlapping stringers, and for the fact that it is of much lower grade as a whole.

Hydrothermal Alteration The following description of hydrothermal alteration is mostly modified and summarized from Evans (2007), and references therein. Wallrock alteration ranges from a few inches to several feet thick, equally pervasive on both sides of the veins. The mineralized zones are usually comprised of a greater proportion of altered wallrock than actual veins. In general, the wallrock is well foliated and has a distinctive dark brown to brownish grey colour due to intense biotite alteration. The brownish alteration is an easily recognizable indicator of potential gold-bearing mineralization. Biotite tends to occur as 1 to 2 mm thick layers of predominantly fine-grained biotite parallel to the foliation. On average the mineralized zones contain about 5% biotite, but can contain over 20% biotite. Generally, zones of biotite alteration accompanied by silicification and sulphidation will yield gold values. Of all the sulphides, arsenopyrite is the most abundant and characteristic of the O’Brien mine. Arsenopyrite occurs mainly in intensely altered wallrock where it can be abundant (2% to 10%). The finer grained and needle-like varieties of arsenopyrite are more likely to contain gold. Coarser grained, euhedral rhombic arsenopyrite is less likely to contain gold (Bisson, personal communication 1998). Fine- to medium-grained, subhedral to euhedral pyrite is frequently observed generally overprinting the foliation (0.5% to 2%). Some pyrite is associated with gold-bearing zones (Hatch, 1998). Minor quantities of pyrrhotite and chalcopyrite are present in the mineralized zones (Bisson, 1995). Carbonate alteration is mainly calcitic in microveinlet form, but it is also found frequently in all lithologies as more massive pervasive replacement. At times, iron carbonate veinlets are visible. Tourmaline is frequently but not always seen. It is generally found in small amounts in association with wallrock xenoliths.

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8. DEPOSIT TYPES

Greenstone-hosted quartz-carbonate vein deposits occur as quartz and quartz-carbonate veins, with valuable amounts of gold and silver, in faults and shear zones located within deformed terranes of ancient to recent greenstone belts commonly metamorphosed at greenschist facies (Dubé and Gosselin, 2007). Greenstone-hosted quartz-carbonate vein deposits are a subtype of lode gold deposits (Poulsen et al., 2000) (Fig. 8.1). They are also known as mesothermal, orogenic. They consist of simple to complex networks of gold-bearing, laminated quartz-carbonate fault-fill veins in moderately to steeply dipping, compressional brittle-ductile shear zones and faults, with locally associated extensional veins and hydrothermal breccias. They can coexist regionally with iron formation-hosted vein and disseminated deposits, as well as with turbidite-hosted quartz-carbonate vein deposits (Fig. 8.2). They are typically distributed along reverse-oblique crustal-scale major fault zones, commonly marking the convergent margins between major lithological boundaries such as volcano-plutonic and sedimentary domains. These major structures are characterized by different increments of strain, and consequently several generations of steeply dipping foliations and folds resulting in a fairly complex geological collisional setting.

Figure 8.1 – Inferred crustal levels of gold deposition showing the different types of lode gold deposits and the inferred deposit clan (from Dubé et al., 2001; Poulsen et al., 2000)

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Figure 8.2 – Schematic diagram illustrating the setting of greenstone-hosted quartz-carbonate vein deposits (from Poulsen et al., 2000) The crustal scale faults are thought to represent the main hydrothermal pathways towards higher crustal level. However, the deposits are spatially and genetically associated with higher order compressional reverse-oblique to oblique brittle-ductile high-angle shear zones commonly located less than 5 kilometres away and best developed in the hanging wall of the major fault (Robert, 1990). Brittle faults may also be the main host to mineralization as illustrated by the Kirkland Lake Main Break; a brittle structure hosting the 25 Moz Au Kirkland Lake deposit. The deposits formed typically late in the tectonic-metamorphic history of the greenstone belts (Groves et al., 2000) and the mineralization is syn- to late-deformation and typically post-peak greenschist facies and syn-peak amphibolite facies metamorphism (cf. Kerrich and Cassidy, 1994; Hagemann and Cassidy, 2000). Stockworks and hydrothermal breccias may represent the main host to the mineralization when developed in competent units such as granophyric facies of gabbroic sills. Due to the complexity of the geological and structural setting and the influence of strength anisotropy and competency contrasts, the geometry of the vein network varies from simple such as the Silidor deposit, Canada, to more commonly fairly complex with multiple orientations of anastomosing and/or conjugate sets of veins, breccias, stockworks and associated structures (Dubé et al., 1989; Hodgson, 1989, Robert et al., 1994, Robert and Poulsen, 2001). Ore-grade mineralization also occurs as disseminated sulphides in altered (carbonatized) rocks along vein selvages. Ore shoots are commonly controlled by: 1) the intersections between different veins or host structures, or between an auriferous

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structures and an especially reactive and/or competent rock type such as iron-rich gabbro (geometric ore shoot); or 2) the slip vector of the controlling structure(s) (kinematic ore shoot). For laminated fault-fill veins, the kinematic ore shoot will be oriented at a high angle to the slip vector (Robert et al., 1994; Robert and Poulsen, 2001). At the district scale, the greenstone-hosted quartz-carbonate-vein deposits are associated with large-scale carbonate alteration commonly distributed along major fault zones and associated subsidiary structures (Dubé and Gosselin, 2007). 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). 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 ore. The main ore minerals are native gold with pyrite, pyrrhotite and chalcopyrite without significant vertical zoning. (Dubé and Gosselin, 2007)

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9. EXPLORATION

InnovExplo Inc. was retained by Radisson to define, classify and recommend drill targets in order to add potential value to the O’Brien Project (Richard and Fallara, 2015). In collaboration with Radisson’s representatives, three types of targets were defined before starting this mandate. Discussions were also held with InnovExplo engineers working on the PEA study, and this led to the identification of six areas where efforts should be focused (Fig. 9.1). Areas were prioritized based on their spatial distribution with respect to historical or planned underground workings. In the short term, Area 1 is more likely to impact project economics than Area 6. It is important to note that targets were defined, classified and prioritized based on the likelihood they would add resources. Although all the proposed drill holes aim for geologically sound targets, they might have been prioritized differently had they been based solely on geological parameters, or if the goal had been to add resources regardless of constraints imposed by economic factors (e.g., proximity to infrastructure).

Type 1 targets Type 1 targets represent the possible extensions of currently identified ore shoots or likely candidates for new ore shoots. Forty-one (41) targets were defined in six different areas referred to as Area 1 to Area 6. For maximum efficiency, drilling was designed to investigate the entire corridor of mineralization rather than single zones; consequently, some holes were lengthened to intersect adjacent zones beyond the intended target.

Type 2 targets Type 2 targets represent the possible extensions of already identified resources in stopes being designed as part of the PEA study. Stope designs are preliminary, but are based on a previously published resource estimate. Forty-seven (47) targets were situated in close proximity to preliminary infrastructure planned in the PEA study.

Type 3 targets Type 3 targets represent exploration targets outside the resource area that may help improve the knowledge of the property and identify new mineralized zones. Type 3 targets are all located outside the large box shown in Fig. 9.1. Twelve (12) exploration target were identified.

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Figure 9.1 – 3D view looking NNE, showing the different areas defined by Richard and Fallara (2015). All the historical infrastructure of the old Kewagama mine, parts of the infrastructure of the old O’Brien mine, and proposed infrastructure from the current PEA are shown. The outlines of the old O’Brien mine stopes are also shown (grey and red shapes to the west). Note that the areas presented on this figure are approximate.

Area 1 Area 2

Area 4

Area 3

Area 5

Area 6

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10. DRILLING

In mid-December 2015, Radisson began a surface diamond drilling program at the O’Brien Project (see Radisson press release dated December 4, 2015). The drilling program is still ongoing. The drilling program consists of 6,200 metres based on the target definition and drill program proposal carried out by InnovExplo and discussed in Section 9 (Exploration). The drilling program targets areas 1 to 5 (see Fig. 9.1) with the purpose of extending known ore shoots and the likely possibility of defining new ore shoots. To date, no gold results have been reported by Radisson from this drilling program.

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11. SAMPLE PREPARATION, ANALYSIS, AND SECURITY

No exploration work or drilling has been done by Radisson since the latest mineral resource estimate in 2013. Sample preparation, analysis and security protocols for the previous exploration program are discussed in de l’Etoile and Salmon (2013).

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12. DATA VERIFICATION

The diamond drill hole database used for the 2015 Resource Estimate presented herein was provided by Radisson. It is referred to as the “Radisson database” in this section. No drilling was underway at the time this report was being produced, and the latest drilling program took place in 2012, before the previous NI 43-101 report on the project in 2013. InnovExplo’s data verification included visits to the project’s office, as well as to the logging and core storage facilities. It also included a review of selected core intervals, drill hole collar locations, assays, the QA/QC program, downhole surveys, information on mined-out areas, and the descriptions of lithologies, alteration and structures. Site visits were completed by Pierre-Luc Richard on January 19 and January 27, 2015.

Historical Work The historical information used in this report was taken mainly from reports produced before the implementation of NI 43-101. In some cases, little information is available about the sample preparation and analytical protocols or the security procedures implemented for the historical work in the reviewed documents. However, InnovExplo assumes that the exploration activities conducted by earlier companies were in accordance with prevailing industry standards at the time.

Radisson Database InnovExplo was granted access to the certificates of assays for all holes in the latest drilling programs, as well as to all logs for historical holes. Assays were verified for more than 5% of the drill holes from these programs. Special care was taken to validate at least 5% of all individual drilling programs over the years, and not simply 5% of the entire database. Minor errors of the type normally encountered in a project database were addressed and corrected. The final database is considered to be of good overall quality. InnovExplo considers the Radisson database for the O’Brien Project (Kewagama and 36E areas) to be valid and reliable. The reader should be aware that the historical O’Brien mine area was not validated. However, no resource has been established for that part of the property.

Radisson Diamond Drilling All surface drill hole collars on the O’Brien Project (resource area) were either professionally surveyed or surveyed using a GPS unit. The collar surveys are considered adequate for the purpose of a resource estimate, although any collar that was only surveyed using a GPS unit should be professionally surveyed. Underground drill holes were compiled by Radisson and slightly adjusted based on the underground void model compiled by InnovExplo.

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Downhole surveys were conducted on the majority of the holes. Tropari, Acid, and Flexit survey information was verified for 5% of all drill holes from the database. A visual verification was performed on 100% of the downhole surveys, and some modifications were made to the database.

Radisson Logging, Sampling and Assaying Procedures The author (Pierre-Luc Richard) reviewed several sections of mineralized core while visiting the onsite core logging and core storage facilities (Figs. 12.1 to 12.4). All core boxes were labelled and properly stored inside or outside. Sample tags were still present in the boxes, and it was possible to validate sample numbers and confirm the presence of mineralization in the half-core reference samples from the mineralized zones. No drilling was underway at the time of the author’s site visit; it was thus not possible to review the path of the drill core from the drill rig to the logging and sampling facility and finally to the laboratory. However, discussions with on-site personnel allowed the author to establish that the protocols in place while drilling was underway were adequate.

Figure 12.1 – Photo of the logging facility building taken during a site visit in January 2015

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Figure 12.2 – Photo of the indoor core storage facilities taken during a site visit in January 2015

Figure 12.3 – Photo of the outdoor core storage facilities taken during an earlier site visit by InnovExplo in 2014

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Figure 12.4 – Photo of the sample preparation facility taken during a site visit in January 2015

Mined-out Voids Underground workings were compiled and updated since the latest resource estimate. In order to take into consideration adequate depletion due to historical mining, all shafts, galleries, raises and stopes within the resource area were modelled and used to update the interpretation of the mineralized zones. Note that workings from the old O’Brien mine area were not compiled at the time this resource estimate was being produced. InnovExplo considers the refinement of the voids triangulation to be of good quality and reliable, despite the fact that some uncertainties remain.

Conclusion Overall, InnovExplo is of the opinion that the data verification process demonstrated the validity of the data and protocols for the Kewagama and 36E areas of the O’Brien Project. InnovExplo considers the Radisson database to be valid and of sufficient quality to be used for the mineral resource estimation herein.

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13. MINERAL PROCESSING AND METALLURGICAL TESTING

There are many historical documents relating to the O’Brien Project area. Several test programs have been carried out since the 1970s. These were executed by various laboratories. The relationship between historical results and the area that is being studied is complex. Most of the time, samples were identified under the name of the zone. However, these names have changed over time, depending on which company owned the deposit. Nevertheless, these data provide an overview of the mineralogy, treatment methods and gold recoveries that may be obtained for samples taken from this area. The O'Brien Project, as currently defined, covers the 36E and Kewagama areas. The 36E area is divided into four zones: Upper West, West Central, West and Lower Central. The Kewagama area covers the eastern sector. In 2014, new laboratory testwork was undertaken on samples from the 36E area by the URSTM (Bouzahzah et al., 2014). A summary of historical results and results achieved in 2014 for the 36E area will be presented in the following sections.

Historical Data Darius mill

In the 1970s and early 1980s, ore from this mining area was treated at the Darius mill (Figure 1).

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Figure 13.1 – Darius mill flowsheet

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The ore was crushed and grinded. Then the pulp was sent to gravity separation to produce a gold concentrate which could be sent directly to smelting. The tails were submitted to flotation. Gold from the flotation concentrate was extracted by cyanidation. Production reports indicate that ore processed between 1979 and 1982 came from two areas (Table 13.1). Ore from the East Zone decreased gravity recovery and the overall recovery. Table 13.1 – Ore processed between 1979 and 1982

Year 1979 1980 1981

Ore origin No East Zone Mixed with 23% East Zone

Mixed with 30% East Zone

Gravity recovery

% 27.3 10.7 No gravity recovery

Overall recovery

% 76.2 77.6 71.7

The gold recovery by flotation approximated 90%, while the recovery by cyanidation approximated 80%. Gravity separation results suggest that the proportion of free gold is lower for the East Zone. The gravity circuit was not used during the last period, which may have had an impact on the overall recovery. In early 1982, ore from the East Zone was milled. A very small amount of free gold was recovered in the gravity circuit and the total recovery was about 67.4%.

Review of historical testwork Several historical laboratory reports are available. This section presents the main elements. In the 1970s, tests were conducted principally to increase gold recovery. Combinations of gravity, flotation with or without concentrate regrind, and cyanidation of both flotation concentrate and tailings, provided an overall recovery of 85%. The cyanidation of flotation tailings increased the overall recovery by 11.4%. Further testwork using gravity, flotation, regrind of the concentrate and carbon-in-leach process yielded a gold recovery of 92%. In the early 1980s, Kewagama ore was submitted to flotation and cyanidation. Regrind of the flotation concentrate increased the overall recovery from 78% to 87%. During the same period, laboratory work identified that the gold contained in the ore occurred as liberated gold (25%), in the gangue (50%) and tied up with arsenopyrite (25%). The weight distribution in sulphide minerals was as follows: pyrite (46.75%), pyrrhotite (29.5%) and arsenopyrite (23.4%).

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In further work, the Darius mill ore feed was tested to optimize gold recovery. The flotation gold recovery ranged around 95%. The Bond ball mill work index was 12.8 kWh/t. Other tests were carried out in the mid-1990s on two different samples: one from the SURFACE stockpile OF THE ESAT ZONE OF OLD O’BRIEN MINE and drill core from the 36-East-Zone. The tests compared flotation, cyanidation and pre-treatment options on both samples. Flotation recoveries were respectively 92% and 94%. Cyanidation results were 47% for the stockpile and 85% for the 36-East-Zone. Two pre-treatment options were tested: roasting and bio-oxidation. The roasting of both samples increased the gold recovery to 90%. The bio-oxidation results were inconclusive. In summary, the sulphide flotation process provides good gold recovery. The results range between 90 and 96.5%. When gravity is included upstream, the recovery is around 95%. The arsenic content in the concentrate is high. This minor element could be problematic to sell the concentrate and is likely to increase the smelting cost. Several tests have been made to extract gold by leaching the flotation concentrate. The results fluctuated from 64% to about 80%. Longer leaching time, smaller grind size and addition of activated carbon all increased the cyanidation recovery. Investigations have shown that a maximum of 80% of the gold could be recovered by cyanidation. The gold particles in pyrite are present as fine inclusions. Gold in inclusions is difficult to leach as the exposed surface is low or nonexistent. Table 13.2 – Summary of gold recoveries based on laboratory results for each extraction method tested Gold recoveries

Gravity / Gravity + Amalgamation 23-75% Flotation 90-95% Gravity + Flotation 95% Cyanidation of concentrate 53-78% Gravity + Amalgamation + Flotation + Cyanidation 85%

Gravity + Flotation + Fine regrind + CIP 92%

Zone 36E AREA Testwork In 2014, the URSTM carried out a series of analyses on samples from the 36E area of the O’Brien Project (Bouzahzah et al., 2014). The objective was to define a process flowsheet and the results are presented in Table 3. The main composite was formed of six samples. The samples were dried then crushed to a particle size of less than 8 mesh. The composite was homogenized and rotary split into half (0.5)-kilogram charges.

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The composite sample consisted mainly of pyrite and arsenopyrite. The gold head grade was calculated for each test and ranged from 9.97 g/mt to 14.59 g/mt. Table 13.3 – Calculated gold head grade

Test number Calculated head grade (g/mt)

KN-F-1 9.97 KN-F-2 10.39 KN-F-3 10.57 KN-F-5 14.59

KN-F-6-R 11.00 KN-CN-F-4 11.23 KN-CN-2 11.79

It should be noted that the samples which were used to test the 36E area were prepared by the client and WSP could not determine whether such samples are representative of the deposit.

Gravity separation Four gravity separation tests were performed at different grind sizes (137, 105, 90 and 74 μm K80). The tests consisted in processing a pre-ground sample in a laboratory Knelson unit. The concentrate thus obtained was then treated on a Mozley table. The recovery results are displayed in Table 13.4. Table 13.4 – Gravity recoveries

K80 Gold recovery

µm %

137 50.4

105 58.9

90 59.0

74 60.2

Gold recovery increases as the particle size decreases, although the difference beyond 105 microns is not significant.

Flotation grind size Four rougher flotation tests were done on the sample. The objective was to evaluate the grind size for the subsequent test program.

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The results presented in Table 13.5 show that a grinding of 80% passing (K80) 73 microns gives the highest recovery. Table 13.5 – Flotation recovery and grind size

K 80 Gold recovery Mass recovery µm % % 139 90.2 10.4 105 94.0 8.8 73 95.8 11.1 37 94.9 18.0

Combination of gravity and flotation Three tests combining gravity separation followed by a rougher flotation were made. It should be noted that only the third test had a cleaning step. All the samples were ground to K80 102 µm then fed to a Knelson unit and a Mozley table. The combined tails were submitted to flotation. Flotation was conducted in Denver lab cells with Potassium Amyl Xanthate (PAX) as the collector and Methyl Isobutyl Carbonyl (MIBC) as the frother. In test 2, the gravity tails were reground to 73 microns before being floated. In test 3, a cleaning step was added. As presented in Table 13.6, the combined recoveries obtained for tests 1 to 3 were respectively 93.6%, 93.4% and 94.4%. Table 13.6 – Summary of gravity and flotation gold recoveries

Gravity Flotation Total K 80 Gold

recovery Mass

recovery Regrind

K80 Gold

recovery Mass

recovery Gold

recovery µm % % % % % %

Test 1 102 55.3 0.03 NA 38.3 8.69 93.6

Test 2 102 54.3 0.06 73 39.1 9.35 93.4

Test 3 102 62.9 0.03 NA 31.5 3.67 94.4

Regrind of gravity tails does not seem to have had an impact on the recovery, as opposed to the preliminary results that were obtained.

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Test 3 yielded the best gold recovery. The cleaning step reduced the mass of concentrate by more than 50%. With the same test condition, the gold recovered by gravity is 8% higher. This influences the final recovery obtained. Additional tests will be needed to validate these results. An arsenic mass balance was done for test 3. Table 13.7 displays the results. Table 13.7 – Arsenic mass balance

Mass recovery As grade As recovery % % %

1st cleaner 1.82 20.2 71.4 Cleaner scavenger 1.85 5.51 19.8 Tailings 96.33 0.05 8.8 Total 100 0.52 100

The arsenic content in the concentrate is 12.79%. This arsenic value is high and will be decisive in the economic evaluation.

Cyclic flotation tests Two locked-cycle tests (4 cycles) were performed to produce metallurgical projection results when one or several products are recirculated in closed loop. The same gravity conditions and flotation reagents were used for these tests. Table 13.8 present the results of the cyclic tests. In the first test, the cleaner tails were returned to the main feed. The cleaner scavenger concentrate was sent to the first cleaner. For the second cyclic test, the cleaner tails were sent to the cleaner scavenger feed. The cleaner scavenger concentrate was returned to the cleaner feed. Table 13.8 – Summary of cyclic tests

K 80 Gravity Flotation Total µm % % %

Cyclic test 1 102 67.4 26.2 93.6 Cyclic test 2 102 60.2 34.4 94.6

The overall recovery is higher for cyclic test 2. The results are similar to those obtained in test 3 of the previous series of tests combining gravity and flotation, which included a cleaning step (94.4%). Additional tests must be performed to determine the best configuration. The choice and the amount of reagent that is used can also be optimized.

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Combination of gravity and cyanidation Two separation tests were performed at 102 micron grind sizes in a Knelson unit. The concentrate thus obtained was then submitted to a Mozley table. Both gravity concentrator and Mozley table tailings were leached with sodium cyanide at pH 11 for a 48-hour period. The gravity tailings were reground at 37 microns before leaching for the second test. The highest overall gold recovery obtained was 92.9% with the 37 μm grind size. A summary of the test conditions and the recovery results is displayed in Table 13.9. Table 13.9 – Summary of gravity and cyanidation test results Test 1 Test 2

Gravity Grind size µm 102 102

Au recovery % 58 60.8

Cyanidation Grind size µm 102 37

Au recovery 24 h % 29.9 30.4

Au recovery 36 h % 30.5 31.5

Au recovery 48 h % 31.6 32.1

Total 89.6 92.9

Reagent consumption

Ca(OH)2 kg/mt 2.08 3.19

NaCN kg/mt 0.33 0.49

There was a difference in the gravity recoveries between the two tests. Cyanidation recovery increases slightly with the particle size reduction. Gold recoveries from the flotation flowsheet were higher. With the latter, gold recoveries ranged from 93.4 to 94.6% but the arsenic content in the concentrate was substantial. The recovery for the cyanidation flowsheet varied from 89.9 to 92.9%. In all cases, the recovery by gravity process is significant and is around 60%.

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14. MINERAL RESOURCE ESTIMATES

The 2015 O’Brien Mineral Resource Estimate herein was prepared by Pierre-Luc Richard, M.Sc., P.Geo., with contributions from Alain Carrier, M.Sc., P.Geo., using all available information (Richard et al., 2015). The main objective of the mandate assigned by Radisson was to update the 2013 Mineral Resource Estimate prepared by RPA and published in a report titled “Technical Report on the O’Brien Project Mineral Resource Estimate, Québec, Canada” (compliant with National Instrument 43-101 and Form 43-101F1) (de l’Étoile and Salmon, 2013). The main reason for the update was the addition of additional ground. The 2013 resource estimate focused solely on the 36E area, whereas the resource estimate presented herein includes the Kewagama area. The Kewagama area was mined in the past as the Kewagama mine, and the 36E area was partially mined as extensions of either the Kewagama or O’Brien mines. The 2015 resource area measures 2.1 km along strike, 0.6 km wide and 0.7 km deep. The resource estimate is based on a compilation of historical and recent diamond drill holes and a litho-structural model constructed by InnovExplo. The mineral resources presented herein are not mineral reserves as they have no demonstrable economic viability. The result of this study is a single Mineral Resource Estimate for 55 gold-bearing zones and two low-grade dilution envelopes (see below for details). The estimate includes indicated and inferred resources for an underground scenario. The effective date of the estimate is April 10, 2015, based on compilation status and cut-off grade parameters.

Drill Hole Database The GEMS diamond drill hole database contains 310 surface diamond drill holes and 1,815 underground drill holes. From these, a subset of 620 holes (279 from surface and 341 from underground) located inside the limits of the resource estimate area were used in this Mineral Resource Estimate, representing the drill holes that had been compiled and validated at the time the estimate was being initiated (Figure 14.1). The majority of the 620 holes contain lithological (n = 566), alteration (n = 255) and structural (n = 206) descriptions taken from drill core logs. A total of 467 holes (63,399 m) contain samples assayed for gold, leaving 153 holes (4,412 m) without any samples in the database. Note that many unsampled holes were drilled in overburden. The 620 drill holes cover the strike-length of the project at a variable drill spacing ranging from 10 to 60 m. This selection of 620 drill holes contains a total of 30,283 sampled intervals taken from 67,811.34 m of drilled core. In addition to the basic tables of raw data, the GEMS database includes several tables containing the calculated drill hole composites and wireframe solid intersections required for the statistical evaluation and resource block modelling.

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43-101 Technical Report – O’Brien Project

Figure 14.1 – Surface plan view of the O’Brien drill hole database. Top: All drill holes in the database (n = 2,125); Bottom: validated holes in the 36E and Kewagama areas used for the 2015 resource estimate (n = 620).

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Interpretation of Mineralized Zones The 2013 model needed to be reviewed in light of the updated database and ongoing compilation work. In order to conduct accurate resource modelling of the deposit, InnovExplo based its mineralized-zone wireframe model on the drill hole database and the author’s knowledge of the O’Brien mine. A total of 4,215 construction lines (1,372 3D rings and 2,843 tie lines) were created in order to produce valid solids. InnovExplo created a total of 55 mineralized solids (coded 101 to 230) that honour the drill hole database. Although currently considered as individual mineralized zones, it is likely that additional work on the property will eventually link some zones that have been broken up by faults. Most of the mineralized zones are included within the dilution envelopes (coded 501 and 502), which were also created by InnovExplo. Overlaps were handled by the “precedence” system used by GEMS for coding the block model. Two surfaces were also created in order to define topography and overburden. These surfaces were generated from drill hole descriptions. Figure 14.2 presents a 3D view of the 55 mineralized solids.

Figure 14.2 – 3D view looking northeast of the 55 mineralized solids.

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Underground Workings Underground workings were compiled and updated since the latest resource estimate. In order to take into consideration adequate depletion due to historical mining, all the shafts, galleries, raises and stopes within the resource area were modelled and used to update the interpretation of the mineralized zones. Note that workings from the old O’Brien mine area were not compiled at the time this resource estimate was being produced. These workings were coded within the block model, and depletion was conducted adequately. Figure 14.3 presents a 3D view of the underground workings considered in the 2015 resource estimate.

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Figure 14.3 – 3D view looking northeast of the underground workings in the 36E and Kewagama areas in relation to resource blocks (red). Note that the compilation of the underground workings to the west (old O’Brien mine) is incomplete.

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High Grade Capping For drill hole assay intervals that intersect interpreted mineralized zones, codes were automatically attributed based on the name of the 3D solids, and these coded intercepts were used to analyze sample lengths and generate statistics for high grade capping and composites. Basic univariate statistics were performed on four (4) raw assay datasets consisting of mineralized zones and dilution envelopes both to the east and west of a major fault crosscutting the deposit. The number of samples for each dataset were as follows: 3,176 (mineralized zones east of the fault), 6,408 (mineralized zones west of the fault), 5,694 (dilution envelope east of the fault) and 11,888 (dilution envelope west of the fault). A total of 59 samples from the mineralized zones and 21 from the dilution envelopes were capped at capping limits varying from 3.5 g/t Au to 65 g/t Au. The capping of high assays affected 0.30% of all samples within the block model. Table 14.1 presents a summary of the statistical analysis for each dataset. Figures 14.4 to 14.7 present graphs supporting the gold assay capping values. Table 14.1 – Summary statistics for the raw assays by dataset

Number Max Uncut High Cut # % % Loss

of (Au g/t) Mean Grade Mean Samples Samples Metal Factor

Samples (Au g/t) Capping (Au g/t) Cut Cut

Mineralized zones west of the fault 101 to 128 6,408 1,019.14 2.30 65.00 1.76 23 0.36% -14.99%

Mineralized zones east of the fault 201 to 230 3,176 853.40 2.65 30.00 1.56 36 1.13% -36.10%

Dilution envelope west of the fault 501 11,888 96.72 0.12 3.50 0.11 13 0.11% -4.03%

Dilution envelope east of the fault 502 5,694 19.46 0.17 4.00 0.16 9 0.16% -4.31%

Total 27,166 1,019.14 0.94 65 0.68 81 0.30% -10.42%

Dataset Block Code

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Figure 14.4 – Different graphs supporting a capping of 30 g/t Au for the mineralized zones east of the fault.

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Figure 14.5 – Different graphs supporting a capping of 30 g/t Au for the mineralized zones west of the fault.

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Figure 14.6 – Different graphs supporting a capping of 4 g/t Au for the dilution envelope east of the fault.

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Figure 14.7 – Different graphs supporting a capping of 3.5 g/t Au for the dilution envelope west of the fault.

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Compositing In order to minimize any bias introduced by the variable sample lengths, the capped gold assays of the DDH data were composited to equal lengths of 0.80 metres (“0.8m composites”) within all intervals that define each of the mineralized zones and dilution envelopes. When the last interval is less than 0.2 m, the composite is rejected. The total number of composites used in the DDH dataset is 83,736. A grade of 0.00 g/t Au was assigned to missing sample intervals. Table 14.2 summarizes the basic statistics for the gold composites.

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Table 14.2 – Summary statistics for the composites

Area Zone Block Code Number of Max Mean Standard Coefficient

Composites (Au g/t) (Au g/t) Deviation of Variation

101 101 544 19.87 0.92 1.82 1.98

102 102 649 57.72 1.11 3.54 3.18

103 103 703 64.43 1.17 3.71 3.16

104 104 434 57.94 1.14 3.99 3.49

105 105 808 65.00 1.40 4.46 3.18

106 106 761 51.22 1.38 3.92 2.85

107 107 746 64.96 1.66 4.28 2.57

108 108 701 64.19 1.10 3.92 3.58

109 109 607 30.10 0.54 2.12 3.93

110 110 578 51.43 0.65 3.16 4.86

111 111 517 24.92 0.45 1.91 4.27

112 112 75 9.87 0.34 1.24 3.60

113 113 75 14.42 0.92 2.82 3.08

114 114 85 4.11 0.26 0.69 2.70

115 115 89 11.06 0.93 1.96 2.11

116 116 91 15.44 0.58 2.06 3.58

117 117 99 60.50 1.55 6.68 4.32

118 118 107 4.76 0.34 0.79 2.33

119 119 120 11.49 0.64 1.78 2.80

120 120 431 44.04 0.65 3.42 5.25

121 121 258 47.37 0.45 3.15 7.07

123 123 368 31.67 1.21 2.54 2.09

124 124 299 10.51 1.06 1.73 1.63

125 125 247 64.96 1.37 4.67 3.42

126 126 175 24.24 1.35 2.86 2.11

127 127 118 30.07 0.47 3.19 6.81

128 128 98 28.38 0.40 2.97 7.38

201 201 64 1.78 0.16 0.33 2.02

202 202 93 8.06 0.46 1.40 3.06

203 203 89 18.81 0.81 2.61 3.24

204 204 165 17.73 0.70 2.12 3.04

205 205 201 8.78 0.56 1.10 1.97

206 206 221 6.17 0.57 0.95 1.66

207 207 530 30.00 1.25 2.46 1.96

208 208 578 18.79 1.08 1.94 1.79

209 209 433 30.00 1.14 3.24 2.84

210 210 457 29.99 0.92 2.54 2.75

211 211 427 20.13 1.10 2.09 1.90

212 212 297 12.25 0.83 1.55 1.87

213 213 181 19.59 1.39 2.52 1.81

214 214 56 15.29 0.88 2.22 2.52

215 215 15 2.92 0.47 0.87 1.84

216 216 71 15.16 1.24 3.04 2.45

219 219 63 3.45 0.42 0.75 1.76

220 220 30 9.76 0.53 1.80 3.44

221 221 25 7.03 1.23 2.14 1.74

222 222 499 30.00 1.01 2.67 2.63

223 223 241 30.00 0.93 3.58 3.84

224 224 158 30.00 1.23 2.96 2.41

225 225 58 10.99 0.83 1.64 1.97

226 226 136 18.17 0.86 2.70 3.12

227 227 59 13.55 0.59 1.87 3.19

228 228 67 6.38 0.33 0.91 2.75

229 229 19 1.44 0.16 0.37 2.33

230 230 43 9.26 1.01 2.03 2.00

Dilution envelope

west of the fault501 501 41,431 3.34 0.03 0.12 3.32

Dilution envelope

east of the fault502 502 27,184 4.38 0.04 0.13 3.27

Mineralized zones

west of the fault

Mineralized zones

east of the fault

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Bulk Density The drill hole database contains limited information regarding density information. Historical mineral resource estimates used a tonnage factor of 12.0 cubic feet per short ton (ft3/ton). The metric equivalent of a tonnage factor of 12.0 ft3/ton is the equivalent of a density factor of approximately 2.67 g/cm3. Although it is believed that this is slightly too low based on the mineralogy of the mineralization, the authors nonetheless used 2.67 g/cm3 for the current resource estimate. No sufficient physical specific gravity determination test work has been carried out to date to confirm this value. Table 14.3 (taken from the previous NI 43-101 report) illustrates how a density factor of approximately 2.75 g/cm3 might be more appropriate based on the typical mineralogy encountered in the deposit. Table 14.3 – Summary statistics for the composites

A density of 2.00 g/cm3 was assigned to the overburden, and 1.00 g/cm3 was assigned to the underground workings. Bulk densities were used to calculate tonnages from the volume estimates in the resource-grade block model.

Block Model A block model was established for the mineralized zones and dilution envelopes. The block model was extended to cover an area sufficient to host an open-pit if necessary. The model has been pushed down to a depth of approximately 1,700 m below surface. The block model was not rotated (Y-axis oriented along a N000 azimuth). The block dimensions reflect the sizes of the mineralized zones and plausible mining methods. Table 14.4 presents the properties of the block model. Table 14.4 – Block model properties

Mineral Specific Gravity (Dana, 1958) Relative abundance (%)

Quartz 2.65 - 2.66 87%

Biotite 2.80 - 3.20 5%

Calcite 2.72 5%

Arsenopyrite 5.90 - 6.20 2%

Pyrite 4.95 - 5.10 1%

Properties X (Columns) Y (Rows) Z (Levels)

Origin coordinates (UTM NAD83) 693500 5344700 500

Block size 3 3 3

Number of blocks 900 465 620

Block model extent (m) 2700 1395 1860

Rotation Not applied

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All blocks with more than 0.001% of their volume falling within a selected solid were assigned the corresponding solid block code in their respective folder. A percent block model was generated, reflecting the proportion of each block inside every solid (each individual mineralized zone, individual dilution envelope, overburden, country rock, underground workings). Precedence was respected during the process. Table 14.5 provides details about the naming convention for the corresponding GEMS solids, as well as the rock codes and block codes assigned to each individual solid. The multi-folder percent block model thus generated was used in the mineral resource estimation.

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Table 14.5 – Block model

NAME1 NAME2 NAME3

Mineralized Zone 101 ZoneClip 101 F150227 3























































Envelope_100 Dilution Envelope 501 ZoneClip 501 F150227 5

Envelope_200 Dilution Envelope 502 ZoneClip 502 F150227 5

Waste All remaining material - - - -

OB Overburden - - - 2

Voids Underground workings - - - 1

Zones_200

Precedence

Zones_100

Work-space DescriptionGEMS Triangulation Name

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Variography and Search Ellipsoids Three-dimensional directional variography was completed on DDH composites of the capped gold assay data for some of the mineralized zones. The study involved 10º incremental searches in the longitudinal plane, followed by 10º incremental searches in the vertical planes of the indicated preferred azimuths, as well as planes normal to the preferred azimuth. The study did not yield results that could be used in the context of the resource estimate presented herein due to strong heterogeneity, which likely reflects a strong nugget effect. The author defined ranges and orientations based on geological and historical development parameters for the project. The obtained ellipsoid for the mineralized zones and the dilution envelope west of the fault is oriented using 110 Principal Azimuth, -75 Principal Dip, and 0 Intermediate Azimuth (according to Gems’ Azimuth–Dip–Azimuth search anisotropy convention). The obtained ellipsoid for the mineralized zones and dilution envelope east of the fault is oriented using 110 Principal Azimuth, -65 Principal Dip, and 0 Intermediate Azimuth. The 3D ellipsoid corresponds to the strike and dip of the mineralized zones.

Grade Interpolation The ellipsoid shape summarized above provided the parameters to interpolate a grade model using the composites from the capped grade data to produce the best possible grade estimate for the defined resources. The interpolation was run on a point area workspace extracted from the DDH dataset. The composite points were assigned block codes corresponding to the mineralized zone or dilution envelope in which they occur. The interpolation profiles specify a single composite block code for each mineralized-zone solid, thus establishing hard boundaries between the mineralized zones and preventing block grades from being estimated using sample points with different block codes than the block being estimated. The interpolation profiles were customized to estimate grades separately for each of the mineralized zones and the dilution envelope. The inverse distance squared (ID2) method was selected for the final resource estimation. Two passes were defined in order to assess easily the effect of different ranges for mineralized zones while one pass was used for the dilution envelopes. Other than the variable ranges, no other parameters were modified from Pass 1 to Pass 2. The ellipsoid radiuses from pass 1 were established using a combination of reasonable assumptions, drill hole spacing, composite lengths, and the true thickness of the mineralized zones. The ellipsoid radiuses from pass 2 were fixed at values equivalent to 2x the ranges of the first pass to interpolate blocks that were not interpolated in the first pass. Pass 1, used for mineralized zones and dilution envelopes, is 50m X 25m X 12.5m. Pass 2, used only for mineralized zones, is 100m X 50m X 25m.

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Figure 14.8 – 3D view looking north-northeast showing Zone 101, all drill holes and the ellipsoid of Pass 1 (50m x 25m x 12.5m).

Figure 14.9 – 3D view looking north-northeast showing Zone 101, all drill holes and the ellipsoid of Pass 2 (100m x 50m x 25m).

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Resource Categories Mineral resource classification definition

The resource classification definitions used for this report are those published by the Canadian Institute of Mining, Metallurgy and Petroleum in their document “CIM Definition Standards for Mineral Resources and Reserves”. Measured Mineral Resource: that part of a Mineral Resource for which quantity, grade or quality, densities, shape, physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough to confirm both geological and grade continuity. Indicated Mineral Resource: that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough for geological and grade continuity to be reasonably assumed. Inferred Mineral Resource: that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes. Due to the uncertainty that may be attached to Inferred Mineral Resources, it cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resource as a result of continued exploration. Confidence in the estimate is insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure. Inferred Mineral Resources must be excluded from estimates forming the basis of feasibility or other economic studies.

Mineral resource classification All interpolated blocks were assigned to the Inferred category during the creation of the grade block model. The reclassification to an Indicated category was done for any blocks meeting all the conditions below:

Blocks interpolated from Pass 1. Blocks from mineralized zones only (not from the dilution envelopes). Blocks for which the distance to the closest composite is less than 20 m.

A series of outline rings (clipping boundaries) were created in long views using the criteria described above, but also keeping in mind that a significant cluster of blocks

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was necessary to obtain an Indicated resource. Within the Indicated category outlines, some Inferred blocks were upgraded to the Indicated category, whereas outside these outlines, some Indicated blocks have been downgraded to the Inferred category. InnovExplo is of the opinion that this was a necessary step to homogenize (smooth out) the resource volumes in each category, and to avoid keeping isolated blocks in the Indicated category. Figures 14.10 and 14.11 show the outlines used for the category classification for some of the main zones, while figures 14.12 and 14.13 show 3D views of the overall indicated resource above the cut-off grade of 3.50 g/t Au. In some areas, interpolated blocks were downgraded to not being assigned a category at all due to lack of confidence in grade and/or continuity. This mainly happens where drill spacing is significantly large or too close to the old O’Brien mine for which the compilation and validation work is incomplete.

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Figure 14.10 – Longitudinal view looking north showing all interpolated blocks of Zone 101 with respective categorization.

Indicated resource (default)

Indicated resource (upgraded)

Inferred category (default)

Inferred resource (downgraded)

Uncategorized material

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Figure 14.11 – Longitudinal view looking north showing all interpolated blocks of Zone 222 with respective categorization.

Indicated resource (default)

Indicated resource (upgraded)

Inferred category (default)

Inferred resource (downgraded)

Uncategorized material

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Figure 14.12 – 3D view looking northeast showing all indicated blocks above the cut-off grade of 3.50 g/t Au.

Figure 14.13 – 3D view looking northeast showing all indicated blocks above the cut-off grade of 3.50 g/t Au among drill holes and historical underground workings.

3.50 to 4.00

4.00 to 4.50

4.50 to 5.00

5.00 to 10.00

> 10.00

Au (g/t)

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Cut-off Grade A cut-off grade was established based on the parameters presented in Table 14.6. Table 14.6 – Input parameters used for the underground cut-off grade estimation

Figure 14.14 shows the variation of gold prices in American dollars, the CAD:USD exchange rate, and the resultant gold price in Canadian dollars. The dashed line presents the values used to determine the cut-off grade for the resource estimate presented in this report.

Input parameter Value

Gold price ($US/oz) 1,200.00

Exchange rate 1.00 USD : 1.20 CAD

Gold price ($C/oz) 1,440.00

Gold sel l ing costs ($C/oz) 5.00

Net gold price ($C/oz) 1,435.00

Mining costs ($C/t) 94.98

Mi l l ing costs ($C/t) 38.30

Total costs ($C/t) 133.28

Process ing recovery (%) 92.50

Mining di lution (%) 15.00

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Figure 14.14 – Graph showing variations of gold prices in $US, the CAD: USD exchange rate, and the resultant gold price in $C. The dashed line presents the values used to determine the cut-off grade for the resource estimate presented in this report (roughly averages of the previous six months). The parameters presented herein lead to a cut-off grade of 3.59 g/t Au. The underground resource estimate presented herein uses a value of 3.50 g/t Au for the underground cut-off grade in order to provide an adequate estimate based on current knowledge. The selected cut-off grade of 3.50 g/t Au allowed the mineral potential of the deposit to be outlined for an underground mining option. Although the open-pit option was briefly investigated, it was not retained.

Mineral Resource Estimate Given the density of the processed data, the search ellipse criteria, the drill hole density, and the specific interpolation parameters, InnovExplo is of the opinion that the current internal mineral resource estimate can be classified as Indicated and Inferred resources. The estimate is compliant with CIM standards and guidelines for reporting mineral resources and reserves. Table 14.7 displays the results of the In Situ Mineral Resource Estimate for the O’Brien Project (55 mineralized zones and 2 dilution envelopes) at the official 3.50 g/t Au cut-off grade (O’Brien and Kewagama claim blocks), as well as the sensitivity at other cut-off scenarios. The reader should be cautioned that the figures presented in Table 14.7, apart from the official scenario at 3.50 g/t Au, should not be misinterpreted as a mineral resource statement. The reported quantities and grade estimates at different cut-off grades are only presented to demonstrate the sensitivity of the resource model to the selection of a reporting cut-off grade.

1.10

1.15

1.20

1.25

1.30

1.35

1.40

1.45

1.50

1.55

1.60

$700

$800

$900

$1,000

$1,100

$1,200

$1,300

$1,400

$1,500

$1,600

$1,700

05/10/2014 04/11/2014 04/12/2014 03/01/2015 02/02/2015 04/03/2015 03/04/2015

Exch

an

ge R

ate

1U

S =

CA

D

Go

ld P

rice

Gold Price and Exchange Rate

Gold Price $USD Gold Price $CAD Exchange rate 1US=CAD

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Table 14.7 – O’Brien Project Mineral Resource Estimate at a 3.50 g/t Au cut-off (O’Brien and Kewagama claim blocks) and sensitivity at other cut-off scenarios

The Independent and Qualified Persons for the Mineral Resource Estimate, as defined by NI 43-101, are Pierre-Luc Richard, P.Geo., M.Sc. and

Alain Carrier. P.Geo., M.Sc., of InnovExplo Inc., and the effective date of the estimate is April 10, 2015. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. The resource model includes the previously named 36E Zone and Kewagama mine areas. The historical O’Brien mine area is not included in this

resource as it had not been compiled or validated at the time this estimate is being prepared. The model includes 55 gold-bearing zones, not all of which include resources at the official cut-off grade. A dilution envelope was also modelled, but no resource at the official cut-off grade is being reported for the envelope.

Results are presented in situ and undiluted. Sensitivity was assessed using cut-off grades of 2.00, 2.50, 3.00, 3.50, 4.00 and 5.00 g/t Au. The official resource is reported at a cut-off of 3.50 g/t Au.

The reader is cautioned that the figures presented herein, apart from the official scenario at 3.50 g/t Au, should not be misinterpreted as a mineral resource statement. The reported quantities and grade estimates at different cut-off grades are only presented to demonstrate the sensitivity of the resource model to the selection of a reporting cut-off grade.

Cut-off grades must be re-evaluated in light of prevailing market conditions (gold price, exchange rate and mining cost). A fixed density of 2.67g/cm3 was used for all zones. A minimum true thickness of 1.5 m was applied, using the grade of the adjacent material when assayed, or a value of zero when not assayed. High grade capping (Au) was done on raw assay data and established on a sector basis (Western zones: 65g/t, Eastern zones: 30g/t, Western

dilution zone: 3.5 g/t Eastern dilution zone: 4.0g/t). Compositing was done on drill hole intercepts falling within the mineralized zones (composite = 0.80 m). Resources were evaluated from drill holes using a 2-pass ID2 interpolation method in a block model (block size = 3 m x 3 m x 3 m). The inferred category is only defined within the areas where blocks were interpolated during pass 1 or pass 2. The indicated category is only defined in areas where the maximum distance to the closest drill hole composite is less than 20m for blocks interpolated

in pass 1. Ounce (troy) = metric tons x grade / 31.10348. Calculations used metric units (metres, tonnes and g/t). The number of metric tons was rounded to the nearest hundred. Any discrepancies in the totals are due to rounding effects. Rounding followed the

recommendations in NI 43-101. InnovExplo is not aware of any known environmental, permitting, legal, title-related, taxation, socio-political, marketing or other relevant issue that

could materially affect the Mineral Resource Estimate.

Zone Cut-off Tonnage Grade Ounces Zone Cut-off Tonnage Grade Ounces

2.00 1,384,700 4.22 188,049 2.00 3,388,500 3.64 396,601

2.50 991,200 5.01 159,770 2.50 2,254,100 4.36 315,725

3.00 748,800 5.75 138,456 3.00 1,525,300 5.12 251,293

3.50 570,800 6.53 119,819 3.50 918,300 6.38 188,466

4.00 444,300 7.33 104,676 4.00 663,500 7.42 158,273

5.00 320,800 8.43 86,939 5.00 486,200 8.52 133,245

Indicated

All

Zones

Inferred

All

Zones

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15. MINERAL RESERVE ESTIMATES

Mineral reserve estimates compliant with the reporting requirements of NI 43-101 have not been prepared for the O’Brien Project.

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16. MINING METHODS

Cautionary Statement The reader is cautioned that this Preliminary Economic Assessment (the “PEA”) is preliminary in nature. The PEA includes inferred mineral resources that are too speculative geologically to have economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the PEA will be realized.

Introduction The proposed mining plan for the O’Brien Project was prepared using the inferred and indicated resources estimated by InnovExplo. Due to the narrow vein nature of the orebody, two (2) underground mining methods were considered in the study, modified Avoca and long-hole mining with captive sublevels. The mining plan for the O’Brien Project comprises a combination of conventional and mechanized mining. The approach in this study has been to prioritize the modified Avoca mining method when possible. When this approach was not convenient, long-hole mining with captive sublevels was selected. The mineralized material will be transported to surface using a combination of 3.5-cubic-yard to 6-cubic-yard scoop trams and 30-tonne trucks. Waste material will be used to backfill mined out stopes as much as possible or will be brought to surface and stored on a dedicated waste pad. The deposit will be accessed via a ramp. The production drifts will be accessed via crosscuts connecting to the ramp. A portion of the resources will be mined using captive methods, however haulage will always be mechanized.

Mineral Resources Considered in the Mining Plan InnovExplo designed the conceptual underground preliminary mine plan based on the indicated and inferred resources presented in an earlier report entitled “NI 43-101 Technical Report for the O’Brien Project”, published on June 3, 2015 and prepared by InnovExplo Inc. Details of the available resources used to generate the preliminary mine plan are presented in Table 16.1. Table 16.1 - Resources considered in the mining plan (cut-off 3.5 g/t)

Category Tonnes (t) Grade (g/t Au)

Contained gold (oz)

Resource classification

Indicated 570,800 6.53 119,819 38% Inferred 918,300 6.38 188,466 62%

Potentially Mineable Mineral Resources The mineral resource block model prepared by InnovExplo was used for the PEA. First, the resources available for mining were defined by creating the stope geometry in the block model at a cut-off grade of 3.5 g/t. Then a second triage was done using

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a diluted cut-off grade of 4.01 g/t. The guideline used in the stope design was a minimum mining width of 1.8 metres for subvertical stopes. The subvertical structures were cut at 18-metre vertical intervals corresponding to access level elevations. The conversion of mineral resources to potential mineral reserves takes into account dilution and losses during mining operations. The mineral resources are already diluted to a minimum width of 1.8 metres. Mining recovery was established at 85%, to take into account pillar requirements. A 30% dilution was also taken into account for stope excavation. Finally, a 95% recovery was applied to account for mining operating losses. For stopes with a diluted grade of less than 5.0 g/t, an evaluation was made to determine the economic viability of each stope, considering the development required to access the stope. If the economic viability could not be justified, the stope was discarded. Following this exercise, that included mine dilution and mine recovery a total of 712,521 tonnes at 6.46 g/t (147,986 oz) was included in the mine plan.

Cut-off grade In order to establish which stopes could potentially be considered in the mine plan, the cut-off grade was evaluated. Each stope was evaluated individually to determine whether it would be included in the study or discarded. For the calculation of this cut-off grade, a gold price of US$1,180 per ounce and an exchange rate of 1.25 CAD/1 USD was used. The remaining parameters used in the cut-off grade estimation are presented in the following table. Table 16.2 – Cut-off grade parameters (CAD) Operating cost $173.28/t Mint cost $5.00/oz Mill recovery 91.50% Mining dilution 30% Cut-off grade 4.01 g/t

Geotechnical Evaluation No additional geotechnical study was conducted for the purposes of this PEA.

Typical ground support patterns These preliminary ground support recommendations are based on standard industry practice. More detailed recommendations will require additional information regarding joint spacing and continuity. Based on Farmer and Shelton (1983), the following proposed bolt lengths for the back are based on the excavation span (bolt length = 0.3 span; Table 16.3).

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Table 16.3 – Bolt length as a function of span Bolt length* Maximum span 5 ft (1.5 m) 16.5 ft (5 m) 8 ft (2.4 m) 26.2 ft (8 m)

* Bolt length indicates the length installed within the rock and excludes any threads or bar outside of the drill hole. The standard support consists of:

Back: rock bolts or rebar (length based on excavation span) on a 1.2 m × 1.2 m (4’ × 4’) pattern with screen as required (based on excavation height);

Wall: rows (number of rows based on excavation height) of rock bolts or friction bolts (length = 1.2-1.5 m) on a 1.2 m × 1.2 m (4’ × 4’) pattern.

Screening of the back to 1.2-1.5 m above the footwall is recommended for all excavations of 3.5 m height or more. The screen is intended as a safety measure where back height will make routine inspections and scaling more difficult. Once additional structural and rock quality information becomes available, it will be possible to optimize the ground support standards.

Mining Methods As mentioned earlier, two mining methods are proposed to accommodate the geometry of the mineralization: modified Avoca and long-hole mining with captive sublevels.

Modified Avoca mining method The modified Avoca mining method was mostly used in the present study. The proposed modified Avoca stope configuration is based on typical industry practice for currently operating mines in deposits with similar vein geometry.

The modified Avoca mining method consists of drilling a series of vertical holes downward into mineralization from one level to another. The mineralization is then blasted in vertical slices, and the broken material ends up in the bottom sill and is extracted using LHDs. For every sill and sublevel horizontal slice, a primary slot opened by drop raise method is excavated at each extremity of the level, and blasting of a first stope 18 to 22 metres in length along strike is achieved using a longitudinal retreating process. All the broken mineralized material is extracted before another slice is blasted to ensure maximum recovery of the mineralized material should any unplanned caving occur. Once the stope is completely mined out, waste rock is dumped in the empty stope as uncemented rock fill. To be able to blast the second stope of the same level, a void must first be created by pulling out some of the backfill of the first fully-backfilled stope. The second stope is then blasted, mucked and backfilled. The process is repeated until all sublevels are mined out. This mining method can also be referred to as longitudinal long-hole retreat mining method. Some parts of the mucking and backfilling steps are performed with remotely operated LHDs for safety reasons. Figures 16.1 illustrate the concept of the mining method.

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Figure 16.1 – Longitudinal view of the modified Avoca mining method: drilling, blasting and mucking activities.

Long-hole method with captive sublevels In some areas, the captive long-hole method was an economically better choice than the modified Avoca mining method. Long-hole stopes will be mined from 3-metre-high sublevels at ±15-metre vertical intervals. It is assumed that stopes will be backfilled. Pillars will be left in place between panels and mining horizons. It was assumed that pillars will have a minimum width of 3 metres or 1.5 times the width of the stopes. The method consists in drilling and blasting 63.5-mm-diameter holes in a pattern parallel to the walls. Holes are drilled upward or downward depending on the context. The development sequence consists in accessing the mineralized zone and excavating a level cut in the mineralized zone. The mining sequence will require the excavation of a raise opening, which is either developed as a conventional raise or as a drop raise when a top access is available. Once development is completed, the mineralized zone is surveyed with precision for the preparation of the drilling and blasting pattern.

16.6.2.1 Mining dilution and recoveries After exclusion of horizontal pillars, a mining recovery factor of 85% was applied in this study to account for the vertical pillars left in place. The average mining dilution factor was estimated at 30% (at 0.0 g/t Au) and the average development dilution factor was estimated at 50% (at 0.0 g/t Au). Then, a 95% mining recovery was applied to consider the general recovery of the mineralized material.

Kewagama shaft dewatering and shaft rehabilitation Prior to any underground rehabilitation or development work, the existing mine infrastructure will need to be dewatered. The total volume of water present in the mine is estimated at 13,139 m3.

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Initial dewatering is expected to be carried out at a rate of 250 USGPM, such that shaft rehabilitation work will take place over a maximum period of 33 days. Mine water will be pumped to the surface for treatment.

Underground mine design Primary development

The current PEA is based on an underground mine with access by decline to a vertical depth of 550 metres in the 36E area and 250 metres in the Kewagama area. The sublevels are developed at 18-metre vertical intervals. Each level or sublevel is accessed using 4.5 m × 4.5 m crosscuts from the main ramp or secondary ramp. The opening dimensions are planned as follows:

Ramp: 4.5 m × 4.5 m; Level: 4.5 m × 4.5 m; Sublevel drift: 3 m × 3 m; Captive sublevel: 3 m × 3 m.

The broken material will be hauled by LHDs from the production area to either a remuck bay or to loading points that will be excavated close to the ramp. The material will then be loaded in 30-tonne trucks to backfill the open stopes or hauled directly to surface. The Kewagama shaft will be rehabilitated to elevation 183 and will serve as the main emergency egress as well as for ventilation purposes. Short ventilation raises will be required as development progresses to accommodate the various production areas.

Secondary development Four permanent refuge stations are planned, one on elevation 201, a second one on elevation 93, a third on elevation -69 and a fourth on elevation -213. These refuge stations will be 4.5 m × 18 m × 4.5 m. One powder magazine will be constructed on elevation 183. Four sump stations are planned, one on elevation 255, a second one on elevation 201, a third on elevation 111 and a fourth on elevation 93. Portable sumps will be used during development.

Stope development In general, the modified Avoca and long-hole stope levels or sublevels will be excavated directly in the mineralized zones. In some cases, to allow for better mine sequencing, bypass drifts and draw points have been considered to preserve access to future resources. The long-hole stopes will be mined by retreating and the broken material will be collected in the lower level or sublevel. When the stopes are located at an elevation higher than the level, short raises will be developed and separated into two compartments by a timber wall, one side to serve as a manway and the other to be used as a chute.

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Stope ground support Stope ground support is used to control dilution. Dilution control can be achieved, to a certain extent, using long-range ground support. Cable bolts are used at the undercut and overcut. Cables of 3 metres and 5 metres are considered.

Mine Sequence Mine development will be accelerated in the first two years of the project to provide a degree of flexibility in terms of access, which should facilitate scheduling during the production period. The development sequence will ensure that many stopes are available for mining at a number of different locations at any given time. However, some of the stopes can only be mined at the end of the mine life since they are located directly over or under the level, therefore preventing any further access on that level when mined.

Mining Rate The expected average daily production rate during the production period is estimated in this PEA between 450 and 500 t/day. The overall project mine life is expected to be approximately 6 years, including a two-year pre-production period. In the opinion of author Laurent Roy, Eng., the mine plan should be achievable given the flexibility and number of available working places. Table 16.4 summarizes the annual tonnage distribution according to the mine plan. Table 16.4 – Mine plan tonnage distribution





Grade (g/t) 7.05 6.47 6.87 7.04 5.50 6.53 6.46

Mine plan schedule criteria Contractors will be used for all mine development, mine production and material haulage activities. A small staff will be hired to provide technical and administrative support and direction to the contractors. The design criteria used to develop the mine plan are as follows:

Overall mechanized drift development: o Single face: 150 metres/month; o Double face: 175 metres/month; o Multi-face: 200 metres/month.

Conventional equipment development: 100 metres/month. Alimak raise: 75 metres/month.

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Development and Production Schedule InnovExplo has prepared a preliminary development and production schedule based on the mineral resources discussed in Section 14. Development and production activities are based on a schedule of two 10-hour shifts per day, 7 days per week, 365 days per year. The underground mine design provides for a 6-year mine plan producing 712,521 tonnes of mineralized material grading 6.46 g/t. Using a mill recovery of 91.5%, this translates to 135,308 ounces of gold produced during this period. The mining plan includes all development required to access and mine the mineralized zones. Estimated development quantities are presented in Table 16.5 and the production schedule is presented in Table 16.6. Figure 16.2 gives a general overview of the total development and mineable zones. The resources included in the mining plan were obtained by applying the mining recovery and dilution factors presented in Section 16.5. Table 16.5 – O’Brien mine development quantities

Pre-production Production Total Year 1 Year 2 Year 3 Year 4 Year 5 Year 6

Development CAPEX Ramp and main level 4.5 × 4.5 (m) 1,053 2,968 3,145 2,357 766 - 10,289 Alimak raise 2.4 × 2.4 (m) 53 107 191 - 351 Development OPEX Sublevel 3.0 × 3.0 (m) 255 2,574 2,482 3,306 2,984 - 11,601 Conventional raise 2.4 × 2.4 (m) 114 18 - 132

Table 16.6 – O’Brien mine production rates

Pre-production Production Total Year 1 Year 2 Year 3 Year 4 Year 5 Year 6

Tonnage summary Mineralized material

(t) 3,196 66,668 158,574 169,891 186,934 127,259 712,522 Waste (t) 67,296 189,667 199,358 168,504 63,649 688,474

Total 70,492 256,335 357,932 338,395 250,583 127,259 1,400,996 Total mineralized material per day (t/day) N/A 183 434 465 512 465 Total mineralized material and waste per day (t/day) 230 702 981 927 687 465 Backfill (t) 37,946 122,692 146,560 86,746 53,076 447,020

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Figure 16.2 – O’Brien mine development and stopes

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Equipment Selection and Requirements A contractor will provide most of the equipment required for development and production. The new equipment that will be acquired by Radisson consists of three pickup trucks and two Kubota.

Manpower Requirements Radisson will hire its own staff for the project’s administrative, technical and surface services. Some positions will be partially staffed at the beginning of pre-production, progressively reaching the final fully-staffed scenario presented in Table 16.7. The Radisson mining staff will only work 9 months during the last year of production. Table 16.7 – Radisson mining staff

Manpower Pre-production Production Year 1 Year 2 Year 3 Year 4 Year 5 Year 6

Administration Manager 1 1 1 1 1 1 Secretary 1 1 1 1 1 1 Senior accountant 1 1 1 1 1 1 Intermediate accountant 1 1 1 1 0.75 Junior accountant 1 1 1 0.75 Senior purchaser 1 1 1 1 1 1 Clerk 1 2 2 2 1 Nurse 1 1 1 1 1 1

Surface services Dryman 1 2 2 2 1.5 1 Gate keeper 4 4 4 4 4 4

Technical services Geology

Chief geologist 1 1 1 1 1 Senior geologist 1 1 1 1 1 1 Junior geologist 1 1 1 1 0.5 Database technician 1 1 1 1 0.5 Senior geology technician 1 1 1 1 1 1

Engineering Chief engineer 1 1 1 1 1 Senior engineer 1 1 1 1 1 1 Junior engineer 1 1 1 1 0.5 Mining technician 1 1 1 1 0.5 Senior surveyor 2 2 2 2 2 2 Surveyor 1 1 1 0.5

The contractor will provide all manpower needed for supervision, maintenance, and production activities.

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The manpower needed on each working shift to achieve the mine schedule includes:

2 long-hole drillers; 2 long-hole blasters; 3 LHD and truck operators; 2-3 conventional development crews; 2 jumbo development crews:

o All crews with a two-boom jumbo for the ramp and early level development; o Each crew consists of 1 jumbo operator and 2 workers for ground support

and services.

Mining Services Ventilation

The existing Kewagama shaft will be rehabilitated and used for mine ventilation as well as an emergency escape way. Main ventilation fans and propane air heaters will be located near the main ventilation raises. For the current study, InnovExplo performed a preliminary simulation to estimate the ventilation equipment required. The simulation is based on the airflows required by the equipment used for development and production. The required ventilation was established at 105 cubic metres per second (220,000 cfm). Fresh air will be heated by propane burner systems and will exhaust via the ramp.

Dewatering The O’Brien mine has an estimated daily water inflow of 1,900 m3. Pumps have a capacity to handle 2,000 m3/day. During the spring thaw, pumps are working at full capacity. The pumping arrangement is a complex cascading system. Maximum head between lifts is 80 metres. The main electric pumps have power ratings between 15 and 50 hp.

Compressed air Three 28.3 m3/min (1,000 cfm), self-enclosed electric compressors will be installed at surface. A network of pipelines will be installed down the shaft and along the ramp and drifts throughout the mine. Compressed air will be provided to various handheld drills and production long-hole rigs, and will also provide emergency air supply to the refuge stations. A parallel network complete with pressure-reducing valves will supply water to the underground operations.

Underground power distribution One (1) 6-8 MVA 25 kV / 4.16 kV transformer will be installed near the portal. Underground, mounted on skids, a 4.16 kV / 600 V transformer and a 600 V distribution panel will be installed at each second level to provide power to the underground loads, such as pumps, fans, lunchroom, etc.

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17. RECOVERY METHODS

In view of potential mining activities, custom milling will be the preferred option. The recent metallurgical testwork has demonstrated the amenability of O’Brien mineralized material to the gravity, leaching and flotation processes. The O’Brien Project is planned for a five-year period at a production rate of 500 tpd. Five gold concentrators located within a 75-km radius were then identified as being able to potentially process the O’Brien material: the Kiena Mill, the Sigma-Lamaque Complex, the Camflo Mill, the Westwood Mill and the Aurbel Mill. Table 17.1 summarizes the main features of these milling options. Table 17.1 - Potential plants for custom milling




48 km Closed NA


Integra Gold

Gravity Concentration &

Leaching/CIP

1,200 to 2,400 tpd


Camflo Mill

Richmont Mines


800 to 1,200 tpd


Westwood Mill

Iamgold Gravity Concentration &

Leaching/CIP Or


Flotation

2,400 tpd

800 tpd

19 km Operating Yes

Aurbel Mill QMX Gravity Concentration &

Flotation & Flotation Concentrate

Leaching

500 to 800 tpd


conditions

The companies were contacted to find out their interest in performing custom milling. The Westwood and Aurbel mills have shown interest. This PEA is based on the use of the Westwood mill. The proximity of the O'Brien Project helps reduce transportation costs. In addition, the plant gives no restriction for environmental treatment. A trade-off study was conducted to compare treatment costs and potential recoveries for the two flowsheets available at the Westwood mill, see Table 17.2.

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Table 17.2 – Trade-off study Gravity/Flotation Gravity/Cyanidation

Gold value


Recovery % 94.52 91.53

Total3 C$/t 289 280

Milling cost



Smelting $/t 45 NA

Total $/t 81.78 36.78 1 Based on mining plan 2 URSTM test KN-F-3 3 Assumption section 17.1.2 BASED ON Gold PRICE at C$1475 /oz

The smelting cost was estimated based on information from similar projects. Preparation and trucking quotations were obtained from suppliers. The budgetary custom milling cost was estimated by the mill based on current knowledge of the ore. However, prices may be adjusted when additional information becomes available. Westwood's gravity and CIP circuit appears to be a good compromise based on the URSTM metallurgical results and the above considerations. The recovery will be lower but the treatment costs are significantly less. However, further work is required to better determine the specific flowsheet that will optimize the metallurgical performance.

Mineral processing description and recovery The ore is first dumped on a heavy-duty grizzly at the mine site. The oversize rock (> 18 inches) will be crushed by a contractor. The ore will be loaded and trucked to the Westwood mill. The plant will process ore at a rate of 2,400 tonnes per day for an entire month. During this period, the circuit will be dedicated to custom milling. Ore extracted at the Westwood mine will be stockpiled for later processing. The O’Brien Project will accumulate material during approximately 145 days before conducting a milling campaign. At this rate, there will be 2 to 3 milling campaigns per year. This section describes the proposed flowsheet at the Westwood mill and discusses the gold recovery that could be obtained in this processing facility considering the metallurgical testwork results obtained so far.

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Process description The crushing circuit is composed of a jaw crusher. The product is transported by conveyor belt to an ore bin for storage. The grinding circuit is composed of a SAG mill and a ball mill. The SAG mill and the ball mill run in closed circuit with their cyclones. The underflow product from the ball mill feeds the gravity recovery circuit. The gravity circuit is composed of a Knelson concentrator to recover free gold. The Knelson concentrate is treated on a shaking table. The gold concentrate is then further treated in the refinery. The cyclone overflow is sent to a trash screen. The underflow goes to the leach circuit after it has been thickened by a thickener. In the leach circuit, cyanide is used to dissolve the gold. Each leach tank is equipped with an agitator mechanism and oxygen lines. The discharge of the leach circuit flows to the carbon-in-pulp (CIP) circuit. The slurry goes from one tank to the other by gravity. Interstage screens prevent carbon from being carried away with the slurry. Carbon is pumped counter-current to the slurry. The combined volume of the leach and CIP circuits provides a residence time of approximately 72 hours. Loaded carbon is pumped from the CIP tank onto a screen, which returns the underflow slurry to the tank. Carbon is then sent to an acid wash column to eliminate carbonates. From there, the carbon goes into an elution column. The cooled pregnant solution is sent to the electrolysis cells located in the refinery. The drying oven and furnace are also used to treat the shaking table concentrate. The doré ingots are stored in a safety vault. Carbon from elution is regenerated in a rotary furnace, cooled, screened and returned to the last carbon-in-pulp tank. Fresh carbon is added as needed. After going through the CIP circuit, the slurry proceeds onto a safety screen to recover any smaller carbon particles that may have passed through the interstage screen. The underflow is sent to the cyanide destruction tank. The Westwood mill uses the SO2-air process. In the tank, reagents and air are used to reduce cyanide concentrations to environmentally acceptable levels. Once through cyanide destruction, the slurry is pumped to the tailings pond.

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Figure 17.1 – Typical gravity / CIP flowsheet

Expected recovery Custom milling with gravity and a leach retention time of 48 hours is considered. Based on URSTM testwork (Bouzahzah et al., 2014), the recoveries obtained are presented in Table 17.3. Table 17.3 – Recoveries obtained in laboratory

Gold Recovery in Laboratory Tests Gravity feed

size K80 Gravity

recovery Regrind K80 Leach - 48 h Total

µm % µm % %

102 58.0 NA 31.6 89.6

102 60.8 37 microns 32.1 92.9

The estimated recovery used for the PEA was calculated from all URSTM gravity tests done to 102 microns (8 tests). The average recovery is 59.7%. The cyanidation particle size at the Westwood facility must be around 70 microns. Since the tests were carried out at 37 and 102 microns, the potential recovery will be between these two values. Additional tests must be conducted to validate this assumption.

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The expected overall recovery for the O’Brien project was estimated at 91.5% (Table 17.4) based on URSTM laboratory testwork and gravity/leaching flowsheet. Certain historical laboratory tests obtained similar results. However, this value is higher than the average of available historical data. Additional testwork must be completed to validate the reproducibility. The amount of free gold recovered by gravity has a significant impact on the global recovery. As demonstrated by historical data, laboratory test work and past production, the quantity of free gold fluctuates depending on the ore zone. A representative sample of the entire ore body will be taken to confirm the free gold content. Table 17.4 – Expected gold recovery

Expected Gold Recovery Gravity feed

size K80 Gravity

recovery Regrind K80 Leach - 48 h Total

µm % µm % %

102 59.7 70 31.8 91.5

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18. PROJECT INFRASTRUCTURE

Surface Water Management Overburden, waste and ore pads

Some of the waste rock will be reused underground. The rest of the waste rock produced by the O’Brien 36E-Kewagama Project will be placed on a new proposed lined waste pad developed on the Kewagama site. The lined waste pad will also hold a mobile crusher as well as the ore pile. Surface runoff water that comes into contact with the ore, the waste rock or the overburden is considered “contact water”. A mean annual flow rate of approximately 6.9 m3/h was estimated for the “contact water” coming from the pads. According to waste rock and ore chemical characteristics described in chapter 20.3.1, runoff water from around the ore, overburden and waste pads will be collected by a network of ditches in order to reach the proposed accumulation pond for further treatment.

Mine dewatering water The underground mine’s dewatering water will be pumped to the surface for treatment. A design flow rate of 56.78 m3/h (250 usgm) was estimated by InnovExplo to dewater the Kewagama shaft and a flow rate of 11.36 m3/h (50 usgm) was estimated to dewater the ramp. The same water treatment system that will be used to treat the “contact water” from the ore, waste and overburden pads will be used to treat the mine dewatering water. Due to the presence of sulphide into the biotite alteration, arsenic trioxide strorage in neaby O'Brien mine and potential ''leachable'' minerals (see chapters 20.1.1 and 20.3.1), tight monitoring of the water quality will be required.

Water treatment plant There are plans to install a modular water treatment system using a physico-chemical process on site. The “contact water” resulting from the ore, waste and overburden pad runoff as well as the mine dewatering water will be directed into a new 5,700-m3 pond. The water treatment system included in the capital costs will be able to treat both flows. A 40 m × 40 m pad will be built to hold the geotubes. The existing ponds will be used for temporary sludge storage. The sludge will be stored permanently in the waste rock pile. Once treated, the water will transit through a 2,000-m3 polishing pond before exiting into the environment. The surface water management system will make sure the quality of water exiting the site into the receiving environment complies with federal and provincial regulations. A provision to fix the existing pond’s geomembrane is included in the capital costs.

Tailings Storage Facility The current O’Brien 36E-Kewagama Project intends to use the Westwood mine’s Tailing Storage Facility (TSF). Therefore, no TSF cells were designed and no costs were estimated for disposal and restoration of the project’s tailings.

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Access Road Main access road

The main access road and all planned infrastructure are shown on drawings 141-20460-00-00-00-0001 and 141-20460-00-00-00-0002 available in Appendix V. The main access road leading to the mine site already exists, turning right at the end of Petit Canada road. This road requires major repairs to allow ore haulage and secure personnel access. The road was initially built 10 metres wide, although it is no longer visible. Ditches and deforestation will be needed over a width of about 2 metres on both sides of the road. A four-inch-thick compacted MG-20 aggregate layer will be applied for the new wearing course. The main access road will be extended to the ore pad access for truck loading. The plan calls for all infrastructure to be erected on the north side of the main access road, thus allowing a bypass for the former road users. For all infrastructure construction, it is assumed that required aggregates MG-20b (0-¾ inch) and MG-112 (0-4 inches) will come from the pit located near the Kewagama site or the surrounding area included in the Project mining lease. For the quantities required, crushing will be executed on site and aggregate will be transported using off-road trucks. Budget quotations were requested from local contractors for this supply, preparation and installation. Access to the existing snowmobile trail will not be obstructed by the mine site. Crossing access over the main access road will be maintained as it is now.

Site access roads On the site, all roads will be designed for regular vehicle access to various locations and buildings, except for the road between the portal and waste and ore pads, which will be built wider and with a higher capacity. A truck scale (100-tonne capacity) will be installed along the main access road for ore transportation weight data gathering. The weight readings will be available on a console inside the gate house. A fence will be installed around the mine site to restrain unwanted access from surrounding trails. A remote controlled motorized gate will also allow site access control and tracking from the gate house.

Garage For the maintenance of mining machinery, a foldaway-type 12 m × 18 m garage is planned. This building will be installed on a concrete slab and equipped with heating, ventilation, lighting and electrical services. Exterior lighting and services will also be available. A 55 m × 25 m parking will be built between the garage and the portal.

Portal and Underground Mine Surface Equipment The required compressed air for the mining equipment will be supplied by three 1000-CFM compressors, one being standby. Compressors will be factory installed in

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containers equipped with air receivers, ventilation, heating, motor starter, control station, electrical lighting and services. Storage of mine equipment, like rods, will be possible in containers installed near the portal. The permanent ventilation equipment (heating system and propane tank) will be installed in Year 2 near the old Kewagama shaft on the existing pad. A new concrete slab will be built for the heating system. Propane tanks will be installed a little further and protected by bollards. This pad will only require a minor overhaul (deforestation and MG-20b wearing layer). The old Kewagama ventilation raise will be kept free of any new installation to allow future rehabilitation. A diesel fuel station, including storage (50,000-litre double-walled tank) and distribution system will be installed near the portal.

Explosives Storage Powder and explosives pads will be installed in the north part of the property, within the fenced perimeter. The building location is suitable for a maximum storage of 5,000 kg, based on the required clearance from mining activities. Both pads will be separated by a 3-metre-high berm. Explosives storage buildings are planned to be supplied in the material contract with the explosives vendor.

Administrative Building and Dry Complex The administrative building and dry complex are planned to be modular-type, installed on tripods and rented for the duration of the Project. These buildings will be installed on a 70 m × 100 m pad. All modules will be installed side-by-side and linked to one another, as shown in Appendix V. The dry will have a capacity of 110 baskets and bins and will be shipped in two modules. An area with 10 baskets and bins is reserved for female staff. The showers and restroom module will be installed between the basket and bin modules. The administrative area will include 26 offices (8 closed), a mine rescue room, an infirmary and a 22-place dining room, all included in 4 modules. The gate house will be an independent module (3 m × 6.1 m) equipped with an office and a restroom. The employees and visitors parking will accept 80 vehicles. Exterior lighting will be installed and service outlets will be available for vehicle heating. Because potable water and sewer systems from the town of Cadillac are not available on the O’Brien and Kewagama properties, independent systems are planned (water well and septic tank). The option to negotiate an agreement with the municipality to connect these services to the Kewagama site should be studied at later stages of advancement of the Project.

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Electrical Distribution In order to have only one electrical metering point, the main connection is planned at the existing O’Brien substation located at the old mill. A connection already exists to the 25-kV line nearby and the existing transformers would be kept to supply the existing O’Brien buildings. A new 25-kV overhead line will be installed along the main access road to feed the Kewagama area. The mandatory clearance from the 120-kV line located south of the main access road was considered in mine infrastructure design. The electrical substation at Kewagama is planned to be installed near the portal and protected by a 2.4-metre-tall fence. It will contain two step-down transformers, one 6-8 MVA 25 kV / 4.16 kV for underground distribution and one 2-3 MVA 25 kV / 600 V for other loads, as well as all the structural, insulators, disconnect switches, switchgear, grounding and other required hardware. Underground mine substations, mounted on skids and including a disconnect switch, a 4.16 kV / 600 V transformer and a 600 V distribution panel will be installed at each second level of the mine to supply mining equipment, fans, pumps, as well as lighting and other services. Also, a leaky feeder system is planned to allow communication throughout the underground mine and decline.

Existing Infrastructure in the O'Brien Area Existing infrastructure in the O'Brien mine area will be accessible by turning left at the end of Petit Canada road. A minor overhaul of this portion of the road is required. The only building that will be reused for the Project is the warehouse, where material requiring dry and tempered conditions can be stored. In particular, electrical spare parts should be stored in this building. The existing core shack will be maintained. Core storage will be possible on the concrete slab of the former administration building. All existing infrastructure in the O'Brien area are expected to be demolished during execution of the mine closure plan.

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19. MARKET STUDIES AND CONTRACTS

Market Studies No market studies were undertaken to support this PEA. The sole mineral considered for revenue within the PEA is gold doré. Markets for doré are readily available and the doré bars produced from Lamaque Project could be sold on the spot market. Gold markets are considered mature, despite a current gold price that is lower than the 3-year trailing average.

Metal Pricing Revenues were calculated using US$1180 and exchange rate of 1.25 $ CDN/$ US.

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20. ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

Previous Work on the Property The O’Brien Project consists of 21 mining claims covering an area of 637 hectares (Fig. 4.2). Two former mines are located on the property: the O'Brien mine is located in the western portion and the Kewagama mine is located approximately 1.5 km east of the O'Brien mine. Section 6.0 presents a detailed history of both mine sites. The next paragraphs will address potential environmental concerns.

Former O’Brien mine The O’Brien mine operated intermittently from 1926 to 1981 under different ownership. The mine is located about 1.5 km west of the 36E and Kewagama Project area. Today, the site is classified as an abandoned mine, and the MERN is responsible for the site even though it is located on Radisson’s mining claims. Surface infrastructure is still present, including:

Concrete foundations for the hoist and the administration building; Wooden structures for the electrical substation; Steel structures for the concentrator; Mechanical shop; Core shack; Concrete slabs/covers securing mine openings.

Between 1926 and 1956, O’Brien Gold Mines Ltd produced a total 587,120.8 ounces of gold from 1,197,147 metric tons milled at an average grade of 15.25 g/t Au (Table 6.2). During this period, a tailings impoundment area, north of the site, stored the tailings produced from the mill. The tailings impoundment is currently inactive and considered abandoned, and the MERN is responsible for the site. During the same period, the O’Brien mine also produced 6,313 metric tons of crude arsenic (arsenic trioxide) from the arsenic-bearing ore, of which 5,176 metric tons were sold prior to 1952. In 1956, with the authorization of the Québec Department of Mines, an estimated 1,150 metric tons of arsenic trioxide was stored in sealed barrels west of the No. 3 Shaft on the 1500' level (-455m agl) in the 15-G-West and 15-F-West drifts. The entrance ways to these storage drifts were sealed with concrete before flooding the mine. In the early 1970s, the mine was acquired by Darius Gold Mine Inc. (Darius) and reopened. From 1974 to 1981, Darius recovered 10,852.4 ounces of gold from 128,373 milled tons at an average grade of 2.63 g/t Au (Table 6.3). During this period, Darius built a second tailings impoundment area, adjacent to the former one. The second tailings impoundment is also currently inactive and considered abandoned, and the MERN is responsible for this area as well. In 1974, Darius believed it had a buyer for the crude arsenic stored on the 1500' level. Access was made through the concrete walls sealing the storage drifts,and Darius reported that the drifts were intact and dry. They reported 1,090 metric tons of arsenic trioxide stored in 45-gallon barrels. Unfortunately, the potential buyer withdrew.

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The mine was closed in 1981 and bought that same year by Sulpetro Minerals Ltd (“Sulpetro”) for the purpose of processing ore from its adjoining Kewagama mine to the east. Sulpetro tried unsuccessfully to find other buyers for the crude arsenic stored on the 1500' level, and in 1985, authorization was given by the Ministère de l’Environnement du Québec (MENVIQ) to confine and seal 1,253 tons of arsenic trioxide in 8,928 barrels in the 15-F-West and 15-G-West drifts (GEOSPEX, 1998). Following this, the mine was flooded.

Former Kewagama mine Activity on the site commenced in 1928, and in 1931, a shaft reaching a depth of 125’ gave access to a 1500’ development drift. This shaft was located 4,800’ east of the O’Brien mine shaft. In 1936, Kewagama Gold Mines Ltd was created. Over time, activities on the site were interrupted occasionally as many companies were involved (see section 6.2). The site was explored and 2,470 tons of ore was processed at the neighbouring Thompson Cadillac Mill. No waste rock piles or tailings impoundment areas were built on the site. In 1978, a temporary mining plant–service building, hoist room and headframe were built, along with a mine dry and a machine shop. In 2012, the site was restored by the MERN with the exception of two water retention ponds. All infrastructure at the Kewagama mine was removed. The Kewagama shaft was secured with a concrete slab.

Liability of Radisson regarding former mine sites Presently, the MERN has exempted Radisson of all liabilities associated with the historical tailings located on the site; however, if Radisson should decide to use the same area for tailings in the future, Radisson would acquire all of the liabilities for the past and present tailings.

Environmental Site Description and Characterization The area where future mining activities will take place has already been affected by previous mining activity. The area planned for development is adjacent to the previous (removed) infrastructure that was present on the Kewagama mine site, which was reclaimed in 2012 by the MERN. The 36E and Kewagama Project’s activities should be constrained to an area less than 15 hectares. An environmental baseline study is required to obtain permits for the 36E and Kewagama Project. The study will define the receiving environment before project development, including the physical, biological and social environment aspects. Since the site has already been affected by human activities, the baseline is also important to identify the potential for or presence of existing contamination. For this type of project, the study area should cover the location of the infrastructure within the project area, which represents 15 hectares. Field data will be collected to address the following subjects:

Hydrology and surface water quality Hydrogeology and groundwater quality Air quality

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Noise Soil Fish and fish habitat Flora Fauna Endangered species

Physical environment

The site is generally flat-lying, with very little topographic relief. The site is located in the Kinojévis River watershed and in the Preissac Lake sub-watershed. On the future mine site, the surface water flows towards a small creek in the northern portion of the site and continues into Blake River, which is a tributary of Preissac Lake. A wetland is present on the western portion of the site and is connected to the small creek in the north mentioned above. On the mine site, there are no streams or lakes. The overburden consists of glacial and fluvioglacial deposits as well as deposits from the proglacial Ojibway Lake. In the Cadillac area, glacial deposits consist primarily of till, observed along the top and sides of mounds and hills, as well as an esker located 6–8 km east of Cadillac. Glaciolacustrine deposits consist of clay and silt sediments in depressions, and reworked sand and gravel deposits at higher elevations. The groundwater could contain arsenic because gold mineralization in the Cadillac area often contains arsenopyrite (FeAsS). No information is yet available concerning the background concentrations of parameters in the groundwater, although this would be required when work starts.

Biological environment The site is located within the boreal forest zone, which covers much of northern Quebec. It is partially covered by black spruce, poplar and minor birch, tamarack and balsam fir trees. The animal species found in the vicinity are typical of the boreal forest and include moose, black bear, otter, marten and wolf. The site is not located in any government-designated protected zones for terrestrial plants or animals.

Management of Waste Rock, Tailings, Ore and Water Information on the environmental characterization of the ore, waste rock and tailings is available in the following reports provided in Appendix VI:

Report from Genivar, July 2012 : “Caractérisation physicochimique du minerai et des stériles à la propriété O’Brien, Cadillac”.

Report from the Unité de Recherche et de Services en Technologie Minérale (URSTM), October 2014 : “Report PU-2013-12-860 : Caractérisation minéralogique, métallurgique et environnementale d’échantillons de la zone 36 du gisement O’Brien”.

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Chemical characteristics of waste rock and ore Tests for acid rock drainage (ARD) potential and metal leaching potential were conducted on all sixteen (16) samples of ore and waste rock taken from a bulk sample: eight samples of ore and eight of waste rock. The potential for waste rock and ore material to generate acid rock drainage was evaluated through the Acid Base Accounting Method (ABA) (Genivar, 2012). Based on the ABA test results, the waste rock had an average total sulphur (S) content of 0.32% S, and the ore had a higher average content of 1.76% S. Although the ore samples had a net neutralization potential (NNP) greater than the specified limit of 20 kg CaCO3/tonne, the NP/AP ratio remains lower than 3 for five of the eight samples. Therefore, the ore is expected to generate acid (Genivar, 2012). Waste rock samples had a high net neutralization potential and a high NP/AP ratio for seven of the eight samples. Based on these results, the waste rock could be considered as non-potentially acid generating (non-PAG). Static tests required under the Québec’s Directive 019 to characterize the metal leaching potential of materials consist of trace metal analysis (MA.200 – Mét. 1.2) combined with short-term leaching tests following the Toxicity Characteristic Leaching Procedure (TCLP – EPA Method 1311: 1992). The TCLP tests use an organic acid (acetic acid) as the lixiviant, but it is not necessarily representative of the likely leaching conditions observed at the site. The Genivar report (2012) provides all the results of the waste rock and ore sample tests for trace metals. In the report, the metal concentrations are compared against the PPSRTC1 soil criteria A, B and C, and concentrations exceeding these criteria are highlighted. Criterion A represents the background values for the Superior Province (geological province) where the O’Brien Project is located. For the waste rock samples, the following elements showed concentrations higher than listed in Criterion A: As, Cd, Cr, Co, Cu, Mg, Mo and Ni. For ore samples, the following parameters showed higher concentrations than Criterion A: As, Cd, Cr, Cu, Mg, Mo and Ni. These exceedances of Criterion A require that the leachability of the waste rock and ore samples be assessed using the TCLP leach test method (as per Directive 019). All of the TCLP results for the 16 samples of ore and waste rock were compiled in the Genivar report (2012), which is provided in Appendix VI. All results were compared to the PPSRTC criteria for groundwater reporting to surface water and to Directive 019 criteria of high-risk mine waste. The following elements exceeded the criteria for leachate (TCLP) from ore samples: As, Cr, Cu, Ni and Pb. For waste rock samples, only Cr and Ni exceeded criteria levels. As per Directive 019, in order to be defined as “leachable”, a mine waste has to exceed levels for both solids (compared with Criteria A of the PPSRTC) and for the leachate composition (TCLP) (MENV, 2003). Using the Quebec Directive 019 criteria, 50% of the waste samples tested are classified as “leachable” for chromium (Cr), and most of the ore samples are considered “leachable” for arsenic (As), chromium (Cr) and nickel (Ni). There were no samples exceeding the high- risk mine waste criteria.

1 Politique de protection des sols et de réhabilitation des terrains contaminés (policy established by the MDDELCC)

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Tailings characteristics A sample of tailings from a flotation test was characterized during metallurgical testing performed by the URSTM in 2014. Based on the ABA testing results, the tailings had an average total sulphur concentration of 0.38%, and the neutralization potential is 94.3 kg CaCO3/tonne. Therefore, the tailings are expected to generate acid (URSTM, 2014). Leaching tests indicated the sample does not qualify as “high-risk” as per Directive 019, although static tests as per Directive 019 to characterize metal leaching potential have indicated that the tailings sample from the flotation test is leachable for Cu, Mn, Ni and Zn. Further tailings characterization is needed as only one sample was tested. The tailings are not expected to be acid generating based on the geology of the site, but they are expected to leach metals under neutral conditions. In the proposed project, the tailings will not be managed at the site. Ore will be processed elsewhere and tailings will be disposed of in an existing facility.

Run-off water management A water management plan will be developed to collect, monitor and treat, if required, all contact water from the mine site. Run-off water will be collected from the waste rock, ore and overburden piles. The run-off water will be collected in ditches before flowing to a sedimentation pond where it will be treated for suspended solids as well as metal content, if required. The overflow will then be discharged into a nearby stream. All other run-off water will follow natural watersheds around these zones and will not be allowed to enter the site’s drainage infrastructure.

Permitting Requirements The Cadillac Region is home to many active and historical mining operations. No EIA will be required for the 36E and Kewagama Project as the proposed output remains less than 2,000 tpd (EQA Q-2, r.23), and none of the physical activities (SOR/2012-147) would trigger the federal process. Mainly two provincial ministries will issue permits: the MERN and the MDDELCC. The following is a list of key permits that will be required. Mining Lease The mining lease is required to extract ore. It will be obtained from the MERN. Closure Plan According to the Mining Act, a closure plan is required in order to obtain a mining lease. The closure plan must be filed and approved by the MERN. The plan will present details on how the site will be reclaimed, and the entire cost will have to be deposited with the MERN within the first three years after obtaining the mining lease.

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Certificates of Authorization Certificates of authorization will have to be obtained from the MDDELCC for most of the activities planned on the site. These permits will contain details regarding design, environmental impacts and monitoring, etc. Other Requirements Other permits or leases will have to be obtained depending on planned development activities at the site. Also, depending on RCM2 or municipal legislation, some permits may also be required from the RCM or the municipality. Federal Government Based on available information, the federal government will not be involved in the permitting process. The 36E and Kewagama Project will not require any federal authorization.

Social or Community Impact The 36E and Kewagama Project is located in the municipality of Rouyn-Noranda in the Abitibi-Témiscamingue administration region. The municipality of Rouyn-Noranda is part of the Rouyn-Noranda RCM in the Cadillac District. The former mining sites are on public land, in the heart of the Abitibi Gold Belt, where municipal zoning allows resource development (mining or forestry). Other active mines and closed sites surround Radisson’s mine site area. The 36E and Kewagama Project is located approximately 1 km from the municipality of Cadillac. The mayor of Cadillac has been informed about the 36E and Kewagama Project and a committee of citizens is also being consulted and informed on a regular basis. An Algonquin community, Abitibiwinni Pikogan, lies approximately 45 km northeast of the 36E and Kewagama Project, near the town of Amos. The community has been informed about the 36E and Kewagama Project. Also, with its 73-km² surface area, Preissac Lake occupies a large portion of Preissac Township. The presence of outfitters, cottages and nautical clubs highlights the fact that Preissac Lake and the surrounding area is home to significant recreational and leisure activities. The Abitibiwinni Pikogan community and residents of the Preissac area are considered to be stakeholders; other stakeholders will be identified during the next development phases. The 36E and Kewagama Project area is accessed by a road that connects to Highway 117. The access road is owned by Radisson but is also used by land users (ATV, snowmobile, etc.). The 36E and Kewagama Project will make a positive economic contribution to the community as it will provide jobs for the local population and also generate economical opportunities for service suppliers. On December 16, 2015, the Quebec government published amendments to the Regulation regarding mineral substances other than petroleum, natural gas and brine. Section 52 of the Amended Act provides that the lessee of a mining lease must

2 Regional county municipality (MRC in French)

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establish a monitoring committee to foster the involvement of the local community within 30 days after the lease is issued. The committee must be maintained until all of the work described in the rehabilitation and closure plan related to the mining lease has been completed. 3

Mine Closure and Rehabilitation In accordance with Québec’s Mining Act, a closure plan must be approved by the MERN before releasing the mining lease. The concept for closure is to have acceptable conditions after restoration work is completed, ensuring that the environment will be protected and the security of stakeholders has been considered. The closure plan will address the rehabilitation of land and areas affected by mining activities (i.e., roads, pads, portals, buildings, water ponds, surface drainage patterns, etc.). The Mining Act has been updated recently, and additional measures were included to ensure that the restoration of mining sites upon closure is enforced. The total amount of rehabilitation costs required as a financial guarantee has been increased to 100%, and the payment schedule has been accelerated into three payments (50%, 25% and 25% of total costs over a 3-year period), with the first payment (half the cost) secured 90 days after the release of the mining lease. The closure plan must address the following items: securement of the mining area, dismantling of infrastructure, reclamation of waste rock and ore disposal areas, an emergency plan and post-closure environmental monitoring. The closure cost estimate for the 36E and Kewagama Project is based on capping the waste rock pile with an impermeable cover to limit water infiltration, which is in turn covered by a re-vegetated overburden layer. The overburden stockpile material will be used to reclaim and cover the waste rock pile. The cost of restoring the site is estimated to be C$3.6M. This cost estimate includes the cost of site restoration as well as post-closure monitoring.

3 http://www.miningprospectslawblog.com/2016/01/08/new-amendments-to-quebec-mining-regulations-come-into-force/#more-1547

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21. CAPITAL AND OPERATING COSTS

The PEA is based on capital pricing as of the third quarter of 2015. The PEA assumes that the development and mining of the mine will be done by contractors and that the latter will supply the mobile equipment.

Capital Costs The capitals costs were estimated using the following sources of information:

Quotes from equipment suppliers; Comparable installations at other mining projects; Contractor costs; and InnovExplo’s internal database.

The capital cost estimates are accurate within ±20%. The pre-production costs are estimated at $36.76 million, net of production revenue received during the second year of the pre-production period ($19.11 million). Pre-production capital costs are minimal given that there is no need to build processing and tailings facilities. Pre-production is anticipated to take 2 years with the majority of proceeds used for ramp construction and for sufficient development of mineralized zones, or working faces, to conduct mining at the proposed mining rate and mill throughput. Sustaining capital is estimated at $21.35 million, including $3.7 million for final closure costs and considering a salvage value of $1.46 million (Table 21.1). Table 21.1 – Capital cost estimate

Description Pre-production

Sustaining capital Total cost

Capitalized operating costs $21.33 M $21.33 M Capitalized revenue -$19.30 M -$19.30 M Royalty payment $1.00 M $1.00 M Development $20.01 M $17.13 M $37.14 M Mobile equipment $0.21 M $0.18 M $0.39 M Surface infrastructure $6.45 M $0.02 M $6.48 M Mine service infrastructure $7.29 M $0.78 M $8.06 M Closure costs $3.70 M $3.70 M Salvage value -$1.46 M -$1.46 M EPCM $0.77 M $0.77 M Total $36.76 M $21.35 M $58.12 M

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Capitalized operating costs Capitalized operating costs include all pre-production development and overall development carried out in Years 1 and 2 of the pre-production period. The capitalized operating costs include definition drilling, stope development, contractor indirect costs, mining costs (10% contingency), O’Brien staff, energy, milling and transportation, and environment (Table 21.2). Table 21.2 – Capitalized operating costs

Description Pre-production Definition drilling and sampling $0.27 M Stope development $4.68 M Contractor indirect costs $2.33 M Mining costs $2.34 M O'Brien staff and general $6.36 M Energy costs $1.96 M Milling and transportation $2.57 M Environment $0.82 M Total $21.33 M

Capitalized revenue During the 24-month pre-production period, it is anticipated that 13,345 ounces of gold will be produced, providing revenue of C$19.3 million (US$1180/oz and CAD/USD of 1.25). The pre-production revenue was capitalized.

Royalties As described in Section 4.5, the Kewagama property consisted of a contiguous block comprising three (3) mining claims covering an aggregate area of 112.07 hectares. Radisson owned a 100% interest in the Kewagama property, with a 2% NSR royalty payable to KWG Resources Inc. in the event of commercial production. In addition, a $1,000,000 payment must be made to Breakwater Resources Ltd (now Nyrstar) upon commencement of commercial production on either one of the O’Brien or Kewagama properties, against which shall be deducted any costs required to restore the O’Brien tailing ponds.

Development costs Development costs include all costs required by the contractor for mobilization and demobilization, portal construction, contractor indirect costs and development (Table 21.3). The contractor costs are based on quotations provided by contractors in the Abitibi region. Indirect costs include costs required by the contractor during the mine life such as indirect manpower and some equipment operating costs.

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Development costs include all costs required by the contractor to develop the ramp, main level drift and raises. It also includes the cost to build four (4) refuge stations, one (1) powder magazine and ventilation walls (SAS). Table 21.3 - Development costs


Indirect costs $7.27 M $3.94 M $11.21 M Development $12.74 M $13.19 M $25.93 M Total $20.01 M $17.13 M $37.14 M

Mobile equipment Most mobile equipment will be provided by the contractor and the cost is included in the development cost. Radisson will only provide three pickup trucks and two Kubota during the life of the mine. An estimated $0.39 million will be necessary for the mobile equipment. This includes a 10% contingency. The equipment cost is based on budgetary quotes obtained from equipment suppliers.

Surface infrastructure Surface infrastructure includes site preparation, buildings, and water management and distribution. Most costs were provided by WSP and a contingency between 15% and 30% was applied (Table 21.4). Table 21.4 – Surface infrastructure costs


Site preparation and installation $4.45 M $4.45 M Buildings $0.56 M $0.02 M $0.58 M Water management $1.44 M $1.44 M

Total $6.45 M $0.02 M $6.48 M

21.1.6.1 Site preparation and installation The site preparation and installation costs were mostly estimated by WSP. These costs include the material and equipment and their installation. A contingency between 15% and 30% was applied. Based on similar projects and quotations, InnovExplo estimated the costs of propane installation, powder and cap magazine installation and safety exits. A 20% contingency was applied to these costs. The total cost is estimated at $4,453,238. Table 21.5 presents the cost breakdown.

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Table 21.5 – Site preparation and installation

Description Total cost ($)

Overburden pad 79,834 Waste pad 1,520,573 Powder magazine 36,361 Cap magazine 32,631 Fuel station (earthwork and concrete) 353,492 Garage (earthwork and concrete) 157,392 Electrical substation (concrete and mechanical) 279,444 Storage containers 5,344 Truck scale 258,899 Compressors (earthwork and concrete) 6,367 Loading station 5,189 Ventilation heating system (earthwork and concrete) 52,902 Propane vessel (earthwork and concrete) 4,666 Ramp portal 45,360 On-site roads 350,733 Office (earthwork) 118,890 Septic system 19,710 Site fencing 286,474 Employee parking 87,825 Potable water well 25,944 Water treatment pond 330,000 Final water treatment system (earthwork) 70,784 Water treatment polishing pond (earthwork) 147,724 Propane installation 18,000 Powder and cap magazine installation 36,000 Safety exit 122,700

Total 4,453,238

21.1.6.2 Buildings Building costs were estimated by WSP and include the installation, transportation and demobilization of the rental buildings in addition to the office furniture. A contingency between 15% and 30% was applied. Table 21.6 presents the cost breakdown.

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Table 21.6 – Buildings

Description Total ($)

Warehouse 7,020 Processing plant 32,760 Garage 167,946 Storage containers 25,208 Office 296,412 Gate 13,000 Powder and cap magazine 34,884 Total 577,230

21.1.6.3 Water management and distribution Water management and distribution costs were estimated by WSP and include a contingency between 15% and 30%. The total cost is estimated at $1,444,902 and the cost breakdown is presented in Table 21.7. Table 21.7 – Water management and distribution

Description Total ($)

Environmental study 500,000 Septic system (mechanical and piping) 129,248 Gate (piping) 6,412 Potable water well (mechanical and piping) 164,233 Water treatment pond (piping) 10,100 Water treatment system (mechanical and environment) 593,949 Water treatment polishing pond (piping) 4,040 Potable water treatment (mechanical) 36,920 Total 1,444,902

Mine service infrastructure The mine service infrastructure cost includes mine dewatering, compressed air distribution, ventilation and air heating, electrical distribution and communication systems (Table 21.8). Mine dewatering costs include the cost of pumping water from the old mine, the development pumping system and the main pumping station. Compressed air distribution was estimated by WSP and includes equipment and installation.

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Ventilation and air heating costs include ventilation fans, surface and underground ventilation set up, rigid conduits and air heating system. The total cost of $1.67 million includes a contingency of 20%. The cost of the electrical distribution & communication systems was estimated by WSP for surface and underground and includes the power distribution, cables and connectors, instruments and communication, lighting and accessories. Table 21.8 – Mine service infrastructure costs

Description Pre-production Production Total

Mine dewatering $1.05 M $0.52 M $1.56 M Compressed air distribution $0.77 M $0.77 M Ventilation and air heating $1.41 M $0.26 M $1.67 M Electrical distribution and comm. systems $4.06 M $4.06 M Total $7.29 M $0.78 M $8.06 M

EPCM cost The engineering, procurement and construction management cost is estimated at $772,200, using a ratio of the direct costs for each discipline (4-6% for engineering, same for procurement and construction management). Also, in order to reduce consultant manpower costs, it was considered that 75% of procurement and construction management would be done by site crew already mobilized on the project.

Closure Costs

Project closure costs for the O’Brien site have been evaluated at $3.7 million. The closure cost includes the dismantling of buildings and the general rehabilitation of the O’Brien mine site.

Salvage value The salvage value was estimated for some of the infrastructure, electrical installations and mobile equipment on a case-by-case basis. For the mobile equipment, it was limited to 25% to 35% depending on the number of years of use.

Operating Costs

Operating costs are estimated in 2015 Canadian dollars with no allowance for escalation. The total operating cost and average unit operating costs are summarized in Table 21.9. The overall unit operating cost is $177.10 per tonne.

Operating costs are summarized below for the production period (Table 21.9).

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Table 21.9 – Summary of operating costs


Definition drilling and sampling $2.47 M 3.85 20.29 Stope development $22.09 M 34.38 181.16 Contractor indirect costs $18.06 M 28.11 148.11 Mining costs $27.30 M 42.48 223.84 O'Brien staff and general $12.38 M 19.27 101.53 Energy costs $5.89 M 9.17 48.32 Milling and transportation $23.64 M 36.78 193.80 Environment $1.97 M 3.06 16.14 Total $113.81 M 177.10 933.18

Definition drilling InnovExplo has estimated the cost of definition drilling at $3.85/t including the cost for sampling. This estimate is based on similar mine operating practices. According to the LOM conceptual mining plan, access for setting up the drill will generally be straightforward. The resulting total estimate for definition drilling is $2.47 million.

Stope development The unit cost for stope preparation stands at $34.38 per tonne milled (based on milled tonnage assigned to production). This cost is based on quotations from contractors. The development costs in the quotation include material (explosives, ground support, installed piping and equipment) and manpower.

Contractor indirect costs The portion of contractor indirect costs attributed to operating costs is estimated at $18.06 million or $28.11/t.

Mining costs Mining costs include stoping, mobile equipment and surface manpower. A 10% contingency has been applied (Table 21.10). Stoping costs include material and manpower for long-hole mining, material and maintenance for haulage and backfill. It also includes the cost of drop raises and cable bolting. The cost for material handling is estimated at $14.99/t, including material, maintenance and manpower. Long-hole stoping costs amount to $24.88/t. The mobile equipment operating cost only includes the cost for equipment provided by Radisson, namely pickup trucks and underground trucks. Manpower only includes the minimal surface crew (dryman and gate keeper) and the material needed. All these costs include 10% contingency.

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Table 21.10 – Mining costs

Description Total cost Unit cost

($/t) ($/oz)

Stoping $25.62 M 39.87 210.07 Mobile equipment $0.43 M 0.67 3.53 Surface manpower and material $1.25 M 1.94 10.24 Total $27.30 M 42.48 223.84

O’Brien staff and general costs Staff and employee salaries and associated expenses include those for administration and technical services. The salaries and the departmental general operating costs are based on costs at other mining operations. To account for benefits, a 33% premium was added and depending on the position, bonuses of 5% to 22% were also included. The estimate of departmental general operating costs was based on comparable mine operating budgets. The average cost for O’Brien staff and general costs is $19.27 per tonne milled. The annual cost during production is $3.58 million, averaging $19.27 per tonne milled. The total cost for O’Brien staff and departmental costs is estimated at $12.38 million. The annual cost is detailed in Table 21.11. Table 21.11 – O’Brien staff salaries

Manpower Annual cost

($) Administration

Manager 271,250 Secretary 82,800 Senior accountant 157,300 Intermediate accountant 117,300 Senior purchaser 128,700 Clerk 138,000 Nurse 103,500 Building rental 121,080 Material and others 877,257

Subtotal 1,997,187 Technical services

Chief geologist 171,600 Senior geologist 128,700 Junior geologist 82,800 Database technician 89,700 Senior geology technician 103,500

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Manpower Annual cost

($) Chief engineer 171,600

Technical services Senior engineer 128,700 Junior engineer 89,700 Mining technician 103,500 Senior surveyor 207,000 Surveyor 69,000 Material and others 240,720

Subtotal 1,586,520 Total 3,583,707

Energy The energy cost includes all electrical consumption, the propane needed to heat the underground air, and the rental of a propane tank. The diesel cost for underground and surface equipment is already included in unit costs (development, mining, and transportation) or departmental general operating costs. The estimated average annual electrical consumption was estimated by WSP. For the production period, an estimated 52,188,930 kWh will be necessary, representing an average annual cost of $1,124,069 (Table 21.12). The electrical consumption cost is based on Hydro-Québec's M rate. The estimated annual propane consumption for Years 3-6 is 1,059,842 litres per year, amounting to $635,905 per year at a price of $0.60/litre (budget quotation from Propane Nord-Ouest). The propane tank rental cost is $3,900/year. As shown in Table 21.12, the estimated total annual energy cost is $1.8 million, representing an average of $9.17 per tonne milled for Years 3-6. Table 21.12 – Annual energy cost (average for Years 3-6)

Description Annual cost ($)

Electricity 1,124,069 Propane 635,905 Propane tank rental 3,900 Total 1,763,874

Milling and transportation Mineralized material from the O’Brien Project will be processed at a mill in the Abitibi area with excess capacity for the duration of the O’Brien mine operation. Potential custom milling partners have been contacted and tentative commitments have been arranged for the processing of mineralized material. For the study, it is assumed that the mineralized material will be trucked to a custom mill located approximately 20 km from the O’Brien Project.

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The Company was able to identify which mills are best suited for material from the O’Brien Project, and this information was taken into account in the determination of the $36.78 per tonne assumption for milling and transportation costs. The unit cost for milling is estimated at $33.10/t, and at $3.68/t for truck loading and transportation.

Environment The environmental operating cost is based on similar operations. The environmental cost covers required manpower, analyses and environmental monitoring of effluent water and underground water quality based on current regulations. Water treatment costs were evaluated based on the projected pumping rate and a unit cost of $0.50/m3. The cost for management and disposal of waste and hazardous material is included. The average environmental cost is estimated at $3.06 per tonne milled. The estimated annual environmental cost is $571,600 (Table 21.13). Table 21.13 – Annual environmental cost

Description Annual cost ($)

Manpower 171,600 Analyses and environmental monitoring 50,000 Sewage system maintenance and monitoring 350,000

Total 571,600

Capitalized operating costs The operating costs incurred during the pre-production period ($21,329,960) were capitalized.

Taxes and royalties The O’Brien Project is subject to the following taxes:

Québec mining rights; Federal and provincial taxes.

An NSR royalty of 2% was considered for the tonnage from the Kewagama property. In addition, a $1,000,000 payment must be made to Breakwater Resources Ltd (now Nyrstar) upon commencement of commercial production on either one of the O’Brien or Kewagama properties, against which shall be deducted any costs required to restore the O’Brien tailing ponds.

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22. ECONOMIC ANALYSIS

Financial Analysis An after-tax model was developed for the O’Brien Project. All costs are in 2015 Canadian dollars with no allowance for inflation or escalation. The O’Brien Project is subject to federal and provincial taxes and taxes relating to Québec mining rights. Income taxes are calculated in accordance with the federal and provincial tax legislations relating to mining companies. The calculations were made by Lucie Chouinard of Raymond Chabot. The federal income tax rate is 15% and the combined provincial income tax rate is 11.9%. Québec mining duties are calculated in accordance with Bill 55, which contains amendments to Québec’s Mining Tax Act and received its first reading in the Québec legislature on November 12, 2013. Under the new regime in the Mining Tax Act, mining operators in Québec are required to pay the higher of a new minimum mining tax applied to the value of the ore at the mine shaft head and a progressive tax on excess profits. The new mining tax introduces progressive mining tax rates ranging from 16% to 28% (replacing the single tax rate of 16%), and a minimum mining tax based on the mine-mouth output value is used. The effective rate of this tax on mining profits starts at the existing 16% rate for mining companies with a profit margin of 35% or less, but rises to 17.8% for mining companies with a profit margin from 35% to 50%, and reaches as high as 22.9% for mining companies with a profit margin of more than 50%. The profit margin is calculated on the operator’s mining profit divided by the total of the gross value of annual output for all the mines it operates. Therefore, the higher a mining corporation’s profit margin, the higher the mining tax. Radisson owns a 100% interest in the Kewagama property, with a 2% NSR royalty payable to KWG Resources Inc. in the event of commercial production. In the cash flow analysis, this royalty was considered on all ounces produced from the Kewagama property. The economic evaluation was performed using the Internal Rate of Return (IRR) and the Net Present Value (NPV) methods. The IRR on an investment is defined as the rate of interest earned on the unrecovered balance of an investment. The discount rate makes the NPV of all cash flows equal to zero. The NPV method converts all cash flows for investments and revenues occurring throughout the planning horizon of a project to an equivalent single sum at present time at a specific discount rate. The discount rate used in the analysis is 5%. According to the NPV method, a positive NPV represents a profitable investment where the initial investment plus any financing interest are recovered. This Preliminary Economic Assessment (PEA) is preliminary in nature as it includes Inferred Mineral Resources that are too speculative geologically to have economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the PEA will be realized.

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The following parameters were considered in the financial analysis (Table 22.1):



produced from the Kewagama property. A residual fiscal base of $ 5.8M was considered in the tax estimation regarding

previous expenses by Radisson on the O’Brien Project.Resources as presented in Section 14.

Future annual cash flow estimates based on grade, gold recoveries and cost estimates as previously discussed in this Report.

69,864 tonnes of mineralized material to be processed during the pre-production period, deemed as capital production and not included in production nor the revenue derived from it.

The main parameters and cash flow analysis results for the entire project are presented in Table 22.1. Details of the cash flow analysis are presented in Table 22.2. Table 22.1 – Cash flow analysis summary












PRE-TAX



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Parameters Results


AFTER-TAX


IRR (%) 3.15%


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Table 22.2 – Economic analysis for the O’Brien Project (figures in Canadian dollars) Radisson - O'Brien ProjectCashflow summary

TotalPRODUCTION

Development (t) 3 196 33 474 32 080 40 298 52 409 - 161 458 Grade (g/t) 7,05 5,74 6,19 5,95 5,11 - 5,70

Long Hole (t) - 33 194 126 494 129 593 134 524 127 259 551 064 Grade (g/t) - 7,20 7,05 7,39 5,66 6,53 6,68

Total (tonne milled) 3 196 66 668 158 574 169 891 186 934 127 259 712 521 Grade (g/t) 7,05 6,47 6,87 7,04 5,50 6,53 6,46 Recovery (%) 91,5% 91,5% 91,5% 91,5% 91,5% 91,5% 91,5%

Gold Produced (oz) all zones 663 12 682 32 057 35 206 30 261 24 439 135 308 Gold Produced (oz) Kewagama 663 11 026 12 808 - - - 24 497 Gold Produced (oz) Royalty 2% Kewagama 13 221 256 - - - 490 Tonne milled assigned to capital 3 196 66 668 69 864 Gold Produced assigned to capital (oz) 663 12 682 13 345 Tonne milled assigned to production 158 574 169 891 186 934 127 259 642 658 Grade (g/t) 6,87 7,04 5,50 6,53 6,45 Gold Produced assigned to production(oz) 32 057 35 206 30 261 24 439 121 963

REVENUESGold Price ($US/oz) 1 180 $ 1 180 $ 1 180 $ 1 180 $ 1 180 $ 1 180 $ 1 180 $Exchange rate ($CAN/$US) 1,25 $ 1,25 $ 1,25 $ 1,25 $ 1,25 $ 1,25 $ 1,25 $Gold Price ($CAN/oz) 1 475 $ 1 475 $ 1 475 $ 1 475 $ 1 475 $ 1 475 $ 1 475 $Gross Revenue 977 850 $ 18 706 105 $ 47 284 419 $ 51 928 456 $ 44 635 278 $ 36 047 344 $ 199 579 452 $Mint (cost 3,00$ per oz) 1 989 $ 38 046 $ 96 172 $ 105 617 $ 90 784 $ 73 317 $ 405 924 $Kewagama royalty (NSR 2%) 19 556,99 $ 325 260 $ 377 849 $ 722 666 $Capitalized revenue 956 304 $ 18 342 799 $ 19 299 103 $Net Revenue 46 810 398 $ 51 822 839 $ 44 544 495 $ 35 974 027 $ 179 151 759 $OPERATING EXPENDITURESDefinition drilling and sampling 12 305 $ 256 670 $ 610 509 $ 654 080 $ 719 695 $ 489 946 $ 2 743 206 $Stope development 413 926 $ 4 267 971 $ 4 420 151 $ 9 271 152 $ 8 403 038 $ 0 $ 26 776 237 $Contractor indirect cost 339 466 $ 1 992 621 $ 3 244 190 $ 4 871 956 $ 5 787 907 $ 4 160 058 $ 20 396 199 $Mining cost 299 255 $ 2 038 487 $ 6 240 115 $ 6 734 672 $ 7 510 252 $ 6 814 711 $ 29 637 492 $O'Brien staff and General 2 808 258 $ 3 553 707 $ 3 583 707 $ 3 583 707 $ 3 173 560 $ 2 041 831 $ 18 744 771 $Energy cost 567 892 $ 1 392 360 $ 1 615 726 $ 1 786 938 $ 1 912 580 $ 577 822 $ 7 853 320 $Milling and transportation 117 549 $ 2 452 035 $ 5 832 348 $ 6 248 590 $ 6 875 427 $ 4 680 579 $ 26 206 527 $Environment 345 860 $ 471 600 $ 571 600 $ 571 600 $ 571 600 $ 253 700 $ 2 785 960 $Capitalized operating cost 4 904 510 $ 16 425 451 $ 21 329 961 $Total operating costs 0 $ 0 $ 26 118 347 $ 33 722 697 $ 34 954 059 $ 19 018 648 $ 113 813 751 $Cash op. cost/tonne $CAN 164,71 $ 198,50 $ 186,99 $ 149,45 $ 177,10 $Cash op. cost/oz $CAN 814,74 $ 957,88 $ 1 155,08 $ 778,21 $ 933,18 $Total cash op. cost/tonne $CAN 167,70 $ 199,12 $ 187,47 $ 150,02 $ 178,26 $Total cash op. cost/oz $CAN 829,53 $ 960,88 $ 1 158,08 $ 781,21 $ 939,28 $Cash op. cost/tonne $US 131,77 $ 158,80 $ 149,59 $ 119,56 $ 141,68 $Cash op. cost/oz $US 651,79 $ 766,30 $ 924,06 $ 622,57 $ 746,55 $Total cash op. cost/tonne $US 134,16 $ 159,29 $ 149,98 $ 120,02 $ 142,60 $Total cash op. cost/oz $US 663,62 $ 768,70 $ 926,46 $ 624,97 $ 751,42 $Operating Cash Flow 0 $ 0 $ 20 692 051 $ 18 100 143 $ 9 590 436 $ 16 955 379 $ 65 338 008 $CAPITAL EXPENDITURESCapitalized operating cost 4 904 510 $ 16 425 451 $ 21 329 961 $Capitalized revenue 956 304 $ 18 342 799 $ 19 299 103 $Royalty payment 1 000 000 $ 1 000 000 $Preproduction contractor 791 912 $ 0 $ 0 $ 0 $ 0 $ 110 000 $ 901 912 $Contractor indirect 3 715 083 $ 3 554 124 $ 2 784 880 $ 1 157 114 $ 0 $ 0 $ 11 211 200 $Development 2 949 809 $ 9 001 010 $ 9 464 855 $ 3 611 313 $ 0 $ 0 $ 25 026 987 $Mobile Equipment 147 847 $ 60 672 $ 60 672 $ 60 672 $ 60 672 $ 0 $ 390 536 $Site preparation and installation 4 453 238 $ 0 $ 0 $ 0 $ 0 $ 0 $ 4 453 238 $Buildings 547 854 $ 7 344 $ 7 344 $ 7 344 $ 7 344 $ 0 $ 577 230 $Water management and distribution - Environment 1 444 902 $ 0 $ 0 $ 0 $ 0 $ 0 $ 1 444 902 $Ventilation and Air heating 480 751 $ 932 762 $ 232 625 $ 28 564 $ 0 $ 0 $ 1 674 702 $Electrical distribution & Communication system 4 057 141 $ 0 $ 0 $ 0 $ 0 $ 0 $ 4 057 141 $Mine dewatering 532 621 $ 516 000 $ 480 000 $ 36 000 $ 0 $ 0 $ 1 564 621 $Compressed Air distribution 768 000 $ 0 $ 0 $ 0 $ 0 $ 0 $ 768 000 $EPCM 386 100 $ 386 100 $ 0 $ 0 $ 0 $ 0 $ 772 200 $Total capital expenditures 24 223 465 $ 12 540 664 $ 14 030 376 $ 4 901 006 $ 68 016 $ 110 000 $ 55 873 527 $All-in sustaining cost/oz $CAN 1 095,96 $All-in sustaining cost/oz $US 876,77 $All-in cost/oz $CAN 1 447,43 $All-in cost/oz $US 1 157,94 $Salvage 1 460 908 $ 1 460 908 $Financial guarantee reimbursement 1 851 634 $ 925 817 $ 925 817 $ 3 703 268 $ 0 $Closure Costs 3 703 268 $ 3 703 268 $Net cashflow 24 223 465 $ 12 540 664 $ 4 810 041 $ 12 273 320 $ 8 596 603 $ 18 306 286 $ 7 222 121 $Cumulative cashflow 24 223 465 $ 36 764 129 $ 31 954 088 $ 19 680 768 $ 11 084 165 $ 7 222 121 $Estimated Mining and Income taxes 2 801 055 $ 1 721 990 $ 2 507 859 $ 2 184 779 $ 1 072 778 $ 2 060 290 $ 3 302 660 $Cash Surplus After Taxes 21 422 410 $ 10 818 674 $ 2 302 182 $ 10 088 541 $ 7 523 825 $ 16 245 997 $ 3 919 461 $Cumulative Cash flow After Taxes 21 422 410 $ 32 241 084 $ 29 938 902 $ 19 850 361 $ 12 326 536 $ 3 919 461 $

Pre-tax NPV (5%) 203 762 $ (0,8) Pre-tax IRR 5,18%

After-tax NPV (5%) 1 908 446 $After-tax IRR 3,15%

ALL ZONES

ProductionPre production

Year 4 Year 5 Year 6Year 1 Year 2 Year 3

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Sensitivity Analysis The parameters in the sensitivity analysis were chosen based on their potential impact on the outcome of the economic evaluation. Key economics were examined by running cash flow sensitivities against:

Operating cost (OPEX); Capital cost (CAPEX); Revenue; Gold price, exchange rate, grade and mill recovery.

Sensitivity analyses were performed on the Project’s after-tax NPV (5%) and IRR by applying a range of variation revenue (±30%) to the parameter values. Results are presented in tables 22.3 to 22.6. The effects on NPV and IRR are shown graphically in figures 22.1, 22.2, 22.3 and 22.4. While project revenues are directly proportional to gold price, mill recovery and grade, the NPV (5%) and IRR of the O’Brien Project are highly sensitive to these factors. They are also highly sensitive to changes in OPEX and moderately sensitive to changes in CAPEX.

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Table 22.3 – Sensitivity analysis of economic parameters on after-tax NPV at 5% (millions $)

Figure 22.1 – Sensitivity analysis of economic parameters on after-tax NPV at 5% (millions $)

-30% -20% -10% Base Case scenario 10% 20% 30%

Revenue (44.47) (29.26) (15.41) (1.91) 8.17 17.84 27.36 Opex 18.40 11.60 5.05 (1.91) (11.11) (20.50) (30.21)

Capex 9.28 5.76 2.00 (1.91) (6.89) (12.21) (17.72)

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Table 22.4 – Sensitivity analysis of grade and Gold Price on after-tax NPV at 5% (millions $)

-30% -20% -10%

Base Case

scenario 10% 20% 30%

Grade (g/t) 8.39 7.75 7.10 6.46 5.81 5.16 4.52 Gold price

(US$/oz) 826 944 1062 1180 1298 1416 1534 Resulting NPV (M$) (44.47) (29.26) (15.41) (1.91) 8.17 17.84 27.36

Figure 22.2 – Sensitivity analysis of grade on after-tax NPV at 5% (millions $)

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Table 22.5 – Sensitivity analysis of economic parameters on after-tax IRR


Revenue -47% -26% -11% 3% 13% 21% 29% Opex 23% 16% 10% 3% -6% -15% -26%

Capex 18% 12% 7% 3% -1% -5% -9%

Figure 22.3 – Sensitivity analysis of economic parameters on after-tax IRR

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Table 22.6 – Sensitivity analysis of grade and Gold Price on after-tax IRR


Grade (g/t) 8.39 7.75 7.10 6.46 5.81 5.16 4.52

Gold price (US$/oz) 826 944 1062 1180 1298 1416 1534

IRR -47% -26% -11% 3% 13% 21% 29%

Figure 22.4 – Sensitivity analysis of grade on after-tax IRR

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23. ADJACENT PROPERTIES

A lot of exploration work and mining has been and continues to be conducted in the vicinity of the O'Brien Project. A number of producers and mineralized occurrences are found on adjacent properties within a few kilometres of the project. For the purposes of this report, the properties adjacent to the O'Brien Project are held by the following companies: Agnico-Eagle Mines (to the north); Cadillac Ventures Inc. (to the west); Globex Mining Enterprises (to the east), and 9265-9911 Québec Inc. (to the south).

Agnico-Eagle Mines Ltd Property Two major deposits are found on this property held by Agnico-Eagle Mines. The Bousquet-1 and -2 deposits are located about 7 km WNW of the resource area presented in this report. They were mined by Lac Minerals Ltd between 1979 and 1996. In 1996, production totalled 10.8 Mt at 5.96 g/t Au (Beaudoin et al, 2014). Along the same stratigraphic horizon as the Bousquet deposits, and less than 2 km to the east, the LaRonde mine has been in operation since 1988, and has produced 4.4 Moz of gold as well as valuable by-products (silver, zinc, copper and lead). The mine still has 3.9 Moz of gold in proven and probable reserves (24 Mt grading 5.0 g/t Au). The deep extension of the LaRonde mine achieved commercial production in November 2011, and is the focus of mining activities going forward with an estimated mine life that will last until 2025 (Agnico-Eagle website). The stratigraphic horizon related to the Bousquet and LaRonde-Dumagami deposits is located within the bimodal volcanics of the Blake River Group. These deposits are described as gold-rich VMS deposits and cannot be compared or associated with the deposits found on the O’Brien Project. They are located on a different stratigraphic horizon, about 2 km north of the resource estimate area presented in this report.

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Figure 23.1 – Adjacent properties of the O’Brien Project, showing past and current producers.

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New Alger Property In January of 2013, Cadillac Ventures Inc. announced that it has entered into an agreement with Renforth Resources Inc. to sell to Cadillac's 100% interest the New Alger property. The New Alger Project consists of two areas of gold occurrences, the Thompson-Cadillac Mine Area and the Pontiac Vein System. The Thompson Cadillac mine was discovered in 1924, and the property was first staked by E. J. Thompson during the same gold rush that discovered the O’Brien mine. The Thompson Cadillac mine is located just over 2 km west of the resource area presented in this report, and a few hundred metres west of the O’Brien property limits. It is located on the same stratigraphic horizon as the resource area presented in this report, and it shares the same orogenic-type geological setting. Gold mineralization is found in quartz veins associated with the CLLFZ, within tension fractures located in a conglomerate unit and basalts from the Piché Group. The mineralization is associated with arsenopyrite, pyrrhotite and pyrite. Free gold is also locally found. A new resource estimate for the New Alger deposit from April 2014 reports inferred resources of 3,007,000 t at a grade of 2.08 g/t Au for 201,000 oz Au (Wellstead and Newton, 2014). The Pontiac Vein System is a recent discovery located south of the mine, this is also a surface occurrence of gold in quartz veins, traced on surface aver 450 m.

Ironwood Project (modified from Pressaco, 2008) The project hosts two former gold producers: the Central-Cadillac mine and the Wood Cadillac mine. The Central Cadillac mine was found in 1933 and is localized 3 km east of the resource area presented in this report. From 1939 to 1943, production from the Central Cadillac mine was 185,541t at 5.14 g/t Au (954 kg Au and 115 kg Ag). From June of 1947 to August of 1949, 233,329 t at 4.33 g/t Au (1,010 kg Au and 130 kg Ag) were extracted but production came mostly from the Wood Cadillac segment without precisions of the contribution from the Central Cadillac. Still, production from these two periods totals 418,870 t at 4.69 g/t Au (1,964 kg Au). Mineralization in this deposit is also orogenic, closely related to the CLLFZ. Most of the mineralization comes from horizontal quartz-tourmaline veins found in a 15-m interval between the CLLFZ and iron formations. The veins and their strongly tourmalinized wallrock are slightly mineralized with pyrite, arsenopyrite and free gold. The veins also contain chalcopyrite and massive scheelite. Late quartz veinlets containing gold crosscut the older mineralized veins as well as silicified greywackes. Gold mineralization associated with arsenopyrite and pyrite was also found in talc-chlorite schists of the CLLFZ. Since 2004, the property has been explored by a joint venture between Globex Mining Enterprises Inc. and Queenston Mining Inc. The exploration work concentrated on the Ironwood deposit where gold mineralization is associated with an alteration assemblage of pyrrhotite-arsenopyrite-pyrite (± calcite/quartz) that is hosted by an oxide iron formation. A mineral resource estimate completed in 2008 indicates that the Ironwood deposit contains 243,200 t of inferred resources grading 17.26 g/t Au (Pressaco, 2008).

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Comments on Item 23 InnovExplo has been unable to verify the above information for adjacent properties near the O’Brien Project. The presence of significant mineralization on these adjacent properties is not necessarily indicative of similar mineralization on the O’Brien Project. Moreover, InnovExplo did not review the technical and economic parameters used to produce the mineral resource estimates for these adjacent properties.

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24. OTHER RELEVANT DATA AND INFORMATION

Additional information is not required to make this technical report understandable and not misleading.

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25. CONCLUSIONS

The principal objective of the issuer requesting a PEA for the O’Brien Project was to validate the technical and logistical advantages of the O’Brien project and estimate the initial investment for a production scenario. This technical report presented herein meet this objective. InnovExplo, WSP and Lamont concludes that the PEA presented herein demonstrate that to advance the O’Brien Project further, additional resources would need to be identify in order to provide economic robustness to lead to the development of a mine.

Mineral Resource Estimate The recent updated mineral resource estimate for the O’Brien Project had the objective of using recently compiled and validated historical diamond drill holes covering the area of the 36E and Kewagama areas. InnovExplo created a litho-structural model of the O’Brien Project using all available geological and analytical information. The following summarizes the approach and methodology used to create the mineralized zone wireframe model:

The new litho-structural model was used as the basis for defining mineralized zones.

Fifty-five (55) mineralized zones defined by grade continuity were modelled. Two dilution envelopes containing lower grade intervals surrounding

mineralized zones were modelled. The interpolation of the mineralized zones was constrained by the wireframes.

After conducting a detailed review of all pertinent information and completing the 2015 Mineral Resource Estimate, InnovExplo concludes the following:

Geological and grade continuity were demonstrated for the 55 gold-bearing zones of the O’Brien Project.

The additional compiled historical drill holes provided sufficient information to update the previous (2013) mineral resource estimate.

The estimate of Indicated Resources now stands at 119,819 oz (570,800 t at 6.53 g/t Au), and total Inferred Resources at 188,166 oz (918,300 t at 6.38 g/t Au).

The 2015 Indicated Resources represent a 13% increase in ounces compared to the 2013 estimate. The 2015 Inferred Resources represent a 181% increase in total ounces compared to the 2013 estimate. Grade increased by 0.6% in the Indicated category, whereas it decreased by 12% in the Inferred category. Note that additional ground was added to the resource area, and these figures should not be solely seen as an increase/decrease within the previous (2013) resource estimate area.

It is likely that additional diamond drilling on multiple zones would upgrade some of the Inferred Resources to Indicated Resources.

There is also the potential for upgrading some of the Indicated Resources to Measured Resources through detailed geological mapping, infill drilling and systematic channel sampling from the underground workings.

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InnovExplo also believes there are several opportunities to add additional resources to the O’Brien Project. After conducting a detailed review of all pertinent information, InnovExplo concludes the following:

41 targets, located down plunge of currently known ore shoots or having a high probability of identifying new ore shoots, were identified in the vicinity of the current resource estimate model.

Additionally, there are 47 targets with a high probability of identifying extensions of already defined resources within close proximity to historical underground workings or preliminary planned stopes.

InnovExplo believes there are several opportunities to make new discoveries outside the current resource estimate with twelve (12) Type 3 exploration targets.

Finally, mineralization is likely to remain in the old O’Brien mine area. Compilation of historical data in this area is likely to yield several exploration targets.

Metallurgy and Milling

There are many historical documents relating to the O’Brien Project area. Several test programs have been carried out since the 1970s. These were executed by various laboratories. The relationship between historical results and the area that is being studied is complex. Most of the time, samples were identified under the name of the zone. However, these names have changed over time, depending on which company owned the deposit. Nevertheless, these data provide an overview of the mineralogy, treatment methods and gold recoveries that may be obtained for samples taken from this area. The O'Brien Project, as currently defined, covers the 36E and Kewagama areas. The 36E area is divided into four zones: Upper West, West Central, West and Lower Central. The Kewagama area covers the eastern sector. In 2014, new laboratory testwork was undertaken on samples from the 36E area by the URSTM (Bouzahzah et al., 2014). In view of potential mining activities, custom milling will be the preferred option. The recent metallurgical testwork has demonstrated the amenability of O’Brien mineralized material to the gravity, leaching and flotation processes. The O’Brien Project is planned for a five-year period at a production rate of 500 tpd. Five gold concentrators located within a 75-km radius were then identified as being able to potentially process the O’Brien material: the Kiena Mill, the Sigma-Lamaque Complex, the Camflo Mill, the Westwood Mill and the Aurbel Mill. The following Table summarizes the main features of these milling options.

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Potential plants for custom milling (Table 17.1)




48 km Closed NA


Integra Gold


Leaching/CIP

1,200 to 2,400 tpd


Camflo Mill

Richmont Mines


800 to 1,200 tpd


Westwood Mill

Iamgold Gravity Concentration &

Leaching/CIP Or


Flotation

2,400 tpd

800 tpd

19 km Operating Yes

Aurbel Mill QMX Gravity Concentration &

Flotation & Flotation Concentrate

Leaching

500 to 800 tpd


conditions

The companies were contacted to find out their interest in performing custom milling. The Westwood and Aurbel mills have shown interest. This PEA is based on the use of the Westwood mill. The proximity of the O'Brien Project helps reduce transportation costs. In addition, the plant gives no restriction for environmental treatment. A trade-off study was conducted to compare treatment costs and potential recoveries for the two flowsheets available at the Westwood mill, see the following table. Trade-off study (Table 17.2)

Gravity/Flotation Gravity/Cyanidation

Gold value


Recovery % 94.52 91.53

Total3 C$/t 289 280

Milling cost



Smelting $/t 45 NA

Total $/t 81.78 36.78 1 Based on mining plan 2 URSTM test KN-F-3 3 Assumption section 17.1.2 BASED ON Gold PRICE at C$1475 /oz

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The selling cost was estimated based on information from similar projects. Preparation and trucking quotations were obtained from suppliers. The budgetary custom milling cost was estimated by the mill based on current knowledge of the ore. However, prices may be adjusted when additional information becomes available. Westwood's gravity and CIP circuit appears to be a good compromise based on the URSTM metallurgical results and the above considerations. The recovery will be lower but the treatment costs are significantly less. However, further work is required to validate the amount of free gold and the recovery by leaching process and then, determine a specific flowsheet that will optimize themetallurgical performance.

Environment

The area where the future mining activities will take place has already been impacted by previous mining activity. The area that is planned for development is adjacent to the previous (removed) infrastructure that was present on the Kewagama mine site. The Project activities should be constrained to an area that is less than 15 hectares. To obtain permits for the project, an environmental baseline study is required. This study will define the receiving environment before project development including the physical, biological and social environmental aspects. For this type of project, the study area should cover the location of the infrastructure within the project area that covers at least 15 hectares. For this project, no Environmental Impact Assessment will be required, as the Project remains lower than 2000 tpd (EQA Q-2, r.23) and no Physical Activities (SOR/2012-147) could trigger the Federal Process. Permits will be mainly issued by the “Ministère de l’Énergie et des Ressources Naturelles” and by the “Ministère du Développment durable, de l’Environnement et de la Lutte contre les Changements Climatiques”. From 2012 to 2014, Radisson conducted a geochemical characterization study of ore and waste rock samples. The majority of waste rock samples show no potential for acid generation but results indicate that all ore samples show a potential for acid generation. Samples of waste rock and ore have also been tested for their metal leaching (ML) potential. According with definition of Quebec’s Directive 019 and TCLP results, both waste rock and ore are leachable for some metals. The management of waste rock pile, ore stockpile as well as surface run-off were deisgned accordingly. Mine closure and rehabilitation cost have been estimated at $ 3.6 M. The closure cost estimate is based on capping the waste rock pile with an impermeable cover to limit infiltration and on the re-vegetation of the overburden layer that will cover the waste rock pile.

Capital and operating cost The PEA is based on capital pricing as of the third quarter of 2015. The PES assumes that the development and mining of the mine will be done by contractors and that they will supply the mobile equipment.

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The capitals costs were estimated using the following sources of information:

Quotes from equipment suppliers Comparable installations at other mining projects Contractor costs InnovExplo’s internal database

The capital cost estimates are accurate within ±20%. The preproduction costs are estimated at $36,76M, net of production revenue received during the second year of the preproduction period ($19,11M). Preproduction capital costs are minimal given that there is no need to build processing and tailings facilities. Preproduction is anticipated to take 2 years with the majority of proceeds used for ramp construction and for sufficient development of mineralized zones, or working faces, to conduct mining at the proposed mining rate and mill throughput. Sustaining capital is estimated at $21.35 million, including $3.7 million for final closure costs and considering a salvage value to $1,46M. Capital cost estimate (Table 21.1)

Description Pre-production Sustaining Total cost

Capitalized operating cost 21.33 M$ 21.33 M$ Capitalized revenue -19.30 M$ -19.30 M$ Royalty payment 1.00 M$ 1.00 M$ Development 20.01 M$ 17.13 M$ 37.14 M$ Mobile Equipment 0.21 M$ 0.18 M$ 0.39 M$ Surface infrastructure 6.45 M$ 0.02 M$ 6.48 M$ Mine service infrastructure 7.29 M$ 0.78 M$ 8.06 M$ Closure cost 3.70 M$ 3.70 M$ Salvage value -1.46 M$ -1.46 M$ EPCM 0.77 M$ 0.77 M$ Total 36.76 M$ 21.35 M$ 58.12 M$

Operating costs are estimated in 2015 Canadian dollars with no allowance for escalation. The total operating cost and average unit operating costs are summarized in the following table. The overall unit operating cost is $177.10 per tonne. Operating costs are summarized below for the production period.

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Summary of total operating costs (Table 21.9)


Definition drilling and sampling 2.47 M$ 3,85 20,29 Stope development 22.09 M$ 34,38 181,16 Contractor indirect cost 18.06 M$ 28,11 148,11 Mining cost 27.30 M$ 42,48 223,84 O'Brien staff and general 12.38 M$ 19,27 101,53 Energy cost 5.89 M$ 9,17 48,32 Milling and transportation 23.64 M$ 36,78 193,80 Environment 1.97 M$ 3,06 16,14 Total 113.81 M$ 177,10 933,18

Mining Plan The proposed mining plan for the O’Brien Project was prepared using the inferred and indicated resources estimated by InnovExplo. Due to the narrow vein nature of the orebody, two (2) underground mining methods were considered in the study, modified Avoca and long-hole mining with captive sublevels. The mining plan for the O’Brien Project comprises a combination of conventional and mechanized mining. The approach in this study has been to prioritize the modified Avoca mining method when possible. When this approach was not convenient, long-hole mining with captive sublevels was selected. The mineralized material will be transported to surface using a combination of 3.5-cubic-yard to 6-cubic-yard scoop trams and 30-tonne trucks. Waste material will be used to backfill mined out stopes as much as possible or will be brought to surface and stored on a dedicated waste pad. The current PEA is based on an underground mine with access by decline to a vertical depth of 550 metres in the 36E area and 250 metres in the Kewagama area. The production drifts will be accessed via crosscuts connecting to the ramp. A portion of the resources will be mined using captive methods, however haulage will always be mechanized. The mineral resource block model prepared by InnovExplo was used for the PEA. First, the resources available for mining were defined by creating the stope geometry in the block model at a cut-off grade of 3.5 g/t. Then a second triage was done using a diluted cut-off grade of 4.01 g/t. The guideline used in the stope design was a minimum mining width of 1.8 metres for subvertical stopes. The subvertical structures were cut at 18-metre vertical intervals corresponding to access level elevations. The conversion of mineral resources to potential mineral reserves takes into account dilution and losses during mining operations. The mineral resources are already diluted to a minimum width of 1.8 metres.

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Mining recovery was established at 85%, to take into account pillar requirements. A 30% dilution was also taken into account for stope excavation. Finally, a 95% recovery was applied to account for mining operating losses. For stopes with a diluted grade of less than 5.0 g/t, an evaluation was made to determine the economic viability of each stope, considering the development required to access the stope. If the economic viability could not be justified, the stope was discarded. Following this exercise, that included mine dilution and mine recovery a total of 712,521 tonnes at 6.46 g/t (147,986 oz) was included in the mine plan. Mine development will be accelerated in the first two years of the project to provide a degree of flexibility in terms of access, which should facilitate scheduling during the production period. The development sequence will ensure that many stopes are available for mining at a number of different locations at any given time. However, some of the stopes can only be mined at the end of the mine life since they are located directly over or under the level, therefore preventing any further access on that level when mined. The expected average daily production rate during the production period is estimated in this PEA between 450 and 500 t/day. The overall project mine life is expected to be approximately 6 years, including a two-year pre-production period. In the opinion of author Laurent Roy, Eng., the mine plan should be achievable given the flexibility and number of available working places. The following table summarizes the annual tonnage distribution according to the mine plan. Mine plan tonnage distribution (Table 16.4)





Grade (g/t) 7.05 6.47 6.87 7.04 5.50 6.53 6.46

Financial analysis An after-tax model was developed for the O’Brien Project. All costs are in 2015 Canadian dollars with no allowance for inflation or escalation. Income taxes are calculated in accordance with the federal and provincial tax legislations relating to mining companies. The calculations were made by Lucie Chouinard of Raymond Chabot. The federal income tax rate is 15% and the combined provincial income tax rate is 11.9%.

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Québec mining duties are calculated in accordance with Bill 55, which contains amendments to Québec’s Mining Tax Act and received its first reading in the Québec legislature on November 12, 2013. The Kewagama property consisted of a contiguous block comprising three (3) mining claims covering an aggregate area of 112.07 hectares. Radisson owned a 100% interest in the Kewagama property, with a 2% NSR royalty payable to KWG Resources Inc. in the event of commercial production. In addition, a $1,000,000 payment must be made to Breakwater Resources Ltd (now Nyrstar) upon commencement of commercial production on either one of the O’Brien or Kewagama properties, against which shall be deducted any costs required to restore the O’Brien tailing ponds. In the cash flow analysis, this royalty was considered on all ounces produced from the Kewagama property. The economic evaluation was performed using the Internal Rate of Return (IRR) and the Net Present Value (NPV) methods. This Preliminary Economic Assessment (PEA) is preliminary in nature as it includes Inferred Mineral Resources that are too speculative geologically to have economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the PEA will be realized. The following parameters were considered in the financial analysis.



produced from the Kewagama property. A residual fiscal base of $ 5.8M was considered in the tax estimation regarding

previous expenses by Radisson on the O’Brien Project.Resources as presented in Section 14;

Resources as presented in Section 14. Future annual cash flow estimates based on grade, gold recoveries and cost

estimates as previously discussed in this Report. 69,864 tonnes of mineralized material to be processed during the pre-

production period, deemed as capital production and not included in production nor the revenue derived from it.

The main parameters and cash flow analysis results for the entire project are presented in the following table.

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Cash flow analysis summary (Table 22.1)












PRE-TAX




AFTER-TAX


IRR (%) 3.15%


Risks and Opportunities Table 25.1 identifies the significant internal risks, potential impacts and possible risk mitigation measures that could affect the economic outcome of the project. The list does not include the external risks that apply to all mining projects (e.g., changes in metal prices, exchange rates, availability of investment capital, change in government regulations, etc.). Significant opportunities that could improve the economics, timing and permitting of the project are identified in Table 25.2. Further information and study is required before these opportunities can be included in the project economics.

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Table 25.1 – Risks of the O’Brien Project RISK Potential Impact Possible Risk Mitigation

Proximity of the historical O’Brien mine where

environmental, economic, and/or technical potential

issues could arise from the presence of 8,938 barrels of

arsenic trioxide stored underground at level 1500'

This underground storage site

is classified as a class 1 dangerous waste material site

by the GERLED group, a government entity with the mandate to catalogue and

monitor all known dangerous waste material sites in the

Province of Québec.

Although the current resources are located away from the storage facility, pumping water (which

would be necessary to bring the O’Brien Project to production) could potentially disturb the groundwater

and therefore affect the current situation, which is believed to be stable.

Historical precautions may have failed to contain the arsenic trioxide within the containment area over the

last 30 years.

In 1985, the Québec Ministry of Environment authorized the installation of new waterproof and

reinforced concrete plugs (2.3 m wide) at the entrance of each drift containing the barrels, and the

subsequent flooding of the mine;

Drilling from either surface or underground locations could breach the confinement facility.

A buffer zone around the drifts where the barrels are stored should be modelled in

3D, and this buffer zone should be excluded from any future drilling

program.

A hydrogeological study could be initiated to establish whether this area poses a risk and to characterize said

risk. Groundwater should be characterized in order to understand the impact that bringing the current resource

to production would have on the area.

Social acceptability Possibility that portions or the entirety of the O’Brien Project could not be explored or exploited.

Develop a pro-active and transparent strategy to identify all stakeholders and

develop a communication plan. Organize information sessions, publish information on the mining project, and meet with host

communities.

Metallurgical recoveries are based on limited testwork

Recovery might differ from what is currently being assumed.

Further variability testing of the deposit to confirm metallurgical conditions and

efficiencies.

The custom milling scenario is based on the fact that the

Westwood mill has expressed interest. The plant has

availability in the near future for custom feed. This scenario

could change.

Operating cost used in the PEA could be higher or lower depending on custom milling option available

at the time of operating the projet.

Free gold recovery The content of free gold recoverable by gravity has a

significant impact on the overall gold recovery. Historical data show that the free gold content varies

from one zone to another

Further metallurgical testwork must be conducted to confirm the gold recoveries for a gravity/CIP flowsheet. Only 2 tests

were done in the recent laboratory program. Most of the historical tests

gave lower recoveries for various cyanidation scenarios.

Limited testwork to determine whether waste rock would be

potentially acid generating (PAG)

Additional capital may be required to prepare a storage site for PAG waste.

Further testing to confirm whether the waste is PAG or non-acid generating

(NAG).

Surface and/or underground geotechnical evaluations not

available

The minimum mining width used for the resource estimate might need to be adjusted if assumptions

differ from reality.

The waste pile design is based on common geotechnical data, therefore footprint & pad

construction requirements might be reduced or enlarged, according to the surface geotechnical

evaluation results.

Geotechnical assessments at a larger scale to confirm rock quality

(underground and at surface) to validate assumptions.

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Table 25.2 – Opportunities of the O’Brien Project OPPORTUNITIES Explanation Potential benefit

Aditional geochemical tests on waste rock

Kinetic leaching tests could be done to confirm the ML potential of waste rock

If waste rock is not leachable, an impervious liner nor an impervious cover will be required.

Conduct specific gravity tests from core samples

Potential to increase the 2.67 g/cm3 specific gravity value currently used for

the resource estimate.

An increase in specific gravity increases the tonnage and therefore the ounces of gold.

Compilation of the old O’Brien mine workings

Potential to locate historical underground stopes, channel samples and drill holes with enough precision to

allow this area to be added to the geological model.

An entirely new area could be added that is not considered in the current resource estimate

presented in this report.

Compilation and validation of all remaining historical drill holes

Potential to upgrade the geological model and identify additional resources.

Adding resources increases the economic value of the mining project.

Compilation and validation of all historical underground channel

samples Potential to upgrade some indicated resources to the measured category.

Adding measured resources increases the economic value of the mining project.

Regarding specifically at Westwood mill opportunities:

A regrind mill could be

refurbished to reduce the grind size before cyanidation.

If the Westwood mill can provide a retention time of 72 hours, higher

recoveries could be achieved.

A bulk sample test should be performed in the Westwood mill.

Surface definition diamond drilling

Potential to upgrade some inferred resources to the indicated category.

Adding indicated resources increases the economic value of the mining project.


Extension of the mineralization within the drilling gap between the historical Kewagama mine

and the 36E area




Extention at depth of the ore

shoot originating in the Kewagama area




Subparallel mineralized zones

north and south of the currently identified zones



Identification of remaining mineralization in the old O’Brine mine area through compilation

and drilling



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26. RECOMMENDATIONS

Based on the PEA results, InnovExplo recommends a two-phase work program with the objective, in Phase 1, of increasing the continuity and tonnage of the resources to potentially improve the economics of the project and update the mineral resource estimate and the PEA. Contingent upon the success of Phase 1, InnovExplo recommends initiating a surface exploration and/or conversion drilling program and updating the resources accordingly. Supported by the new resource estimate, InnovExplo also recommends an underground development program. Phase 1 The property-scale compilation should be updated. As part of this compilation, the Company should complete a 3D compilation of the remaining historical openings of the old O’Brien mine, which would have a positive impact on locating all remaining historical underground drill holes and channel samples. The remaining historical data (drill holes, channel samples, etc.) should also be compiled, and the results used to upgrade the current model and resource estimate. Exploration drilling should target the currently identified areas of interest described in this report, but also target the discovery of additional zones over the entire project. If additional work proves to have a positive impact on the project, the current resource estimate should be updated to include compiled and validated historical drill holes, future drill holes, underground channel samples and updated 3D models of voids and mineralized zones. Based on the results of the updated resource estimate, the PEA should be updated. Regarding environmental matters, WSP recommends that additional site investigations, data collection, surveys and analyses be initiated as the project progresses to subsequent levels of design, to confirm or revise the current assumptions used for this study. Here is a non-exhaustive list of studies that are recommended:

Geochemical characterization of the waste rock, the ore and the tailings; Characterization of the mine water (groundwater); A baseline study of the receiving environment will be required for the

permitting application process; On-site evaluation of the current water management infrastructure (ponds,

ditches, liners, etc.); Geotechnical and hydrogeological studies for the waste rock, ore and

overburden pads; In an effort to potentially improve mill recovery, WSP recommends:

To conduct a metallurgical study to confirm and improve gold recoveries with a gravity/CIP flowsheet for 36E and Kewagama mineralized material: o Sample the entire mineralized area to evaluate the free gold content

per area/level;

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o Measure ball mill and abrasion work indexes to better estimate power and grinding media consumption;

o Conduct metallurgical tests in line with the Westwood mill flowsheet (gravity concentration followed by cyanidation of gravity tails) to optimize reagent consumption;

o Conduct metallurgical tests with a longer retention time; o Conduct further diagnostic testing (via QEMSCAN or other) to

determine the nature of the unleached gold; o Conduct a trade-off study to evaluate whether refurbishing the regrind

mill to obtain a finer grind and thus improve recoveries would be economically advantageous;

o Conduct corresponding metallurgical tests to determine the expected recoveries.

Phase 2 Contingent upon the success of Phase 1, InnovExplo recommends a Phase 2 that includes conversion drilling, which should be devoted to upgrading part of the inferred resources to the indicated category. It is recommended to update the mineral resource estimate to include all drilling results. Provision for an underground development program, namely including a bulk sampling campaign aimed at confirming the metallurgy and the continuity of mineralized zones, is considered in the recommended budget. It is recommended to obtain more detailed information about the Westwood process to better evaluate the gold recovery. Additional metallurgical testing should be initiated to improve knowledge through targeted laboratory tests on the cyanidation and gravity circuit conditions and to analyze the mineralogy of gold in discharges. There is a significant amount of data on flotation recovery. However, results for the two cyanidation tests conducted by URSTM are higher than reported historical data. These values should be confirmed to increase the level of confidence in the recovery rate. In addition, the two zones (36E and Kewagama) should be tested individually. The presence of free gold is crucial to recovery. Several historical tests indicate that recovery varies according to the mineralized zone.

InnovExplo and WSP have prepared a cost estimate for the recommended two-phase work program to serve as a guideline for the project. The budget for the proposed program is presented in Table 26.1. Expenditures for Phase 1 are estimated at C$3,772,000 (including 15% for contingencies). Expenditures for Phase 2 are estimated at C$19,280,000 (including 15% for contingencies). The grand total is C$23,050,000 (including 15% for contingencies). Phase 2 is contingent upon the success of Phase 1.

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Table 26.1 – Estimated costs for the recommended work program


Budget

Description

Cost

1a Property-scale compilation including 3D compilation of all remaining historical openings and historical data $100,000

1b Surface exploration drilling (all inclusive) 25,000 m $2,500,000

1c Stakeholder mapping, communication plan $50,000

1d Environmental studies $300,000

1e 3D model and resource estimate update $80,000

1f PEA update $250,000 Contingencies (~ 15%) $490,000 Phase 1 subtotal $3,770,000


Budget

Description

Cost 2a Surface exploration and/or conversion drilling (all inclusive) 25,000 m $2,500,000

2b 3D model and resource estimate update $80,000

2c Provision for an underground development program $13,500,000

2d Provision for environmental and hydrogeological characterization studies

$600,000

2e Metallurgical testing $100,000 Contingencies (~ 15%) $2,5000,000 Phase 2 subtotal $19,280,000

TOTAL (Phase 1 and Phase 2) C$ 23,050,000

InnovExplo is of the opinion that the recommended two-phase work program and proposed expenditures are appropriate and well thought out, and that the character of the O’Brien Project is of sufficient merit to justify the recommended program. InnovExplo believes that the proposed budget reasonably reflects the type and amount of the contemplated activities.

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27. REFERENCES

Ayer, J., Amelin, Y., Corfu, F., Kamo, S., Ketchum, J.F., Kwok, K., and Trowell, N.F., 2002a, Evolution of the Abitibi greenstone belt based on U-Pb geochronology: Autochthonous volcanic construction followed by plutonism, regional deformation and sedimentation: Precambrian Research, v. 115, p. 63–95.

Ayer, J.A., Ketchum, J., and Trowell, N.F., 2002b, New geochronological and neodymium isotopic results from the Abitibi greenstone belt, with emphasis on the timing and the tectonic implications of Neoarchean sedimentation and volcanism: Ontario Geological Survey Open File Report 6100, p. 5-1–5-16.

Ayer, J.A., Trowell, N.F., Amelin, Y., and Corfu, F., 1998, Geological compilation of the Abitibi greenstone belt: Toward a revised stratigraphy based on compilation and new geochronology results: Ontario Geological Survey Miscellaneous Paper 169, p. 4-1–4-14.

Barrie, C., 2006. High Resolution Horizontal Magnetic Gradient & XDS-VLF-EM Airborne Survey, O’Brien Kewagama Project, Cadillac, Quebec. Operations Report prepared by Terraquest Ltd for Radisson Mining Resources Inc. Report # B-193. 23 pages.

Bateman, R., Ayer, J.A., and Dubé, B., 2008, The Timmins-Porcupine gold camp, Ontario: Anatomy of an Archean greenstone belt and ontogeny of gold mineralizations. Economic Geology, v. 103, p. 1285–1308.

Beaudoin, G., Mercier-Langevin, P., Dubé, B., Taylor, B. E., 2014, Low-Temperature Alteration at the World-Class LaRonde Penna Archean Au-Rich Volcanogetic Massive Sulfide Deposit, Abitibi Subprovince, Québec, Canada: Evidence from Whole-Rock Oxygen Isotopes, Economic Geology, v. 109, pp. 167-182.

Bell, L. V., 1937. O’Brien Gold Mines Limited and Kewagama Gold Mines Ltd. In Mining Properties and Development in the Rouyn-Bell River District during 1936. Department of Mines and Fisheries, Province of Quebec, Canada. P. R. No. 116. Pages 34-35.

Bell, L. V., and MacLean, A., 1929. Cadillac O’Brien Property of M. J. O’Brien Co. In Report on the Geology of Bousquet-Cadillac Gold Area. Abitibi district, Annual Report Quebec Bureau of Mines for the Calender Year 1929, Part C. Pages 48-52.

Benn, K., and Peschler, A.P., 2005, A detachment fold model for fault zones in the Late Archean Abitibi greenstone belt: Tectonophysics, v. 400, p. 85–104.

Benn, K., Miles, W., Ghassemi, M. R., Gillet, J., 1994. Crustal structure and kinematic framework of the north-western Pontiac Subprovince, Québec: an integrated structural and geophysical study. Canadian Journal; of Earth Sciences, Vol. 31, pages 271-281.

Bisson, Y., 1994. Historical Facts of the O’Brien Mine. Internal Report of Radisson Mining Resources Inc. 14 pages.

Bisson, Y., 1995. 1994-1995 Winter Drilling Campaign Report. O’Brien Property Zone 36 East. Internal Report of Radisson Mining Resources Inc. 45 pages.

Bisson, Y., 1996. Rapport de campagne 1995-1996, Propriété O’Brien, Canton Cadillac. Ressources Minières Radisson Inc. Internal Report. 36 pages.

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Bisson, Y., 1998. Rapport de campagne 1996-1997, Propriété O’Brien, Canton Cadillac. Ressources Minières Radisson Inc. 77 pages. GM 56042.

Bisson, Y., 2004. Rapport préliminaire de campagne de forage 2004, Propriétés O’Brien et Kewagama, Canton Cadillac. Ressources Minières Radisson Inc. 42 pages. GM 61529.

Bouzahzah, H., Lelièvre, J., Villeneuve, M., Caractérisation minéralurgique, métallurgique et environnementale d’échantillons de la zone 36 du gisement O’Brien, October 2014.

Brereton, W. E., 1973. Report on Overburden Drilling, Kewagama Mine, Cadillac Township, Quebec. Report prepared by Driftex Limited for Derry, Michener & Booth. 6 pages. GM 30024.

Brethour, G. W., 1974. Underground Work report to April 1, 1974. Darius Gold Mines Ltd. 1 page. GM 29876.

Brethour, G. W., 1975a. Summary of U/G Exploration to date February 28, 1975. Darius Gold Mines Ltd. 1 page. GM 30838.

Brethour, G. W., 1975b. DHH logs of underground drilling. Darius Gold Mines Ltd. 39 pages. GM 31977.

Brethour, G. W., 1976. DHH logs of underground drilling. Darius Gold Mines Ltd. 102 pages. GM 32969.

Browm, R. A, 1948. O’Brien Mine. In Structure of Geology of Canadian Ore Deposits, A symposium Arranged by a Committee of the Geology Division Canadian Institute of Mining and Metallurgy Pages 809-816.

Charlton, J. D., 1994. O’Brien Property Reevaluation. Report prepared for Radisson Mining Resources Inc. Internal Report. 12 pages.

Chown, E. H., Daigneault, R., Mueller, W., and Mortensen, J., 1992. Tectonic evolution of the Northern Volcanic Zone of Abitibi Belt. Canadian Journal of Earth Sciences, v. 29, pp. 2211-2225.

Cooke, H. C., James, W. F., and Mawdsley, J. B., 1931. O’Brien Claims. In Geology and Ore Deposits of Rouyn-Harricanaw Region, Quebec, Memoir 166, Geological Survey, Canada Department of Mines. Pages 267-269.

Daigneault, R., Mueller, W.U., Chown, E.H., 2004. Abitibi greenstone belt plate tectonics: the diachronous history of arc development, accretion and collision. In Eriksson, P.G., Altermann, W., Nelson, D.R., Mueller, W.U., Catuneanu, O. (Eds.). The Precambrian Earth: Tempos and Events, Series: Developments in Precambrian geology, vol. 12, Elsevier, pages. 88–103.

Darius Gold Mines Inc., 1979. DDH logs of holes GF-79-1 to GF-79-24. 288 pages. GM 35984.

David, M., and Gauthier, E., 2012. Forage au diamant, propriétés O’Brien et Kewagama. Ressources Minières Radisson Inc. 3 pages. GM 67344.

de l’Étoile, R. and Salmon, B., 2013. Technical Report on the O’Brien Project Mineral Resource Estimate, Quebec, Canada. NI-43-101 Report prepared for Radisson Mining Resources Inc. by RPA Inc. 99 pages.

www.innovexplo.com


Dimroth, E, Imrech, L., Rocheleau, M., Goulet, N., 1982. Evolution of the south-central part of the Archean Abitibi Belt, Quebec. Part I: stratigraphy and paleostratigraphic model. Canadian Journal of Earth Sciences, Vol. 19, pages 1729-1758.

Dimroth, E, Imrech, L., Rocheleau, M., Goulet, N., 1983. Evolution of the south-central part of the Archean Abitibi Belt, Quebec. Part III: plutonic and metamorphic evolution and geotectonic model. Canadian Journal of Earth Sciences, Vol. 20, pages 1374-1388.

Dresser, J. A., and Denis, T. C., 1949. . O’Brien Gold Mines Limited and Kewagama Gold Mines Ltd. In Geology of Quebec, Vol III, Economic Geology, Department of Mines, Province of Quebec, Canada. , Geological Report 20. Pages 197-200.

Dubé, B. and Gosselin, P. 2007, Greenstone-Hosted Quartz-Carbonate Vein Deposits. In Goodfellow, W. D., ed. Mineral Deposits of Canada: A Synthesis of Major Deposit-Types, District Metallogeny, the Evolution of Geological Provinces, and Exploration Methods. Geological Association of Canada, Mineral Deposits Division, Special Publication No. 5. p. 49-73.

Dubé, B., O’Brien, S., and Dunning, G. R., 2001. Gold deposits in deformed terranes: examples of epithermal and quartz-carbonate shear-zone-related gold systems in the Newfoundland Appalachians and their implications for exploration. In North Atlantic Symposium, St-John’s, NF, Canada. Extended abstracts volume, May 27-30, 2001. p. 31-35.

Dubé, B., Poulsen K.H., and Guha, J., 1989. The effects of layer anisotropy on auriferous shear zones: The Norbeau mine, Quebec: Economic Geology, v. 84, p. 871-878.

Dugas, G., Duquette, G, and Latulippe, M., 1967. Annotated Bibliography on Metallic Mineralization in the regions of Noranda, Matagami, Val-d’Or, Chibougamau. Quebec Department of Natural Resources, Special Paper2. Page 41. ES 002.

Evans, L., 2007. Technical Report on the O’Brien mine Zone 36 East Mineral Resource Estimate, Cadillac, Quebec, Canada. NI-43-101 Report prepared for Radisson Mining Resources Inc. by Scott Wilson Roscoe Postle Associates Inc.94 pages.

Gagné, M. R., and Masson, J., 2013. A Step Foward! An Act to Amend the Minig Act (2013 S.Q., c. 32). Mining Bulletin. Fasken Martineau. 7 pages.

GEOSPEX Sciences inc., July 1998, « Caractérisation environnementale site minier O’Brien (Darius), Cadillac (Québec) ENVIRONNEMENTALE »

Glover, M. J., 1989. 1989 Diamond drilling summary report, O’Brien mine property, Project 5036. Breakwater Resources Ltd. Internal report. 31 pages.

Goutier, J., 1997, Géologie de la région de Destor: Ministère des Ressources naturelles du Québec 37 pages. RG 96-13.

Goutier, J., and Melançon, M., 2007, Compilation géologique de la Sous-province de l’Abitibi (version préliminaire): Ministère des Ressources naturelles et de la Faune du Québec.

www.innovexplo.com


Groves, D.I., Goldfarb, R.J., Know-Robinson, C.M., Ojala, J., Gardoll, S., Yun, G., and Holyland, P., 2000, Late-kinematic timing of orogenic gold deposits and significance for computer-based exploration techniques with emphasis on the Yilgarn block, Western Australia: Ore Geology Reviews, v. 17, p. 1-38.

Gunning, H. C., 1937. O’Brien Gold Mines Limited and Canadian Gold Operators Ltd. In Cadillac Area, Quebec, Memoir 206, Geological Survey, Canada Department of Mines. Pages 49-57.

Guzon, V., 2012. Mining Rights in the Province of Quebec. Blakes Bulletin Real Estate – Mining Tenures July 2012. Blake, Cassels & Graydon LLP. 7 pages.

Hagemann, S.G., and Cassidy, K.F., 2000, Archean orogenic lode gold deposits, in Hagemann, S.G., and Brown, P.E., eds., Gold in 2000: Society of Economic Geologists, Reviews in Economic Geology, v. 13, p. 9-68.

Hodgson, C.J., 1989, The structure of shear-related, vein-type gold deposits: A review: Ore Geology Reviews, v. 4, p. 635-678.

Jolly, W. T., 1978. Metamorphic history of the Archean Abitibi Belt. In Metamorphism in the Canadian Shield. Geological Survey of Canada, Paper 78-10, pp. 63-78.

Karpoff, B. S., and Evans, L., 1998. Resource Estimate of the O’Brien Mine Zone 36 East, Quebec. Report prepared by Roscoe Postle Associates Inc. for Radisson Mining Resources Inc. Internal Report. 100 pages.

Kelly, D., 2003. Résultats préliminaires des travaux de forages, propriété Kewagama. Ressources Minières Radisson. 12 pages. GM 60151.

Kerrich, R., and Cassidy, K.F., 1994. Temporal relationships of lode gold mineralization to accretion, magmatism, metamorphism and deformation – Archean to present: A review: Ore Geology Reviews, v. 9, p. 263-310.

Kroon, A. S., 1996. Qualification Report, O’Brien Property, Cadillac Township. Report prepared by Kilborn SNC-Lavalin for Radisson Mining Resources, Internal Report. 57 pages.

Kroon, A. S., 1997. Qualification Report Update, O’Brien Property, Cadillac Township. Report prepared by Kilborn SNC-Lavalin for Radisson Mining Resources, Internal Report. 66 pages.

Laboratoire de traitement du minerai de la mine Noranda, Tests de flottation sur le minerai de la mine d’or Darius, August 1980.

Lafleur, P. J., 1980. DDH logs of surface holes. Darius Gold Mines Inc. 22 pages. GM 38542.

Lafrance, B, Mercier-Langevin, P. Dubé. B., Galley, A.G., Hannington, M.D., Davis, D.W., Moorhead, J., Pilote, P., Mueller, W.U., 2003a. Carte synthèse de la Formation de Bousquet: partie ouest. Ministère des Ressources naturelles, de la Faune et des Parcs, Québec; DV 2003-08, echelle 1 : 20 000.

Lafrance, B, Mercier-Langevin, P. Dubé. B., Galley, A.G., Hannington, M.D., Davis, D.W., Moorhead, J., Pilote, P., Mueller, W.U., 2003a. Carte synthèse de la Formation de Bousquet: partie est. Ministère des Ressources naturelles, de la Faune et des Parcs, Québec; DV 2003-08, echelle 1 : 20 000.

Lafrance, B., and Doucet, P., 2005. The deep-seated gold potential of the Cadillac mining camp. Géologie Québec. Ministère des Ressources naturelles du Québec. 8 pages. PRO 2005-02.

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Lakefield Research of Canada Ltd, An Investigation of the Recovery of Gold from Samples Submitted by Darius Gold Mines Inc., Progress Report no. 1, May 12, 1976.

Lakefield Research of Canada Ltd, Microscopic Examination of Cadillac Gold Ore Samples Submitted by Azcon Corporation, Progress Report no.1, March 7, 1977.

Lakefield Research of Canada Ltd, Microscopic Examination of Cadillac Gold Ore Samples Submitted by Azcon Corporation, Progress Report no. 2, June 10, 1977.

Lakefield Research of Canada Ltd, Microscopic Examination of Cadillac Gold Ore Samples Submitted by Azcon Corporation, Progress Report no. 3, August 17, 1977.

Laronde, D. J., 1980. Report of Kewagama Gold Mines Ltd. Kewagama Gold Mines (Quebec) Limited. 13 pages. GM 36988.

Lelièvre, J., 1994. Revue de la métallurgie du minerai O’Brien (Analyses et recommandations). Centre spécialisé en technologie minérale, division de l’Abitibi-Témiscamingue. Internal report. 21 pages

Ludden, J.N., Hubert, C., and Gariépy, C., 1986, The tectonic evolution of the Abitibi greenstone belt of Canada: Geological Magazine, v. 123, p. 153-166.

MERQ-OGS, 1984, Lithostratigraphic map of the Abitibi subprovince: Ontario Geological Survey and Ministère de l’Énergie et des Ressources, Québec, Map 2484 and DV 83–16.

National Assembly, 2013. Bill 70 (2013, chapter 32) An Act to amend the Mining Act. Québec Official Publisher 2013. 32 pages.

O’Brien Gold Mine Ltd, 1929. Cross-sections and location plans of holes 1 to 12 drilled in 1929. GM 0751- A.

O’Brien Gold Mine Ltd, 1934-1945. DDH logs of holes 13 to 33 drilled between 1934-1945. 116 pages. GM 01641.

Paquet, A., 1989. Rapport de caractérisation Lieu 08-28B « dépôt de trioxide d’arsenic O’Brien-Darius ». Groupe de caractérisation et de surveillance des lieux d’élimination de et d’entreposage de déchets dangereux. Ministère de l’Environnement du Québec. Internal report. 55 pages.

Pelchat, C., 1996. 1995 Geological Mapping Report, Kewagama Property. Brewater Resources Ltd. 16 pages. GM 53820.

Pouliot, J.-L., 1964. Sondage au diamant. Kewagama Gold Mines Ltd.1 pages. GM 14780.

Poulsen, K. H., Robert, F., and Dubé, B., 2000. Geological classification of Canadian cold deposits. Geological Survey of Canada, Bulletin 540, 106 pages.

Powell, W. D., Carmichael, D. M., and Hodgon, C. J., 1993. Thermobarometry in a subgreenschist to greenschist transition in metabasite of the Abitibi greenstone belt, Superior Province, Canada. Journal of Metamorphic Geology, Vol. 11, pages 165-178.Quan, W. Ng See, 1987. O’Brien property, 2130, Diamond Drilling. Novamin Ressources Inc. 31 pages. GM 45791.

www.innovexplo.com


Pressacco, R., 2008, Technical Report for the Mineral Resource Estimate, Ironwood Project, Cadillac Township, Québec, (NTS 32D/01), Globex Mining Enterprises Inc and Queenston Mining Inc., 93 p.

Report from Genivar, July 2012: “Caractérisation physicochimique du minerai et des stériles à la propriété O’Brien, Cadillac”.

Report from the Unité de Recherche et de Services en Technologie Minérale (URSTM), October 2014: “Report PU-2013-12-860: Caractérisation minéralogique, métallurgique et environnementale d’échantillons de la zone 36 du gisement O’Brien”.

Richard, P.-L., Turcotte, B., and Carrier, A., 2015. Technical Report for the O’Brien Project, Abitibi, Quebec. Report prepared by InnovExplo Inc. for Radisson Mining Resources Inc. 126 pages.

Richard, P.-L., and Fallara, F., 2015. Target definition and drill program proposal for the O’Brien Projet. Internal report prepared by InnovExplo Inc. for Radisson Mining Resources. 51 pages.

Rive, M., 1980. Darius. In Rapport des géologues résidents 1979. Direction Générale de la recherche géologique et minérale. Ministère de l’Énergie et des Ressources. Pages 70-71. DPV 737.

Rive, M., 1981. Darius. In Rapport des géologues résidents 1980. Direction Générale de la recherche géologique et minérale. Ministère de l’Énergie et des Ressources. Page 4. DPV 814.

Rive, M., 1982. Darius. In Rapport des géologues résidents 1981. Direction Générale de la recherche géologique et minérale. Ministère de l’Énergie et des Ressources. Pages 10-11. DPV 868.

Robert, F., 1990, Structural setting and control of gold-quartz veins of the Val d'Or area, southeastern Abitibi subprovince, in Ho, S.E., Robert, F., and Groves, D.I., eds., Gold and Base-Metal Mineralization in the Abitibi subprovince, Canada, with Emphasis on the Quebec Segment: University of Western Australia, Short Course Notes, v. 24, p. 167-210

Robert, F., and Poulsen, K.H., 2001. Vein formation and deformation in greenstone gold deposits, in Richards, J.P., and Tosdal, R.M., eds., Structural Controls on Ore Genesis: Society of Economic Geologists, Reviews in Economic Geology, v. 14, p. 111-155.

Robert, F., Poulsen, K.H., and Dubé, B., 1994, Structural analysis of lode gold deposits in deformed terranes and its application: Geological Survey of Canada, Short course notes, Open File Report 2850, 140 p.

Sauvé, T., and Trudel, P., 1991. Géologie de la mine O’Brien, région de Cadillac. Ministère des Ressources naturelles du Québec. 39 pages. ET 89-07.

Schaaf, R. E., 1972. Report of Ferris (O’Brien property). Internal Report, 20 pages.

Schaaf, R. E., 1976a. Mineral Inventory Compilation, O’Brien Project, Cadillac Township, Quebec. Darius Gold Mines Inc. Progress Report 1 (No. 1 S Vein). 11 pages. GM 32217.

Schaaf, R. E., 1976c. Mineral Inventory Compilation, F9 Vein, No. 2 Shaft Zone, O’Brien Project, Cadillac Township, Quebec. Darius Gold Mines Inc. Progress Report 3. 5 pages. GM 32217. Schaaf, R. E., 1976b. Mineral Inventory

www.innovexplo.com


Compilation, O’Brien Project, Cadillac Township, Quebec. Darius Gold Mines Inc. Progress Report 2 (No. 1 N Vein). 3 pages. GM 32217.

Schaaf, R. E., 1976d. Mineral Inventory Compilation, F9 Vein, No. 3 Shaft Zone, O’Brien Project, Cadillac Township, Quebec. Darius Gold Mines Inc. Progress Report 4a. 1 pages. GM 32217.

Schaaf, R. E., 1976e. Mineral Inventory Compilation, F9 Vein, No. 3 Shaft Zone, O’Brien Project, Cadillac Township, Quebec. Darius Gold Mines Inc. Progress Report 4b. 11 pages. GM 32217.

Schaaf, R. E., 1976f. Mineral Inventory Compilation, H-4-14 Vein, O’Brien Project, Cadillac Township, Quebec. Darius Gold Mines Inc. Progress Report 5. 7 pages. GM 32217.

Schaaf, R. E., 1979. Exploration and development Merits Kewagama Gold Mines (Quebec) Ltd, Cadillac Gold Property, Cadillac Township, Quebec. Kewagama Gold Mines Ltd. 74 pages. GM 36987.

Scobie, A. G., 1972. An Investigation of the Recovery of Gold from a sample submitted by R. E. Schaaf, O’Brien Mill tailing Project. Report prepared by Lakefield research of Canada Limited. Internal Report

Thompson, I. S., 1974. Report on Diamond Drilling, August-September 1974, Cadillac Township, Quebec. Kewagama Gold Mines (Quebec) Ltd. 11 pages. GM 30587.

Thurston, P.C., and Chivers, K.M., 1990, Secular variation in greenstone sequence development emphasizing Superior province, Canada: Precambrian Research, v. 46, p. 21–58.

Thurston, P.C., Ayer, J.A., Goutier, J., and Hamilton, M.A., 2008, Depositional gaps in the Abitibi greenstone belt stratigraphy: A key to exploration for syngenetic mineralization. Economic Geology, v. 103, p. 1097−1134.

Trudel, P., Sauvé, P., Tourigny, G., Hubert, C., and Roy, L., 1992. Synthèse des caractéristiques géologiques des gisements d’or de la région de Cadillac (Abitibi). Ministère des Ressources naturelles du Québec.106 pages. MM 91-01.

Vaillant, R. L., and Hutchinson, R. W., 1982. Stratisgraphic and genesis of gold deposits, Bousquet region, northwestern Quebec. Canadian institute of Mining and metallurgy, special volume 24, pages 27-40.

Van the Wall, M., 1980. Darius. In Rapport des géologues résidents 1979. Direction Générale de la recherche géologique et minérale. Ministère de l’Énergie et des Ressources. Pages 70-71. DPV 737.

Vincent, R., 2009. Journeaux de sondages des campagnes de forages de 2006 à 2008, propriétés O’Brien et Kewagama, Cadillac Québec. GM-64406.

Wellstead, M., Newton, B.H., 2014, Technical Report on the 2014 ddh Program and Mineral Resource Estimate, New Alger Property, Abitibi-Temiscamingue, Québec, Renforth Resources Inc., Billiken Management Services, 133 p.

Wright, J. L., 1986. Magnetometer, VLF, and IP Geophysical Surveys, Fall 1985, O’Brien Mine property. Novamin Resources Inc. 18 pages. GM 43306.

www.innovexplo.com


Wyslouzil, D.M., Gochnauer, K., Mineralogical Examination of one Kewagama Mine Project Combined Mineralogical Sample, Progress Report no. 1, December 24, 1980.

Wyslouzil, D.M., Yen, W.T., An investigation of the recovery of gold from Kewagama Mine Project Samples, Progress Report no. 2, February 4, 1981.

Wyslouzil, D.M., Yen, W.T., An investigation of the recovery of gold from Kewagama Mine Project Samples, Progress Report no. 3, March 3, 1981.

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APPENDIX I – UNITS, CONVERSION FACTOR, ABBREVIATION

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Units Units in this report are metric unless otherwise specified. Precious metal content is reported in grams of metal per metric ton (g/t Au) except otherwise stated. Tonnage figures are dry metric tons unless otherwise stated. The ounces are in Troy ounces. Conversion factors for measurements

Imperial Unit Multiplied by Metric Unit 1 inch 25.4 mm

1 ft 0.3048 m 1 acre 0.405 ha

1 ounce (troy) 31.10348 g 1 pound (avdp) 0.454 kg

1 ton (short) 0.907 t 1 ounce (troy) / t (short) 34.286 g/t

Abbreviations

°C degrees Celsius oz troy ounces ha hectares avdp avoirdupois pound g grams st short ton kg kilograms oz/t ounces per short ton

mm millimetres t metric ton (tonne) cm centimetres Mt millions of tonnes m metres t.milled tonnes milled km kilometres t.moved tonnes moved

masl metres above sea level t.mined tonnes mined ’ or ft ft tpd / tpy metric tons per day/year cfm cubic ft per minute g/t grams per metric ton

m3/min cubic metres per minute ppb parts per billion usgpm US gallons per minute ppm parts per million

Mbs megabytes per second hp horsepower LOM life-of-mine MW megawatts $M millions of dollars kWh/t kilowatt-hours per tonne

$ or C$ or CAD Canadian dollars kV/kVA kilovolts/kilovolt-amps US$ or USD American dollars kPa/MPa kilo/mega pascals

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APPENDIX II – MINING RIGHTS IN THE PROVINCE OF QUÉBEC

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II.1 Mining Rights in the Province of Québec

The following discussion on the mining rights in the province of Québec was largely taken from

Guzon (2012) and Gagné and Masson (2013), and from the Act to Amend the Mining Act (“Bill

70”) assented on December 10, 2013 (National Assembly, 2013).

In the Province of Québec, mining is principally regulated by the provincial government. The

Ministry of Energy and Natural Resources (“MENR”; Ministère de l’Énergie et des Ressources

naturelles du Québec) is the provincial agency entrusted with the management of mineral

substances in Québec. The ownership and granting of mining titles for mineral substances are

primarily governed by the Mining Act (the “Act”) and related regulations. In Québec, land surface

rights are distinct property from mining rights. Rights in or over mineral substances in Québec

form part of the domain of the State (the public domain), subject to limited exceptions for privately

owned mineral substances. Mining titles for mineral substances within the public domain are

granted and managed by the MENR. The granting of mining rights in privately owned mineral

substances is a matter of private negotiations, although certain aspects of the exploration for and

mining of such mineral substances are governed by the Act. This section provides a brief overview

of the most common mining rights for mineral substances within the domain of the State.

II.1.1 The Claim A claim is the only exploration title for mineral substances (other than surface mineral substances,

or petroleum, natural gas and brine) currently issued in Québec. A claim gives its holder the

exclusive right to explore for such mineral substances on the land subject to the claim, but does

not entitle its holder to extract mineral substances, except for sampling and in limited quantities.

In order to mine mineral substances, the holder of a claim must obtain a mining lease. The

electronic map designation is the most common method of acquiring new claims from the MENR

whereby an applicant makes an online selection of available pre-mapped claims. In a few areas

defined by the government, claims can be obtained by staking.

A claim has a term of two years, which is renewable for additional two-year periods, subject to

performance of minimum exploration work on the claim and compliance with other requirements

set forth by the Act. In certain circumstances, if the work carried out in respect of a claim is

insufficient, or if no work has been carried out at all, it is possible for the claimholder to comply

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with the minimum work obligations by using work credits for exploration work conducted on

adjacent parcels, or by making a payment in lieu of the required work.

Additionally, since May 6, 2015, claim holder must submit to the MENR, on each claim registration

anniversary date, a report of the work performed on the claim in the previous year. Moreover, the

amount to be paid to renew a claim at the end of its term when the minimum prescribed work has

not been carried out now corresponds to twice the amount of the work required. Any excess

amount spent on work during the term of a claim can only be applied to the six subsequent renewal

periods (12 years in total). Holders of a mining lease or a mining concession are no longer able to

apply work carried out in respect of a mining lease or mining concession to renew claims.

II.1.2 The Mining Lease Mining leases and mining concessions are extraction (production) mining titles which give their

holder the exclusive right to mine mineral substances (other than surface mineral substances, or

petroleum, natural gas and brine). A mining lease is granted to the holder of one or several claims

upon proof of indications that a workable deposit could be present on the area covered by such

claims, and that the holder has complied with other requirements prescribed by the Act. A mining

lease has an initial term of 20 years, but may be renewed for three additional periods of 10 years

each. Under certain conditions, a mining lease may be renewed beyond the three statutory

renewal periods.

The Act (as amended by Bill 70) states that an application for a mining lease must be accompanied

by a project feasibility study, as well as a scoping and market study as regards to processing in

Québec. Holders of mining leases must then produce such a scoping and market study every 20

years. Bill 70 adds, as an additional condition for granting a mining lease, the issuance of a

certificate of authorization (CA) under the Environment Quality Act. The Minister may nevertheless

grant a mining lease if the time required to obtain the CA is unreasonable. A rehabilitation and

restoration plan must be approved by the Minister before any mining lease can be granted. In the

case of an open-pit mine, the plan must contain a backfill feasibility study. This last requirement

does not apply to mines in operation as of December 10, 2013. Bill 70 sets forth that the financial

guarantee to be provided by a holder of a mining lease be for an amount that corresponds to the

anticipated total cost of completing the work required under the rehabilitation and restoration plan.

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II.1.3 The Mining Concession Mining concessions were issued prior to January 1, 1966. After that date, grants of mining

concessions were replaced by grants of mining leases. Although similar in certain respects to

mining leases, mining concessions granted broader surface and mining rights, and they are not

limited in time.

A grantee must commence mining operations within five years from December 10, 2013. As is the

case for a holder of a mining lease, a grantee may be required by the government, on reasonable

grounds, to maximize the economic spinoffs within Québec of mining the mineral resources

authorized under the concession. It must also, within three years of commencing mining

operations and every 20 years thereafter, send the Minister a scoping and market study as regards

to processing in Québec.

II.1.4 Other Information

The claims, mining leases, mining concessions, exclusive leases for surface mineral substances,

and the licences and leases for petroleum, natural gas and underground reservoirs obtained from

the MENR may be sold, transferred, hypothecated or otherwise encumbered without the MENR’s

consent. However, a release from the MENR is required for a vendor or a transferee to be released

from its obligations and liabilities owing to the MENR related to the mine rehabilitation and

restoration plan associated with the alienated lease or mining concession. Such release can be

obtained when a third party purchaser assumes those obligations as part of a property transfer.

The transfers of mining titles, and the grants of hypothecs and other encumbrances in mining

rights, must be recorded in the register of real and immovable mining rights maintained by the

MENR and other applicable registers.

Under Bill 70, a lessee or grantee of a mining lease or a mining concession, on each anniversary

date of such lease or concession, must send the Minister a report showing the quantity and value

of ore extracted during the previous year, the duties paid under the Mining Tax Act and the overall

contributions paid during same period, as well as any other information as determined by

regulation.

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APPENDIX III – DETAILED LIST OF MINING TITLES

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Type of Mining Titles

Title Number NTS sheet Status Area (ha) Registration Date Expiration Date Holder

CDC 2169717 32D01 Active 12.33 August 7, 2008 August 6, 2016 Radisson Mining Resources Inc. (100%)

CDC 2169718 32D01 Active 35.61 August 7, 2008 August 6, 2016 Radisson Mining Resources Inc. (100%)

CDC 2429679 32D01 Active 57.37 July 30, 2015 March 1, 2017 Radisson Mining Resources Inc. (100%)



















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APPENDIX IV – DETAILED LIST OF HISTORICAL MINING TITLES

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Property Type of Mining Titles Title Number NTS sheet Township Status Area (ha) Holder Royalty

O'Brien CDC 2169717 32D01 Cadillac Active 12.33 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CDC 2169718 32D01 Cadillac Active 35.61 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CDC 2169719 32D01 Cadillac Active 26.71 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CDC 2169720 32D01 Cadillac Active 17.45 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CDC 2169721 32D01 Cadillac Active 3.91 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CDC 2169722 32D01 Cadillac Active 6.80 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CDC 2169723 32D01 Cadillac Active 19.90 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CDC 2169724 32D01 Cadillac Active 26.42 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CDC 2169725 32D01 Cadillac Active 36.85 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CDC 2169726 32D01 Cadillac Active 8.62 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3295141 32D01 Cadillac Active 81.47 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350491 32D01 Cadillac Active 17.47 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350492 32D01 Cadillac Active 14.55 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350493 32D01 Cadillac Active 16.26 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350494 32D01 Cadillac Active 14.64 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350495 32D01 Cadillac Active 15.92 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350501 32D01 Cadillac Active 16.42 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350502 32D01 Cadillac Active 8.92 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350504 32D01 Cadillac Active 9.86 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350505 32D01 Cadillac Active 6.87 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350511 32D01 Cadillac Active 6.78 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350512 32D01 Cadillac Active 9.76 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350513 32D01 Cadillac Active 11.47 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350514 32D01 Cadillac Active 9.76 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350515 32D01 Cadillac Active 6.93 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350521 32D01 Cadillac Active 6.55 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350522 32D01 Cadillac Active 6.96 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350523 32D01 Cadillac Active 13.98 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 3350524 32D01 Cadillac Active 4.80 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 4261152 32D01 Cadillac Active 9.84 Radisson Mining Resources Inc. (100%) No Royalty O'Brien CL 5274288 32D01 Cadillac Active 16.00 Radisson Mining Resources Inc. (100%) No Royalty Elmac CL 5274289 32D01 Cadillac Active 16.36 Radisson Mining Resources Inc. (100%) No Royalty Elmac CL 5274290 32D01 Cadillac Active 8.36 Radisson Mining Resources Inc. (100%) No Royalty

Kewagama CL C005451 32D01 Cadillac Active 21.17 Radisson Mining Resources Inc. (100%) 2% NSR to KWG Resources Inc.

Kewagama CL C006763 32D01 Cadillac Active 7.88 Radisson Mining Resources Inc. (100%) 2% NSR to KWG Resources Inc.

Kewagama CLD P007770 32D01 Cadillac Active 83.02 Radisson Mining Resources Inc. (100%) 2% NSR to KWG Resources Inc.

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APPENDIX V – SURFACE PLANS

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APPENDIX VI – ENVIRONMENTAL CHARACTERIZATION

Rapport final PU-2013-12-860

Caractérisation minéralogique,

métallurgique et environnementale d’échantillons de la zone 36

du gisement O’Brien

Pour :

Monsieur Mario Bouchard Ressources minières Radisson

93, chemin Trémoy Case postale 307

Rouyn-Noranda (Québec) J9X 1W4

Par :

Hassan Bouzahzah, Ph.D. Jean Lelièvre, ing., M.Sc.

Mathieu Villeneuve, M.Sc.A.

Unité de recherche et de service en technologie minérale

445, boul. de l’Université, Rouyn-Noranda (Québec) J9X 5E4 Téléphone : 819 762-0971, poste 2558 - Télécopieur : 819 797-6672

Octobre 2014

Caractérisation minéralogique, métallurgique et environnementale d’échantillons de la zone 36 du gisement O’Brien - PU-2013-12-860

Page i

Table des matières Page

Introduction …................................................................................................................................. 1

Partie 1 : Caractérisation métallurgique ......................................................................................... 2

1. Description des échantillons reçus ........................................................................................... 2

2. Évaluation de la proportion d’or libre récupérable par méthode gravimétrique ................... 4

3. Essais de flottation .................................................................................................................... 6

3.1 Essais préliminaires pour déterminer la granulométrie de flottation (F-1 à F-4) ........... 6

3.2 Essais combinant la concentration gravimétrique et la flottation .................................. 8

3.2.1 Essai KN-F-1 ............................................................................................................ 8

3.2.2 Essai de flottation KN-F-2 ....................................................................................... 9

3.3 Essais combinant la concentration gravimétrique et la flottation avec étapes de nettoyage .................................................................................................................. 11

3.3.1 Essais KN-F-3 ......................................................................................................... 11

3.3.2 Essai KN-F-4 .......................................................................................................... 14

3.3.3 Essai cyclique KN-F-5 ............................................................................................ 14

3.3.4 Essai cyclique KN-F-6-R ......................................................................................... 18

3.3.5 Comparaison des résultats des essais KN-F-3, KN-F-5 et KN-F-6 ......................... 20

3.4 Essais combinant la concentration gravimétrique et la cyanuration ............................ 21

3.4.1 Essais KN-CN-F-4 .................................................................................................. 21

3.4.2 Essai KN-CN-2 ....................................................................................................... 23

4. Conclusions – caractérisation métallurgique.......................................................................... 24

Partie 2 : Caractérisation environnementale ................................................................................ 26

1. Échantillons ............................................................................................................................. 26

2. Méthodes .................................................................................................................................... ................................................................................................................................ 28

2.1 Caractérisations chimiques des solides ......................................................................... 28

2.2 Essais PGA ...................................................................................................................... 29

2.3 Essais de lixiviation ......................................................................................................... 29

3. Résultats ................................................................................................................................ 31

3.1 Caractérisations chimiques des solides ......................................................................... 31

3.2 Essais statiques de détermination du PGA .................................................................... 33

3.3 Essais de lixiviations ....................................................................................................... 34

4. Conclusions – caractérisation environnementale .................................................................. 37

5. Recommandations .................................................................................................................. 37

6. Références .............................................................................................................................. 38

Partie 3 : Caractérisation minéralogique ...................................................................................... 39


Page ii

1. Étude …….. .............................................................................................................................. 39

2. Préparation de l’échantillon ................................................................................................... 39

3. Caractérisation de l’échantillon .............................................................................................. 39

3.1 Caractérisation chimique ............................................................................................... 39

3.2 Caractérisation minéralogique par microscopie optique et microsonde électronique ................................................................................................................... 39

4. Résultats ................................................................................................................................ 39

4.1 Analyse chimique ............................................................................................................. 39

4.2 Microscopie optique ........................................................................................................ 41

4.3 Microsonde électronique ................................................................................................ 41

4.4 Quantification de l’or associé à l’arsénopyrite dans l’échantillon « concentré de flottation » ................................................................................................................ 43

5. Conclusion – caractérisation minéralogique .......................................................................... 44

Annexe 1 : Essais métallurgiques détaillés

Annexe 2 : Protocole de cyanuration

Annexe 3 : Certificats d’analyses chimiques

Annexe 4 : Compositions chimiques élémentaires des pyrites, chalcopyrites et sphalérites par microsonde électronique

Annexe 5 : Photographies au microscope optique de tous les minéraux sulfurés analysés à la microsonde électronique

Liste des tableaux

Tableau 1 : Liste d'échantillons utilisés pour réaliser le lot composite utilisé pour les essais métallurgiques .......................................................................................... 2

Tableau 2 : Synthèse des essais d’évaluation de la proportion d’or libre récupérable .............. 5

Tableau 3 : Résumé des résultats obtenus des essais F-1 à F-4 ................................................. 7

Tableau 4 : Bilan (Au) métallurgique de l’essai KN-F-1 ............................................................... 8

Tableau 5 : Bilan métallurgique de l’essai KN-F-2 ..................................................................... 11

Tableau 6 : Bilan métallurgique (Au) de l’essai KN-F-3 ............................................................. 13

Tableau 7 : Bilan métallurgique (As) de l’essai KN-F-3 ............................................................. 13

Tableau 8 : Bilan métallurgique de l’essai cyclique KN-F-5....................................................... 17

Tableau 9 : Bilan métallurgique de l’essai KN-F-6-R ................................................................. 19

Tableau 10 : Tableau comparatif des résultats des essais KN-F-3, KN-F-5 et KN-F-6-R .............. 20

Tableau 11 : Bilan métallurgique de l’essai KN-CN-F-4 avec K80 = 102µ à la cyanuration ........ 22

Tableau 12 : Récupération et consommation en réactifs de la cyanuration uniquement (CN-F-4) ................................................................................................................... 22

Tableau 13 : Bilan métallurgique de l’essai KN-CN-2 avec K80 = 37µ à la cyanuration.............. 24


Page iii

Tableau 14 : Récupération et consommation en réactifs de la cyanuration uniquement (KN-CN-F-2) ............................................................................................................. 24

Tableau 15 : Liste des échantillons bruts constituant les quatre composites de l’étude environnementale .................................................................................................. 27

Tableau 16 : Différentes moutures des échantillons pour les essais environnementaux .......... 28

Tableau 17 : Analyses chimiques réalisées sur les échantillons de l’étude environnementale ................................................................................................................... 32

Tableau 18 : Comparaison des résultats des analyses chimiques avec les critères de la PPSRTC ................................................................................................................ 33

Tableau 19 : Bilan des essais statiques de détermination du PGA ............................................. 34

Tableau 20 : Résultats des lixiviations MA.100-Lix.com.1.1 (TCLP) réalisées sur tous les matériaux de l’étude environnementale ............................................................... 36

Tableau 21 : Analyses chimiques totale par ICP-AES de l’échantillon «concentré de flottation» ............................................................................................................... 40

Tableau 22 : Analyses chimiques des métaux lourds par ICP-AES de l’échantillon «concentré de flottation» ...................................................................................... 40

Tableau 23 : Résumé des observations au microscope optique des minéraux sulfurés de l’échantillon « concentré de flottation » ........................................................... 42

Tableau 24 : Récapitulatif des calculs pour l’estimation de l’or structural associé à l’arsénopyrite dans l’échantillon «concentré de flottation» ................................. 44

Liste des figures

Figure 1 : Localisation du site O’Brien de Ressources minières Radisson ............................... 1

Figure 2 : Diviseur rotatif utilisé pour la division en lots homogènes ...................................... 3

Figure 3 : Schéma de concassage des échantillons .................................................................. 3

Figure 4 : Schéma expérimental utilisé pour l’évaluation de la proportion d’or libre............. 4

Figure 5 : Montage expérimental pour l’évaluation de la proportion d’or libre ..................... 4

Figure 6 : Graphique de la récupération gravimétrique de l’or vs la granulométrie de broyage ................................................................................................................ 5

Figure 7 : Photographie de l’or libre sur la table de Mozley (Essai KN-4) ................................ 6

Figure 8 : Protocole de flottation utilisé pour les essais F-1 à F-4 ........................................... 6

Figure 9 : Graphique de la récupération en or et en argent, en fonction de la granulométrie de broyage ........................................................................................ 7

Figure 10 : Protocole expérimental de l’essai KN-F-1 ................................................................ 9

Figure 11 : Protocole expérimental de l’essai KN-F-2 .............................................................. 10



Figure 14 : Protocole de l’essai cyclique KN-F-5 ...................................................................... 16


Page iv

Figure 15 : Protocole expérimental de l’essai cyclique KN-F-6-R ............................................. 18

Figure 16 : Protocole de l’essai KN-CN-F-4 ............................................................................... 21

Figure 17 : Protocole expérimental de l’essai KN-CN-2 ........................................................... 23

Figure 18 : Photographies au microscope optique montrant les trois statuts de l’or ............. 41

Figure 19 : Représentation graphique des teneurs en or dans l’arsénopyrite dans l’échantillon «concentré de flottation» ................................................................ 42

Caractérisation minéralogique, métallurgique et environnementale d’échantillons de la zone 36 Est du gisement O’Brien - PU-2013-12-860

Page 1

Introduction

Monsieur Mario Bouchard, président de Ressources minières Radisson (ci-après « le client »), a contacté l’Unité de recherche et de service en technologie minérale de l’Université du Québec en Abitibi-Témiscamingue (URSTM-UQAT) au sujet de la réalisation de travaux de caractérisa-tions minéralogique, métallurgique et environnementale d’échantillons provenant de la zone 36 de l’ancienne mine O’Brien. Le site O’Brien est situé à moins d’un kilomètre au nord du village de Cadillac en Abitibi et à 50 km à l’est de la ville de Rouyn-Noranda (QC).

Figure 1 : Localisation du site O’Brien de Ressources minières Radisson

Le minerai de la zone 36 Est, de la mine O’Brien, est un minerai d’or composé de pyrite et d’arsénopyrite et contenant une certaine proportion d’or libre.


Page 2

Partie 1 : Caractérisation métallurgique

L’objectif principal des essais métallurgiques consistait à définir un schéma de traitement permettant la récupération gravimétrique maximale de l’or libre et la production d’un concentré « sulfures-or », par flottation, avec une teneur élevée. Le principal avantage anticipé de la production d’un concentré « sulfures-or » par flottation concerne la possibilité de traiter le concentré dans une autre usine ou directement à la fonderie. Ainsi, une telle méthode de traitement réduirait considérablement les coûts de la future usine étant donné la non nécessité du circuit de cyanuration. En cours de projet, il a été convenu d’évaluer l’alternative « concentration gravimétrique de l’or libre, suivie par la cyanuration » pour des fins de comparaison. Les essais métallurgiques ont été réalisés entre les mois d’avril et d’août 2014, dans les laboratoires du Cégep de l’Abitibi-Témiscamingue, par Jean Lelièvre, ing., M. Sc., pour l’Unité de recherche et de service en technologie minérale (URSTM). Les pyroanalyses ont été effectuées par Laboratoire Expert et les autres analyses chez Multilab, deux entreprises de Rouyn-Noranda. Les caractérisations minéralogique et environnementale ont été effectuées à partir des rejets et des concentrés provenant des essais métallurgiques (parties 2 et 3 de ce rapport).

1. Description des échantillons reçus

L’échantillon reçu pour l’essai minéralurgique était composé d’un total de six poches d’échantillons, combinés pour former un lot global. Le tableau 1 présente l’identification de chaque poche d’échantillons.

Tableau 1 : Liste d'échantillons utilisés pour réaliser le lot composite utilisé pour les essais métallurgiques

Poche #1 (PS)

Poche #2 (PC)

Poche #3 (PN)

Poche #4 (1S)

Poche #5 (1N)

Poche #6 (1X) + 8(FV) La totalité des échantillons reçus a tout d’abord été soumise à un concassage à moins de huit mailles, pour ensuite être homogénéisée et divisée en lots uniformes de 0,5 kg, à l’aide d’un diviseur rotatif.


Page 3

Figure 2 : Diviseur rotatif utilisé pour la division en lots homogènes

Figure 3 : Schéma de concassage des échantillons

Concasseur à mâchoires

1re division avec l’échantillonneur rotatif

2e division avec l’échantillonneur rotatif

Division en lots de 0,5 kg

Séchage 40o C

Concasseur à rouleaux


Page 4

2. Évaluation de la proportion d’or libre récupérable par méthode gravimétrique

Un total de quatre évaluations de la proportion d’or libre récupérable a été réalisé à quatre granulométries différentes. La figure 4 présente la démarche expérimentale utilisée pour l’évaluation de la proportion d’or libre récupérable.

Table de Mozzley

Pompe Masterflex

Knelson

Evaluation de la proportion d'or libre récupérable

concentré

Rejet table de Mozzley

Concentré or libre

Broyage

Rejet Knelson

Figure 4 : Schéma expérimental utilisé pour l’évaluation de la proportion d’or libre

Figure 5 : Montage expérimental pour l’évaluation de la proportion d’or libre

Pompe Masterflex


Page 5

Le tableau 2 présente les résultats obtenus des évaluations de la proportion d’or libre en fonction de la granulométrie de broyage.

Tableau 2 : Synthèse des essais d’évaluation de la proportion d’or libre récupérable

Réc. Au

Teneur du conc. or libre

(g Au/tm)

Rejet g Au/mt

Alim. calc.

g Au/tmRéc. Ag

Teneur du conc. or libre

(g Au/tm)

Rejet g Ag/tm

Alim. calc.

g Ag/tm

KN-1 137 µ 58,8% 50,4% 18905,2 5,47 11,03 15,5% 451,42 0,72 0,86

KN-2 105 µ 66,0% 58,9% 19961,7 6,00 14,58 27,4% 518,69 0,59 0,81

KN-3 90 µ 70,9% 59,0% 20968,0 4,37 10,67 23,4% 797,52 0,78 1,02

KN-4 74 µ 80,4% 60,2% 18158,6 4,60 11,54 44,9% 1071,83 0,50 0,91

Moyenne: 11,96 0,90

Au Ag

Evaluation proportion d'or libre

récupérable

Description Essai K80 (µm)

% < 200 mailles

On constate que la proportion d’or libre récupéré augmente avec la granulométrie de broyage mais très peu après 105µ. À partir d’environ 105µ, on atteint une récupération de l’or de 58,9 %, ce qui constitue une valeur assez habituelle pour les minerais contenant de l’or libre.

On remarque également les très faibles proportions d’argent libre qui sont récupérées. Un maximum de 44,9 % de l’argent libre a été récupéré à une granulométrie de 74µ.

La figure 6 présente la relation graphique entre la proportion d’or libre récupéré et la granulométrie de l’alimentation.

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

30 40 50 60 70 80 90 100 110 120 130 140 150

% o

r lib

re ré

cupé

ré

K80 (µm)

% or libre récupéré Vs K80Zone 36-Est - Ressources Radisson

Figure 6 : Graphique de la récupération gravimétrique de l’or vs la granulométrie de broyage


Page 6

Le graphique précédent montre une courbe très régulière. On constate que la récupération gravimétrique de l’or libre plafonne vers 100µ.

Figure 7 : Photographie de l’or libre sur la table de Mozley (Essai KN-4)

L’or libre observé sur la table de Mozley montre la présence d’or libre très grossier. La plus grosse particule observée a une dimension de 3299µ (3,3 mm).

3. Essais de flottation

3.1 Essais préliminaires pour déterminer la granulométrie de flottation (F-1 à F-4)

Les essais F-1 à F-4 ont été réalisés pour évaluer la récupération en fonction de la granulométrie de broyage. La figure 8 représente le protocole utilisé pour ces essais.

Broyag:

Conditionnement5 min pH nat = 8,2

250 g/tm CuSO4

15 g/tm A-40720 g/tm KAX

13 g/tm MIBC

1,0min 1,15mi 3,0min 2,0min

10 g/tm A-40715 g/tm KAX13 g/t MIBC

Conc. 1 Conc. 2 Conc. 3 Conc. 4

Rejet


Figure 8 : Protocole de flottation utilisé pour les essais F-1 à F-4


Page 7

Le tableau 3 résume les résultats obtenus des essais de flottation F-1 à F-4.

Tableau 3 : Résumé des résultats obtenus des essais F-1 à F-4

Réc. Au

Teneur moy. du

conc. (g Au/tm)

Rejet g Au/mt

Alim. calc.

g Au/tmRéc. Ag

Teneur moy. du

conc. (g Ag/tm)

Rejet g Ag/tm

Alim. calc.

g Ag/tm

F-1 139 µ 56,5% 10,4% 90,2% 64,9 0,82 7,47 46,0% 5,88 0,80 1,33

F-2 105 µ 66,0% 8,8% 94,0% 122,4 0,76 11,47 67,1% 8,43 0,40 1,11

F-3 73 µ 81,2% 11,1% 95,8% 100,9 0,55 11,69 55,8% 5,06 0,50 1,01

F-4 37 µ 97,7% 18,0% 94,9% 53,9 0,63 10,26 66,5% 4,53 0,50 1,23

Moyenne: 10,22 1,17

Ag

Essais de flottation

Description Essai K80 (µm)

% massique

au concentré

Au

% < 200 mailles

On constate que la récupération de l’or, pour tous les essais effectués, dépasse la valeur de 90 %.

La récupération maximale de 95,8 % a été obtenue avec une granulométrie de 73µ.

Les récupérations en argent sont beaucoup plus faibles avec une récupération maximale observée de 67 %.

La teneur calculée de l’alimentation est de 10,22 g Au/tm et de seulement 1,17 g Ag/tm. Donc, il y a très peu d’argent dans ce minerai.

La figure 9 illustre la relation entre la récupération par flottation et la granulométrie de broyage.

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

30 40 50 60 70 80 90 100 110 120 130 140 150

% ré

cupé

ratio

n

K80 (µm)

Essais de flottation F-1 à F-4% récupération Vs K80

Zone 36-Est - Ressources Radisson

Au

Ag

Figure 9 : Graphique de la récupération en or et en argent, en fonction de la granulométrie de broyage


Page 8

On observe une relation très régulière entre la récupération en or et la granulométrie de broyage. Celle obtenue pour l’argent est beaucoup moins précise mais moins importante, compte tenu de la valeur relative de l’or versus celle de l’argent.

On peut constater de nouveau qu’une granulométrie d’environ 73µ semble adéquate pour la flottation.

3.2 Essais combinant la concentration gravimétrique et la flottation

3.2.1 Essai KN-F-1

Deux essais ont été réalisés en combinant la concentration gravimétrique (Knelson + Mozley) et la flottation (voir figure 10). L’essai KN-F-1 a été réalisé avec une granulométrie de 102µ en effectuant d’abord la concentration gravimétrique de l’or, suivie de la flottation du rejet de la concentration gravimétrique sans étape préalable de rebroyage. Il faut également noter que cet essai ne comporte pas d’étape de nettoyage.

Tableau 4 : Bilan (Au) métallurgique de l’essai KN-F-1

Masse (g) % poids Teneur mg Au % distribution

0,28 0,03% 19700 g Au/tm 5,52 mg Au 55,3%

86,20 8,69% 44,27 g Au/tm 3,82 mg Au 38,3%

905,7 91,28% 0,71 g Au/tm 0,64 mg Au 6,4%

992,18 100% 10,05 g Au/tm 9,97 100%

55,3%38,3%93,6%Récupération or combinée:

Concentré d'or libre:

Concentré flottation

Rejet flottation

Alimentation calculée

Récupération or dans concentré or libre:Récupération or dans concentré flottation:

On observe une excellente récupération en or de 93,6 %;

La récupération gravimétrique de l’or de cet essai se situe à 55,3 %;

La teneur du concentré est de 44,27 g Au/ tm peut sembler un peu faible mais il faut considérer qu’aucune étape de nettoyage permettant de hausser la teneur n’a été incluse dans cet essai.


Page 9

Essai: KN-F-1

Description:

992,2 g

K80 = 102 µ

13,1 kPa

Table de Mozzley

0,28 g

% or récupéré = 55,3%

Essai Knelson + flottation (sans rebroyage avant la flottation)

(66,5% < 200 mailles)

Pompe Masterflex

Knelson

19700 g Au/tm

concentré


Concentré or libre

Broyage: : 20 kg de tiges 6' 45"50% solide1,0 kg de minerai/broyage

Rejet Knelson


250 g/tm CuSO4


13 g/tm MIBC



S-11 et S-18

S-10 et S-16


Figure 10 : Protocole expérimental de l’essai KN-F-1

3.2.2 Essai de flottation KN-F-2

L’essai KN-F-2 a été réalisé avec une granulométrie de 102µ en effectuant d’abord la concentration gravimétrique de l’or, suivie d’une étape de rebroyage et, ensuite, par la flottation du rejet de la concentration gravimétrique.


Page 10

Il faut également noter que cet essai ne comporte pas d’étape de nettoyage.

Essai: KN-F-2

Description:

997,2 g

K80 = 102 µ

13,1 kPa

Table de Mozzley

0,55 g


10263 g Au/tm

Pompe Masterflex

Knelson

Essai Knelson + flottation (avec rebroyage avant la flottation)

(66,5% < 200 mailles)

concentré


Concentré or libre


Rejet Knelson

Broyage: 2' 00" K80 = 73µ


250 g/tm CuSO4


13 g/tm MIBC



S-14 et S-19

S-13 et S-17




Page 11

Tableau 5 : Bilan métallurgique de l’essai KN-F-2


0,55 0,06% 10263 g Au/tm 5,64 mg Au 54,3%

93,20 9,35% 43,65 g Au/tm 4,07 mg Au 39,1%

903,4 90,60% 0,75 g Au/tm 0,68 mg Au 6,5%

997,15 100% 10,42 g Au/tm 10,39 100%

54,3%39,1%93,5%Récupération or combinée:


Rejet flottation


Concentré flottation

Récupération or dans concentré or libre:Récupération or dans concentré flottation:

La récupération combinée en or est de 93,5 %, ce qui est quasi identique à la récupération obtenue pour l’essai KN-F-1 sans rebroyage.

À partir uniquement de ces deux essais (KN-F-1 et KN-F-2), il pourrait sembler que le rebroyage du rejet gravimétrique ne soit pas justifié. Le seul bémol concerne cependant la relation démontrée précédemment par les essais F-1 à F-4, qui montraient une récupération supérieure à une granulométrie de 73µ. D’autres essais seraient requis pour trancher cette question.

3.3 Essais combinant la concentration gravimétrique et la flottation avec étapes de nettoyage

3.3.1 Essais KN-F-3

L’essai KN-F-3 a été réalisé en intégrant une étape de nettoyage du concentré de dégrossissage et en ajoutant une étape de nettoyage de l’épuiseur. Les deux concentrés sont ensuite combinés pour former le concentré final. Étant donné que ce protocole de flottation ne contient pas de recirculation, il n’est pas nécessaire d’effectuer d’essais cycliques pour ce protocole de flottation.


Page 12

Essai: KN-F-3Date: 17-mai-14

Description:

994,0 g

K80 = 102 µ

13,1 kPa

Table de Mozzley

0,29 g


S-30 et S-340,03% As0,57 g Au/tm

S-330,30% As1,5 g Au/tm

S-31 S-3520,0% As 5,51% As156,7 g Au/tm 27,0 g Au/tm

Essai Knelson + flottation (avec étapes de nettoyage)

(65,8% < 200 mailles)

Pompe Masterflex

Knelson

22947 g Au/tm

concentré


Concentré or libre


Rejet Knelson


250 g/tm CuSO4


13 g/tm MIBC

1,5min 1,5min 2,0min 2,0min




1,0min 2,0min



Page 13

Tableau 6 : Bilan métallurgique (Au) de l’essai KN-F-3


0,29 0,03% 22947 g Au/tm 6,65 mg Au 62,9%

18,10 1,82% 156,7 g Au/tm 2,84 mg Au 26,8%

18,40 1,85% 27,0 g Au/tm 0,50 mg Au 4,7%

36,50 3,67% 91,3 g Au/tm 3,33 mg Au 31,5%

45,30 4,56% 1,5 g Au/tm 0,07 mg Au 0,7%

911,9 91,74% 0,57 g Au/tm 0,52 mg Au 4,9%

957,2 96,30% 0,61 g Au/tm 0,58 mg Au 5,5%

993,99 100% 10,64 g Au/tm 10,57 mg Au 100%

62,9%31,5%94,5%Récupération combinée or libre + concentrés flottation :

Rejet nettoyeur épuiseur

Rejet épuiseur



Concentré 1er nettoyage

Concentré nettoyeur épuiseur

Récupération or dans concentré or libre:Récupération combinée des conc. 1er nettoyeur et conc. nettoyeur épuiseur:

Concentrés flottation combinés

Rejets combinés

La récupération obtenue de l’essai KN-F-3 est de 94,5 %, ce qui est excellent.

La concentration gravimétrique obtient une récupération de 62,9 %. Le tableau 7 présente le bilan métallurgique pour ce qui concerne l’arsenic.

Tableau 7 : Bilan métallurgique (As) de l’essai KN-F-3

Masse (g) % poids Teneur % As g As% distribution

As

18,10 1,82% 20,20% 3,66 g As 71,4%

18,40 1,85% 5,51% 1,01 g As 19,8%

36,50 3,67% 12,79% 4,67 g As 91,15%

45,30 4,56% 0,30% 0,14 g As 2,7%

911,9 91,77% 0,03% 0,32 g As 6,2%

957,2 96,33% 0,05% 0,45 g As 8,8%

993,70 100% 0,52% 5,12 g As 100%Alimentation calculée

Concentrés flottation combinés

Rejets combinés

Concentré 1er nettoyage

Concentré nettoyeur épuiseur

Rejet nettoyeur épuiseur

Rejet épuiseur

On observe tout d’abord que la teneur calculée de l’alimentation se situe à 0,52 % As.


Page 14

La teneur des concentrés combinés de flottation est de 12,79 % As. Cette valeur en arsenic est assez élevée et sera déterminante dans l’évaluation économique du coût de « smeltage » de ce concentré.

3.3.2 Essai KN-F-4

L’essai KN-F-4 a été réalisé dans le but d’aider à préciser la meilleure façon de recirculer les produits de la flottation. La figure 13 montre le protocole qui a été élaboré. Un des principes guidant la recirculation des produits est de tenter de combiner les flux ayant des teneurs similaires. En observant de plus près la figure 13, on peut constater que la recirculation du rejet du premier nettoyeur a avantage à se faire vers l’étape du nettoyeur/épuiseur. Pour ce qui concerne la recirculation du concentré du nettoyeur/épuiseur, la réponse n’est pas évidente mais il semble préférable, à priori, de le combiner directement au concentré final du premier nettoyage plutôt que le combiner à l’alimentation du premier nettoyage et, ainsi, diluer la teneur combinée de l’alimentation de l’étape de nettoyage. Cette façon de faire correspond au protocole utilisé pour l’essai KN-F-3; essai qui a obtenu d’excellents résultats.

3.3.3 Essai cyclique KN-F-5

L’essai KN-F-5 constitue un essai comportant quatre cycles (voir figure 14). Les essais cycliques permettent de prédire les résultats obtenus à partir d’un protocole comportant une recirculation, en circuit fermé, de certains produits. Pour l’essai KN-F-5, la recirculation du rejet du nettoyeur est effectuée à l’alimentation, tandis que celle du concentré du nettoyeur/épuiseur est effectuée à l’alimentation du premier nettoyeur.


Page 15

Essai: KN-F-4Date: 24-juin-14

Description:

992,7 g

K80 = 102 µ

13,1 kPa

Table de Mozzley

0,22 g


S-34-B0,55 g Au/tm

Rejets combinés0,57 g Au/tm

S-31-B S-33-B5,1 g Au/tm 1,30 g Au/tm

S-32-B13,7 g Au/tm

S-30-B119,7 g Au/tm

47,8 g Au/tm 4,2 g Au/tm

Essai Knelson + flottation (avec étapes de nettoyage)

(65,8% < 200 mailles)

Pompe Masterflex

Knelson

27095 g Au/tm

concentré


Concentré or libre


Rejet Knelson


250 g/tm CuSO4


13 g/tm MIBC





1,0min 2,0min

1er nettoyage Nettoyeur-épuiseur



Page 16

Essai: KN-F-5Date: 25-juin-14

Description:

4 kg

K80 = 102 µ

13,1 kPa

Table de Mozzley

0,63 g


Rejet 10,02% As0,82 g Au/mt

S-48 Rejet 20,51% As 0,17% As3,58 g Au/mt 6,79 g Au/mt

S-471,75% As

Pompe Masterflex

7,26% As 114,3 g Au/tm 94,5 g Au/tm

Essai Knelson + flottations cycliques (Cycle 4)

(65,8% < 200 mailles)

Knelson

62143 g Au/tm

Concentré final

concentré


Concentré or libre


Rejet Knelson


250 g/tm CuSO4


13 g/tm MIBC





1,5min 2,0min


Figure 14 : Protocole de l’essai cyclique KN-F-5


Page 17

Tableau 8 : Bilan métallurgique de l’essai cyclique KN-F-5

Cycle 1 Cycle 2 Cycle 3 Cycle 4

0,16 g 0,16 g 0,16 g 0,16 g

62143 g Au/mt 62143 g Au/mt 62143 g Au/mt 62143 g Au/mt

24,60 g 30,90 g 38,60 g 40,20 g

165,1 g Au/mt 122,9 g Au/mt 115,9 g Au/mt 94,5 g Au/mt

8,25% As 2,6% As 8,3% As 7,3% As

929,30 g 931,80 g 925,30 g 929,90 g

0,88 g Au/mt 0,92 g Au/mt 0,97 g Au/mt

0,02% As 0,02% As 0,02% As 0,02% As

962,86 g 964,06 g 994,76 g


0,24% As 0,11% As 0,35% As 0,32% As

% distribution or libre récupéré 67,9% 64,8% 67,4%

% récupération globale Au 94,3% 94,4% 93,6%

Concentré or libre

% distribution Au concentré flottation

26,4% 29,6% 26,2%

Alim. calc.

Rejet 1 + Rejet 2

Concentré final

La récupération en or de cet essai semble légèrement inférieure à celle obtenue notamment par l’essai KN-F-3.

À la lumière des informations obtenues de l’essai KN-F-4, la recirculation du rejet du nettoyeur, à la tête du circuit de flottation, n’est probablement pas une bonne idée. La teneur du rejet du premier nettoyeur est trop faible comparativement à celle de l’alimentation.

La recirculation du concentré du nettoyeur/épuiseur vers l’alimentation du premier nettoyeur n’est également pas une bonne alternative, à cause de sa teneur en or très élevée. Le concentré du nettoyeur/épuiseur gagnerait à combiner directement au concentré du nettoyage (de la même façon qu’effectué lors de l’essai KN-F-3).

En outre, cet essai comporte certaines anomalies au niveau des teneurs en arsenic, qui semblent sous-évaluées comparativement aux autres essais où l’arsenic a été analysé. En effet, la teneur calculée en arsenic de cet essai est de seulement 0,32 % As, comparativement aux autres essais réalisés dont la teneur de l’alimentation est plutôt de 0,5 % As.

La teneur calculée en or est également très élevée et se démarque de tous les autres essais effectués.


Page 18

3.3.4 Essai cyclique KN-F-6-R

L’essai cyclique KN-F-6-R a été réalisé en recirculant, cette fois, le rejet du nettoyeur vers le nettoyeur/épuiseur et en recirculant le concentré du nettoyeur/épuiseur vers l’alimentation du nettoyeur.

Essai: KN-F-6-RDate: 06-juil-14

Description:

4 kg

K80 = 102 µ

13,1 kPa

Table de Mozzley

0,69 g



Rejet final0,03% As0,62 g Au/mt


1,61% As15,9 g Au/mt

Reprise de l'essai KN-F-6 :Essai Knelson + flottations cycliques (Cycle 4)

(65,8% < 200 mailles)

Pompe Masterflex

Knelson

38114 g Au/tm

Concentré final12,09% As

96,1 g Au/tm

concentré


Concentré or libre


Rejet Knelson


250 g/tm CuSO4


13 g/tm MIBC





1,5min 2,0min


Figure 15 : Protocole expérimental de l’essai cyclique KN-F-6-R


Page 19

Tableau 9 : Bilan métallurgique de l’essai KN-F-6-R

Cycle 1 Cycle 2 Cycle 3 Cycle 4

0,17 g 0,17 g 0,17 g 0,17 g

38114 g Au/mt 38114 g Au/mt 38114 g Au/mt 38114 g Au/mt

31,60 g 33,30 g 40,40 g 39,10 g


13,96% As 13,4% As 12,5% As 12,1% As

891,80 g 951,20 g 947,70 g 951,00 g


0,02% As 0,04% As 0,03% As 0,03% As

923,40 g 984,50 g 988,10 g 990,10 g


0,50% As 0,49% As 0,54% As 0,51% As

Alim. calc.

Concentré or libre

Concentré final

Rejet 1 + Rejet 2


38,2% 32,4% 34,4%



34,4%

61,2%

95,6%

La récupération en or est établie à 94,6 % pour l’essai KN-F-6-R, ce qui est sensiblement la même que celle obtenue par l’essai KN-F-5. Cependant, les valeurs recueillies sont très cohérentes avec l’ensemble des essais réalisés, particulièrement au niveau de la teneur calculée en or et en arsenic de l’alimentation.

La teneur en or du concentré final se situe à 96,1 g Au/tm, avec une teneur en arsenic de 12,1 % As.

La concentration gravimétrique de l’or se situe à 60,2 %; valeur très similaire à la plupart des essais réalisés dans ce projet.


Page 20

3.3.5 Comparaison des résultats des essais KN-F-3, KN-F-5 et KN-F-6

Les essais KN-F-3, KN-F-5 et KN-F-6 représentent trois alternatives de schéma de traitement. Le tableau 10 résume les principaux résultats obtenus par ces trois essais.

Tableau 10 : Tableau comparatif des résultats des essais KN-F-3, KN-F-5 et KN-F-6-R

KN-F-3 KN-F-5 (cycle 4)

KN-F-6-R (cycle 4)

Knelson + flottation sans recirculation

Knelson + flottation cyclique avec

recirculation du rejet du 1er nett. au dégrossissage

Knelson + flottation cyclique avec

recirculation du rejet du 1er nett. au

nettoyeur-épuiseur

22947 g Au/mt 62143 g Au/mt 38114 g Au/mt

36,50 g 40,20 g 39,10 g


12,79% As 7,3% As 12,1% As

957,20 g 929,90 g 0,00 g


0,05% As 0,02% As 0,03% As

993,99 g 994,76 g 990,10 g


0,52% As 0,32% As 0,51% As



31,5% 26,2% 34,4%


Concentré or libre

Concentré final

Rejet 1 + Rejet 2

Alim. calc.

À priori, les résultats de ces trois essais sont relativement similaires au niveau de la teneur en or du concentré final.

Les essais KN-F-3 (essai en circuit ouvert) obtiennent une récupération similaire à l’essai KN-F-6-R, qui constitue un schéma comportant une recirculation en circuit fermé. Il est difficile de déterminer, à ce stade, lequel parmi ces deux essais obtient les meilleurs résultats métallurgiques, étant donné que les résultats obtenus sont très similaires. D’autres essais seraient nécessaires pour valider davantage le choix de la meilleure configuration du schéma de flottation.


Page 21

En dépit du fait que les résultats obtenus soient très bons, il y a toujours place à une certaine optimisation de ces résultats. Par exemple, le choix des réactifs utilisés n’a pas fait l’objet d’essais spécifiques.

Par ailleurs, il pourrait s’avérer judicieux d’évaluer la pertinence de rebroyer le rejet du nettoyeur et le concentré de l’épuiseur avant de réaliser l’étape du nettoyeur/épuiseur.

3.4 Essais combinant la concentration gravimétrique et la cyanuration

3.4.1 Essais KN-CN-F-4

La dernière série d’essais consistait à évaluer la récupération en combinant la concentration gravimétrique, suivie par une cyanuration. L’essai KN-CN-F-4 a été réalisé à une granulométrie de 102µ, sans étape de rebroyage. La cyanuration du rejet a été réalisée avec prélèvement de la solution à 25, 34 et 48 heures.

Essai: KN-CN-F-4Description:

K80 = 102 µ

13 kPa

Nettoyage sur table Mozley

concentré Rejet solide

Solution58,0%30508 g Au/tm 87,9%

Récupération gravimétrique de l'or suivie de la cyanuration du rejet gravimétrique

Pompe Masterflex

Knelson

d'or libre

Récupération globale en or =

concentré

Rejet de table de Mozley

% or libre récupéré

Broyage: 20 kg de tiges 7' 00"1,0 kg de minerais50% solide

Rejet Knelson

Cyanuration 48h

Figure 16 : Protocole de l’essai KN-CN-F-4


Page 22

Tableau 11 : Bilan métallurgique de l’essai KN-CN-F-4 avec K80 = 102µ à la cyanuration


0,210 0,02% 30508 g Au/tm 6,41 mg Au 58,0%

3,30 mg Au 29,9%

3,37 mg Au 30,5%

3,48 mg Au 31,6%

983,1 99,98% 1,17 g Au/tm 1,15 mg Au 10,4%

983,3 100% 11,23 g Au/tm 11,04 mg Au 100%

87,9%88,6%89,6%58,0%

Rejet solide de cyanuration


% or libre récupéré =Récupération globale en or après 48 h =

Consommation en cyanure de sodium = 0,33 kg NaCN / tmConsommation en chaux hydratée = 2,08 kg Ca(OH)2 / tm

Solution après 34 h de cyanuration


Récupération globale en or après 25 h =Récupération globale en or après 34 h =

Concentré or libre


On observe que la récupération augmente de 1,7 %, pour une durée de cyanuration de 48 heures, comparativement à une cyanuration de 25 heures.

Il faut noter, ici, que la cyanuration a été réalisée sur le rejet gravimétrique qui est relativement grossier (K80 = 102µ).

Le tableau 12 présente la récupération et la consommation en réactifs pour l’étape de la cyanuration, exclusivement.

Tableau 12 : Récupération et consommation en réactifs de la cyanuration uniquement (CN-F-4)

0,0 hres 0,0% 0,00 0,00

25,3 hres 71,2% 0,19 1,50

34,5 hres 72,7% 0,23 1,60

48,0 hres 75,2% 0,33 2,08

Durée Récupération Au

NaCN kg/tm de minerai

Ca(OH)2 kg/tm de minerai

On observe que les récupérations pour l’étape de cyanuration sont plus faibles mais cela est assez habituel étant donné que la majeure partie de l’or libre a déjà été récupérée lors de la concentration gravimétrique.

Les consommations en cyanures et en chaux hydratée se situent dans les moyennes observées pour un minerai d’or.


Page 23

3.4.2 Essai KN-CN-2

L’essai KN-CN-2 est semblable à l’essai précédent, à l’exception du rebroyage fin qui a été réalisé avant l’étape de la cyanuration.

Essai: KN-CN-F-2Description:

K80 = 102 µ

13 kPa

Nettoyage sur table Mozley K80 = 37 µ

concentré Rejet solide

Solution60,8%25598 g Au/tm 92,9%

Récupération gravimétrique de l'or suivie de la cyanuration du rejet gravimétrique avec broyage du rejet Knelson

Pompe Masterflex

Knelson

d'or libre

Récupération globale en or =

concentré

Rejet de table de Mozley

% or libre récupéré

Broyage: 20 kg de tiges 7' 00"1,0 kg de minerais50% solide

Rejet Knelson

Cyanuration 48 h

Figure 17 : Protocole expérimental de l’essai KN-CN-2


Page 24

Tableau 13 : Bilan métallurgique de l’essai KN-CN-2 avec K80 = 37µ à la cyanuration


0,280 0,03% 25598 g Au/tm 7,17 mg Au 60,8%

3,59 mg Au 30,4%

3,72 mg Au 31,5%

3,79 mg Au 32,1%

994,3 99,97% 0,84 g Au/tm 0,84 mg Au 7,1%

994,6 100% 11,86 g Au/tm 11,79 mg Au 100%

91,2%92,3%92,9%60,8%

Récupération globale en or après 48 h =

0,49 kg NaCN / tmConsommation en cyanure de sodium =Consommation en chaux hydratée = 3,19 kg Ca(OH)2 / tm

Concentré or libre

Solution après 25,6 h de cyanuration

Rejet solide de cyanuration


% or libre récupéré =

Solution après 38,3 h de cyanuration


Récupération globale en or après 25,6 h =Récupération globale en or après 38,3 h =

Tableau 14 : Récupération et consommation en réactifs de la cyanuration uniquement (KN-CN-F-2)

0,0 hres 0,0% 0,00 0,00

25,6 hres 77,6% 0,41 2,39

38,3 hres 80,4% 0,46 3,04

48,0 hres 81,9% 0,49 3,19

Durée Récupération Au

NaCN kg/tm de minerai

Ca(OH)2 kg/tm de minerai

On constate qu’un broyage fin augmente de façon significative la récupération globale qui atteint ici 92,9 %. Cependant, une granulométrie aussi fine (K80 = 37µ) implique des coûts de broyage importants.

La consommation en cyanure de sodium se situe, pour cet essai, à près de 0,5 kg de NaCN/tm, ce qui constitue une consommation plus élevée que la moyenne des minerais d’or.

On constate, encore ici, qu’une durée de cyanuration de 48 heures est justifiée pour maximiser la récupération.

4. Conclusions – caractérisation métallurgique

L’étude réalisée sur l’échantillon provenant de la zone 36 Est, de Ressources Radisson, a permis de définir un schéma de traitement permettant la concentration gravimétrique de l’or libre et la production d’un concentré à teneur en or assez élevée, par flottation.


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Les essais KN-F-3 et KN-F-6-R, réalisés avec un niveau de broyage relativement grossier (K80 = 102µ), ont obtenu les meilleurs résultats avec une récupération globale en or de 94,5 et 94,6 % respectivement. Les concentrés de flottation ont atteint, pour ces deux essais, des teneurs respectives de 91 et 96 g Au/tm. La concentration de l’or libre, réalisée à l’aide du concentreur Knelson et de la table de Mozley, ont atteint respectivement 62,9 et 60,2 % pour ces deux essais.

Il est difficile, à ce stade, de définir la meilleure configuration du schéma de traitement, puisque l’essai KN-F-3 (circuit ouvert) a obtenu des résultats similaires à l’essai KN-F-6-R (essai cyclique réalisé en circuit fermé avec recirculations). D’autres essais seraient requis pour préciser le choix définitif.

De façon générale, les récupérations gravimétriques de l’or libre ont été assez semblables (se situant à environ 60 %), ce qui est fort intéressant. Cette valeur obtenue en laboratoire pourrait être légèrement inférieure en usine selon la configuration des équipements utilisés.

Les essais de récupération gravimétrique de l’or libre ont montré qu’une granulométrie assez grossière (K80 = 102µ) était suffisante pour récupérer l’or libre présent dans ce minerai.

La granulométrie optimale pour la flottation se situe entre 102 et 75µ. D’autres essais seraient requis pour préciser davantage cette granulométrie.

Certaines variations importantes de la teneur en arsenic, dans le concentré de flottation, ont été observées aux cours des essais. Cependant, selon notre interprétation, il faut considérer que la teneur en arsenic, dans le concentré de flottation sera, de l’ordre de 12 % As, ce qui constitue une valeur assez élevée pour un traitement à la fonderie. La pénalité attribuée au traitement de ce concentré à la fonderie pourrait être importante.

Les essais de flottation ont été réalisés en utilisant les mêmes réactifs (KAX et A-407). Des gains de récupération pourraient être possibles en réalisant des essais spécifiquement dédiés à la sélection des meilleurs réactifs de flottation pour ce minerai.

Les deux essais combinant la récupération gravimétrique de l’or libre et la cyanuration ont respectivement obtenus des récupérations globales de 89,6 (K80 =102µ) et 92,9 % (K80 =37µ). On constate ainsi qu’une granulométrie très fine augmente la récupération de façon significative.

Les essais de cyanuration montrent également qu’un temps de séjour élevé à la cyanuration est souhaitable pour maximiser la récupération en or. Aucun essai n’a été réalisé avec une durée de cyanuration supérieure à 48 heures. Ainsi, d’autres essais de laboratoire pour-raient être fait dans un futur projet afin d’évaluer les durées de cyanuration supérieures (p. ex. : 72 et 96 heures).


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Partie 2 : Caractérisation environnementale

Cette section du rapport fait état des caractérisations environnementales qui ont été réalisées sur quatre composites représentant le minerai des différentes épontes, ainsi qu’un rejet de flottation type produit au cours des travaux de minéralurgie du présent projet. L’approche utilisée pour l’évaluation environnementale est de comparer les résultats des différentes caractérisations aux définitions de résidus miniers contenues à l’annexe 2 de la Directive 019 (2012).

1. Échantillons

Les échantillons ont été reçus à l’URSTM-UQAT sous forme de sacs contenant de la roche concassée (passant ~ 1 cm pour certains et passant ~ 1-2 mm pour d’autres). La liste complète des échantillons reçus est présentée au tableau 15. Selon les indications du client, ces échantillons ont intégralement été mélangés selon les assemblages du tableau 15, afin de former les quatre composites pour l’étude environnementale. Ces derniers ont été assemblés par le client de manière à représenter différentes épontes dans la mine. Les composites ont été mélangés par roulage (40 fois par coin) et divisés ensuite par séparateurs à riffles, de manière à produire un sous-échantillonnage représentatif qui a été pulvérisé passant 200 µm pour les analyses chimiques et les essais statiques. Le tableau 16 montre les différentes moutures des composites utilisées dans l’étude environnementale. En plus des composites représentant les épontes, deux rejets de traitement métallurgiques, produits par Jean Lelièvre suite aux travaux de la partie 1 de ce rapport, ont aussi été caractérisés dans cette étude. Le premier est l’échantillon « Rejet de flottation », qui provient de l’essai sur concentrateur gravimétrique Knelson et une flotte subséquente de son rejet (ce rejet a été produit par l'essai KN-F-5 de l’étude métallurgique). Le deuxième rejet testé (seulement pour le bilan des essais statiques), nommé S21-S23, est un rejet de concentration Knelson suivie d’une cyanuration (essai métallurgique KN-CN-2). Ces deux échantillons de rejets ont été caractérisés à la granulométrie telle que reçue.


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Tableau 15 : Liste des échantillons bruts constituant les quatre composites de l’étude environnementale

Nom échantillon brut #URSTM Remarques

Volc Centre = #3 (45236) 32145Concassé passant ~ 1 cm

Masse approx: 2.85 kg

Volc Centre = #3 (45237) 32146Concassé passant ~ 1 cm


Volc Sud Éponte #1A (49918) 32147Concassé passant ~ 2 mm


Volc Sud #1B (49934) 32148Concassé passant ~ 2 mm

Masse approx: 2 kg

Volc Sud #1B (49935) 32149Concassé passant ~ 2 mm

Masse approx: 2,07 kg

Porph Nord (45221) 32150Concassé passant ~ 1 cm






Porph Sud (45248) 32153Concassé passant ~ 1 cm


Porph Sud (45249) 32154Concassé passant ~ 1 cm


WCgl (45228) 32155Concassé passant ~ 1 cm








W-Volc-Sud N (45204) 32159Concassé passant ~ 1 cm

Masse approx: 1 kg











Co

mp

osi

te É

po

nte

s 1

+ 3

Co

mp

osi

te É

po

nte

s 2

+ 5

Co

mp

osi

te É

po

nte

4C

om

po

site

Ép

on

te 6


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Tableau 16 : Différentes moutures des échantillons pour les essais environnementaux

Nom #URSTM Mouture Essais

Composite Épontes 1 + 3 grossier 32165 < 1 cm TCLP

Composite Épontes 2 + 5 grossier 32166 < 1 cm TCLP

Composite Éponte 4 grossier 32167 < 1 cm TCLP

Composite Éponte 6 grossier 32168 < 1 cm TCLP

Composite Épontes 1 + 3 pulv 32169 < 200 µmAnalyses chimiques, Essais

statiques

Composite Épontes 2 + 5 pulv 32170 < 200 µmAnalyses chimiques, Essais

statiques

Composite Éponte 4 pulv 32171 < 200 µmAnalyses chimiques, Essais

statiques

Composite Éponte 6 pulv 32172 < 200 µmAnalyses chimiques, Essais

statiques

Rejet de flottation 34290 < 200 µmTCLP, Analyses chimiques,

Essais statiques

2. Méthodes

2.1 Caractérisations chimiques des solides

Pour évaluer les éléments majeurs, les échantillons ont été envoyés dans un laboratoire sous-traitant pour une analyse « roche totale » en fluorescence des rayons-X. Les échantillons ont aussi été soumis à une analyse chimique par ICP-AES, suite à une digestion complète par HNO3/Br2/HF/HCl aux laboratoires de l’URSTM-UQAT. Cette analyse comprend les éléments suivants : Al, Ba, Ca, Cd, Co, Cr, Cu, Fe, Mg, Mn, Mo, Ni, Pb, Sn, Ti et Zn. L’analyse de Ag, B, K, Hg et Na a été réalisée par un laboratoire sous-traitant, à partir de la solution provenant de cette digestion totale. L’analyse des métaux lourds (As, Be, Bi, Sb, Se, Te) par ICP-AES, suite à une digestion acide adaptée, a aussi été réalisée. L’analyse du Stotal et du Ctotal a été exécuté par fournaise à induction. Le texte de l’annexe 2 de la Directive 019 contient la définition suivante :

« Résidus miniers à faibles risques Il s’agit de résidus miniers dont les concentrations en métaux n’excèdent pas les critères de niveau A indiqués au tableau 1 de l’annexe 2 de la Politique de protection des sols et de réhabilitation des terrains contaminés. Ces critères représentent les teneurs de fond qui prévalent pour la province géologique des Basses-Terres du Saint-Laurent. Pour les autres provinces géologiques, les teneurs de fond sont présentées au tableau 2 de cette même annexe. […]


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Les résidus qui lixivient en deçà des critères établis pour désigner des résidus miniers lixiviables sont également considérés comme des résidus miniers à faibles risques. »

Les concentrations en métaux obtenues ici sont donc comparées aux critères de la Politique de protection des sols et de réhabilitation des terrains contaminés (PPSRTC) pour la province géologique du Supérieur, afin de valider si les échantillons étudiés peuvent être considérés comme « résidus miniers à faibles risques ».

2.2 Essais PGA

Les essais statiques de détermination du potentiel de génération d’acide (PGA) ont été réalisés. Ces essais dressent le bilan entre le potentiel de génération d’acidité (PA) d’un matériel, qui est relié aux minéraux sulfureux, et son potentiel de neutralisation de cette acidité (PN), qui est relié aux minéraux carbonatés et à certains silicates. Ces essais incluent :

la détermination du soufre total (Stotal) par fournaise à induction;

la détermination du soufre sous forme sulfate (Ssulfate) par lixiviation acide et lecture ICP-AES;

l’analyse du carbone total (Ctotal) par fournaise à induction;

le bilan du carbone par lessivage acide des carbonates et lecture du carbone résiduel à la fournaise à induction (Ccarbonates = Ctotal - Crésiduel; ici le Crésiduel peut être attribué au Cgraphite puisque les échantillons sont des roches fraiches);

le calcul du potentiel de génération d’acidité (PA = 31,25x%Ssulfures, où %Ssulfures = %Stotal - %Ssulfates);

la détermination du potentiel de neutralisation (PN) par la méthode de Sobek (1978) modifiée par Lawrence et Wang (1996);

le calcul du potentiel net de neutralisation (PNN = PN - PA) et le calcul du ratio PN/PA. Les critères contenus dans la Directive 019 sont ensuite utilisés pour interpréter les données du PNN et du PN/PA. Le texte de l’annexe 2 de la Directive 019 (2012) se lit comme suit :

« Résidus miniers acidogènes Il s’agit de résidus miniers contenant du soufre (Stotal) en quantité supérieure à 0,3 % et dont le potentiel de génération acide a été confirmé par des essais de prévision statiques, en répondant à au moins une des deux conditions suivantes :

le potentiel net de neutralisation (PNN) d’acide est inférieur à 20 kg CaCO3/tonne de résidus;

le rapport du potentiel de neutralisation d’acide sur le potentiel de génération d’acide (PN/PA) est inférieur à 3.

Des essais de prévision cinétiques peuvent aussi être réalisés pour confirmer ou infirmer le caractère acidogène obtenu à la suite des résultats des essais de prévision statiques qui ont été réalisés. »

2.3 Essais de lixiviation

Le texte de l’annexe 2 de la Directive 019 (2012) se lit :


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« Résidus miniers lixiviables Il s’agit de résidus miniers qui, lorsqu’ils sont mis à l’essai conformément à la méthode d’analyse de lixiviation MA.100-Lix.com.1.1 (TCLP), produisent un lixiviat contenant un contaminant dont la concentration est supérieure aux critères applicables pour la protection des eaux souterraines, sans toutefois produire un lixiviat contenant un contaminant dont la concentration est supérieure aux critères énoncés dans le tableau 1 ci-dessous. Les critères de référence définis en fonction des récepteurs sont présentés à l’annexe 2 de la Politique de protection des sols et de réhabilitation des terrains contaminés. Soulignons que la liste des critères présentés à l’annexe 2 de cette politique

n’est pas limitative.» Les critères de la Politique de protection des sols et de réhabilitation des terrains contaminés (PPSRTC) cités sont :

l’eau souterraine aux fins de consommation (résumé ESFC);

les résurgences dans les eaux de surface (résumé RESIE). Toujours à l’annexe 2 de la Directive 019 (2012), on peut lire la définition suivante :

« Résidus miniers à risques élevés Il s’agit de résidus miniers […] qui, lorsqu’ils sont mis à l’essai conformément à la méthode d’analyse de lixiviation MA.100-Lix.com.1.1 (TCLP), produisent un lixiviat contenant un contaminant dont la concentration est supérieure aux critères énoncés dans le tableau 1 ci-dessous ». Les valeurs de ce tableau 1 sont donc aussi utilisées dans les interprétations des résultats des essais TCLP de l’étude.

Ces lixiviations TCLP de la méthode MA.100-Lix.com du CEAEQ ont été réalisées conformément au texte de la Directive 019. Suite aux lixiviations, les solutions ont été analysées pour en déterminer le pH final et les concentrations en métaux et ions (Ag, Al, Sb, As, Ba, B, Be, Bi, Cd, Ca, Cr, Co, Cu, Fe, Hg, Pb, Mg, Mn, Mo, Ni, K, Se, Si, Ti, Zn, Sulfates, Stotal, Na). La méthode MA.100-Lix.com du CEAEQ fixe les granulométries maximales à utiliser pour les essais de lixiviations. Pour le TCLP, nous mentionnons une granulométrie passant 9,5 mm. Les composites des épontes ont donc été analysés tels que reçus (tableau 16) car leur distribution granulométrique était adéquate. L’échantillon de rejet de flottation a aussi été traité tel quel pour l’extraction TCLP.

Bien que les essais MA.100-Lix.com.1.1 (TCLP) soient la norme au point de vue environ-nemental, ces essais ne doivent être utilisés seulement que pour identifier des éventuels problèmes de mobilité de métaux à partir de rejets miniers. Ils ne représentent en aucun cas les qualités d’eau réelles auxquels on pourrait s’attendre dans les conditions de terrain. Il ne s’agit que d’essais diagnostiques et ne représentent pas des conditions naturelles d’altération des matériaux (Kandji, 2014).


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3. Résultats

3.1 Caractérisations chimiques des solides

Le tableau 17 présente l’ensemble des résultats des analyses chimiques réalisées sur les échan-tillons. Selon les analyses chimiques, les échantillons semblent principalement composés d’alumino-silicates de calcium. Ces derniers semblent aussi comporter une certaine concentration de carbonates, notable par la présence de magnésium (qui pourrait aussi être inclus dans les aluminosilicates) et des pertes au feu entre 3 et 8 %. La présence de carbonates est d’ailleurs confirmée selon le bilan du carbone (tableau 5). Les échantillons semblent en général peu sulfureux (Stotal entre 0,36 et 0,44 % en ICP-AES). Concernant les éléments possiblement problématiques pour l’environnement, les présences d’As (entre 88,9 et 1 690 mg As/kg), de Ba (entre 218 et 753 mg Ba/kg), de Co (entre 27 et 74,3 mg Co/kg), de Cr (entre 187 et 497 mg Cr/kg), de Cu (entre 32,9 et 104 mg Cu/kg), de Mn (entre 496 et 1274 mg Mn/kg), de Mo (entre 38,4 et 58,7 mg Mo/kg) et de Ni (entre 82,9 et 204 mg Ni/kg), ont été détectées. Notons que la flottation ne semble pas retirer une portion importante de ces éléments, car les concentrations au rejet de flottation sont très semblables à celles des composites des quatre épontes. Le tableau 18 compare les concentrations élémentaires obtenues aux critères de la PPSRTC. On observe plusieurs dépassements des critères A, B et C de la Politique pour tous les matériaux testés. Parmi les dépassements aux critères, notons :

As, qui dépasse le critère C pour les quatre composites et les rejets de flottation;

Ba, qui dépasse le critère B dans trois des cinq échantillons et le critère A pour les autres;

Cr, qui dépasse le critère B dans 3/5 échantillons et le critère A pour les autres;

Cu, qui dépasse le critère B pour le composite Éponte 1+3 et le critère A pour les échantillons Éponte 4, Éponte 6 et le rejet de flottation;

Mn, qui dépasse les critères A et B pour les Éponte 1+3 et 6;

Mo, qui dépasse le critère C dans tous les échantillons, sauf le composite Éponte 2+5 où il dépasse tout de même le critère B;

Ni, qui dépasse le critère B pour 3/5 échantillon et le critère A pour les deux autres. Nous devons cependant attendre de voir si ces éléments sont mobiles sous les conditions de l’essai MA.100-Lix.com (TCLP) avant de de ne pouvoir statuer si les matériaux satisfont quand même la définition de « Résidus miniers à faible risques » de la Directive 019.


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Tableau 17 : Analyses chimiques réalisées sur les échantillons de l’étude environnementale

Paramètre Unités LDM Composite

Éponte 1+3

pulv.

Composite

Éponte 2+5

pulv.

Composite

Éponte 4 pulv.

Composite

Éponte 6 pulv.

Rejet de

flottation

U32169 U32170 U32171 U32172 U34290

Fe2O3 % p/p 0,01 11,20 5,12 8,93 8,90 5,90

SiO2 % p/p 0,01 48,92 62,26 59,55 49,79 57,46

Al2O3 % p/p 0,01 13,51 14,06 14,60 13,23 13,40

Na2O % p/p 0,01 2,37 4,56 2,82 1,69 3,05

MgO % p/p 0,01 6,27 3,29 2,88 5,54 4,04

K2O % p/p 0,01 1,77 1,74 2,60 1,53 2,33

CaO % p/p 0,01 8,29 3,86 3,35 8,29 5,16

P2O5 % p/p 0,01 0,14 0,16 0,16 0,13 0,14

MnO % p/p 0,01 0,20 0,08 0,16 0,17 0,12

TiO2 % p/p 0,01 0,85 0,35 0,61 0,74 0,49

Cr2O3 % p/p 0,01 0,02 0,04 0,03 0,02 0,05

V2O5 % p/p 0,01 0,04 0,01 0,02 0,03 0,02

LOI % p/p 0,01 6,78 4,31 3,00 8,56 5,59

Mass Balance % p/p 0,01 100,35 99,84 98,70 98,62 97,74

Ag 1 mg/kg 2 <2 <2 <2 <2 <2

Al mg/kg 60 66130 47300 62380 62760 68940

As mg/kg 30 88,9 1690 382 257 251

B 1 mg/kg 0,01 10,0 10,5 7,2 3,0 <0,01

Ba mg/kg 5 266 753 542 218 588

Be mg/kg 1 <1 <1 <1 <1 <1

Bi mg/kg 30 <30 <30 <30 <30 <30

Ca mg/kg 60 60960 25360 23830 59550 41430

Cd mg/kg 5 <5 <5 <5 <5 <5

Co mg/kg 5 74,3 28,7 52,3 61,7 27

Cr mg/kg 5 187 497 401 217 301

Cu mg/kg 10 104 32,9 77,2 93,8 72

Fe mg/kg 10 78660 27080 54590 60780 39220

Hg 1 mg/kg 0,01 <0,01 0,10 <0,01 <0,01 0,02

Mg mg/kg 15 32820 8991 11960 27810 20420

Mn mg/kg 5 1274 496 908 1026 775

Mo mg/kg 5 58,7 38,4 50,2 53,3 48

Ni mg/kg 5 82,9 204 166 86,1 120

Pb mg/kg 5 <5 <5 <5 <5 <5

Sb mg/kg 4 <4 30,8 10 <4 <4

Se mg/kg 3 <3 <3 <3 <3 <3

Sn mg/kg 5 <5 <5 <5 <5 <5

Stotal (ICP) mg/kg 200 4389 3815 5563 4977 3558

Te mg/kg 2 <2 <2 <2 <2 <2

Ti mg/kg 25 4677 1227 3174 3026 9359

Zn mg/kg 55 <55 <55 <55 <55 <55

1 : la méthode de digestion pour ces éléments doit être cons idérée partiel le, mais conforme à MA. 200 – Mét. 1.2

n/a : non analysé

Ro

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Tableau 18 : Comparaison des résultats des analyses chimiques avec les critères de la PPSRTC

Éléments Unités Critère A* Critère B* Critère C*

Éponte 1+3 Éponte 2+5 Éponte 4 Éponte 6 Rejet

fottation

U32169 U32170 U32171 U32172 U34290

Argent (Ag) mg/kg <2 <2 <2 <2 <2 0,5 20 40

Arsenic (As) mg/kg 88,9 1690 382 257 251 5 30 50

Baryum (Ba) mg/kg 266 753 542 218 588 200 500 2000

Cadmium (Cd) mg/kg <5 <5 <5 <5 <5 0,9 5 20

Chrome (Cr) mg/kg 187 497 401 217 301 85 250 800

Cobalt (Co) mg/kg 74,3 28,7 52,3 61,7 27 20 50 300

Cuivre (Cu) mg/kg 104 32,9 77,2 93,8 72 50 100 500

Étain (Sn) mg/kg <5 <5 <5 <5 <5 5 50 300

Manganèse (Mn) mg/kg 1274 496 908 1026 775 1000 1000 2200

Mercure (Hg) mg/kg <0,01 0,1 <0,01 <0,01 0,02 0,3 2 10

Molybdène (Mo) mg/kg 58,7 38,4 50,2 53,3 48 6 10 40

Nickel (Ni) mg/kg 82,9 204 166 86,1 120 50 100 500

Plomb (Pb) mg/kg <5 <5 <5 <5 <5 40 500 1000

Selenium (Se) mg/kg <3 <3 <3 <3 <3 3 3 10

Zinc (Zn) mg/kg <55 <55 <55 <55 <55 120 500 1500

*Cri tères de la Pol i tique de protection des sols et de réhabi l i tation des terra ins contaminés pour la province du supérieur

Échantillon

3.2 Essais statiques de détermination du PGA

Les résultats du bilan complet des essais statiques de détermination du potentiel de génération d’acide (PGA) sont présentés au tableau 19. On observe que la majorité du carbone contenu dans les composites des épontes et le rejet de flottation est lié à la présence de carbonates, tel que suspecté par la perte au feu plus haut. Seul le composite Éponte 6 contient des traces (0,20 % C) de carbone résiduel. Ce carbone résiduel peut être attribué au graphite vu que la roche est fraîche et ne contient visuellement pas de carbone issu de décompositions organiques. L’analyse des sulfates a été peu probante ici, car l’extraction acide (40 % v/v HCl) a attaqué les sulfures, phénomène confirmé par l’odeur de H2S lors de l’essai. Ceci arrive lorsque des sulfures plus réactifs, tels la pyrrhotite ou la sphalérite sont présents. C’est le soufre total qui a donc été utilisé pour calculer le potentiel maximal de production d’acide (PAM, plutôt que l’habituel PA calculé à partir du Ssulfure = Stotal – Ssulfate). Néanmoins, on observe que les échantillons sont peu sulfureux, avec des teneurs en soufre total variant entre 0,332 et 0,557 % Stotal, ce qui a mené à des PAM se situant entre 8,6 kg CaCO3/t (pour le rejet de flottation) et 30,5 kg CaCO3/t (pour le rejet de cyanuration). La présence de carbonates a conféré aux matériaux testés des potentiels de neutralisation (PN) se situant entre 44,9 kg CaCO3/t (pour le composite Éponte 4) et 150 kg CaCO3/t (valeur limite de l’essai modifié de Lawrence et Wang [1996], pour le composite Éponte 6).


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La présence de PN et les faibles concentrations en soufre ont mené à des verdicts « non acidogènes » pour tous les matériaux, sauf pour le composite Éponte 4 qui, bien qu’il ne possède qu’un PNN de 28 kg CaCO3/t, il affiche un rapport PN/PA de 2,6 (la Directive 019 exige un rapport > 3). Seul le composite Éponte 4 doit donc être considéré « résidus miniers acidogènes » selon les définitions de la Directive 019 (2012).

Tableau 19 : Bilan des essais statiques de détermination du PGA

Paramètre Unités LDM Composite

Éponte 1+3

pulv.

Composite

Éponte 2+5

pulv.

Composite

Éponte 4 pulv.

Composite

Éponte 6 pulv.

Rejet de

flottation

Rejet

cyanuration

(S-21 S-24 )

U32169 U32170 U32171 U32172 U34290 U34625

Ctotal % p/p 0,05 1,92 1,08 0,60 2,40 1,40 1,36

Cgraphite % p/p 0,04 <0,04 <0,04 0,04 0,20 <0,04 n/d

Ccarbonates % p/p 0,04 1,92 1,08 0,56 2,20 1,40 n/d

Stotal (Leco) % p/p 0,009 0,332 0,438 0,557 0,441 0,379 1,12

Ssulfates 1

% p/p n/d 0,141 0,298 0,429 0,176 0,103 0,147

Ssulfures % p/p n/d 0,191 0,140 0,128 0,265 0,276 0,98

PAM 1 kg CaCO3/t n/d 10,4 13,7 17,4 13,8 8,6 30,5

PN 2 kg CaCO3/t n/d 115 74,9 44,9 150 94,3 112

PNN kg CaCO3/t n/d 105 61 28 136 85,7 81,5

PN/PA - - 11,1 5,5 2,6 10,9 10,9 3,7

Acidogène 3 - - Non Non Oui Non Non Non

1 : la matrice des sul fures a été touchée par l 'extraction des sul fates , le PAM est donc ca lculé à parti r du S total

3 : selon les cri tères de la Directive 019, mars 2012

n/d : non déterminé

2 : La l imite de l 'essa i de Lawrence et Wang 1997 est de 150 kgCaCO3/t, i l se pourra i t que le PN véri table soi t

plus élevé que la va leur rapportée dans ce cas .

3.3 Essais de lixiviations

Le tableau 20 présente les résultats complets des analyses des solutions post-lixiviations MA.100-Lix.com.1.1 (TCLP) réalisées sur tous les échantillons de l’étude. Les résultats d’analyses y sont comparés :

Aux critères d’eau souterraine aux fins de consommation (résumé ESFC);

Aux critères de résurgences dans les eaux de surface (résumé RESIE); Un dépassement d’un des critères parmi ces deux séries qualifie l’échantillon de

« résidus miniers lixiviables »;

Aux critères du tableau 1 de l’annexe 2 de la Directive 019 (2012); Un dépassement d’un de ces critères qualifie l’échantillon de « résidus miniers à risques

élevés ».


Page 35

Mise en garde : Une contamination en antimoine (Sb) a été observée dans les blancs des solutions TCLP (tableau 20). Les dépassements des critères ESFC rencontrés ici sont donc des faux positifs, d’autant plus que les concentrations en Sb étaient très faibles (sous ou près de la limite de détection) dans les échantillons solides.

La même conclusion ne peut cependant pas être tirée pour la présence de zinc (Zn). Bien que présent dans les blancs (~0,06 mg Zn/l), les concentrations en Zn ont au moins doublé lors des essais de lixiviations, donc proviennent bien du lessivage des échantillons testés. En observant les données du tableau 20, on constate que :

Ag : des dépassements du critère de RESIE (0,00062 mg Ag/l) pour les matériaux Éponte 2+5 et Éponte 6;

As : un dépassement du critère d’ESFC pour l’As (0,025 mg As/l) pour le matériel Éponte 4 (avec 0,083 mg As/l);

Ba : un dépassement du critère d’ESFC pour le Ba (1 mg Ba/l), à la suite de la lixiviation du matériel Éponte 2+5 (avec 1,09 mg Ba/l);

Cu : seul le rejet de flottation produit un dépassement au critère de RESIE (0,0073 mg Cu/l) pour le cuivre (avec 0,53 mg Cu/l);

Mn : tous les échantillons ont produit un dépassement du critère esthétique pour le Mn de la ESFC (0,05 mg Mn/l) avec des concentrations entre 8,23 et 21,3 mg Mn/l dans les lixiviats;

Hg : deux dépassements du critère de RESIE (0,00013 mg Hg/l) sont observés pour les composites Éponte 1+3 et Éponte 2+5;

Ni : le composite Éponte 6 et le rejet de flottation ont produit des dépassements du critère d’ESFC (0,02 mg Ni/l);

Zn : tous les matériaux lixiviés (même si on retire l’apport de la solution de lixiviation) ont produit des dépassements au critère de RESIE pour le Zn (0,0167 mg/l), avec des concentrations corrigées entre 0,057 mg Zn/l (Éponte 1+3) et 0,209 mg Zn/l (rejet de flottation).

À la lumière de ces résultats et en vue des définitions de l’annexe 2 de la Directive 019 (2012) :

Tous les matériaux testés se qualifient de « résidus miniers lixiviables »;

Aucun matériel testé ne se qualifie en tant que « Résidus miniers à risques élevés ».


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Tableau 20 : Résultats des lixiviations MA.100-Lix.com.1.1 (TCLP) réalisées sur tous les matériaux de l’étude environnementale

Paramètre Symbole Unités LDR

- - U32165 U32166 U32167 U32168 U34290 U34290 (d)

Solution # - - - 1 2 2 1 1 1 2 2 s/o s/o s/o

pH initial pH - - 4,88 2,90 2,90 4,88 4,88 4,88 2,90 2,90 s/o s/o s/o

pH final pH - - 4,9 2,89 5,64 6,15 5,55 6,34 5,18 5,18 s/o s/o s/o

Aluminium (Al) mg/l 0,01 0,099 0,051 <0,01 0,123 0,338 0,102 0,580 0,588 s/o 0,75 s/o

Antimoine1 (Sb) mg/l 0,09 0,752 0,647 0,167 0,399 0,550 0,555 0,697 <0,09 0,006 s/o s/o

Argent (Ag) mg/l 0,0001 <0,0001 <0,0001 <0,0001 0,0017 <0,0001 0,0020 <0,0001 <0,0001 0,1 0,00062 s/o

Arsenic (As) mg/l 0,06 <0,06 <0,06 <0,06 <0,06 0,083 <0,06 <0,06 <0,06 0,025 0,34 5,0

Baryum (Ba) mg/l 0,001 0,072 0,074 0,017 1,09 0,773 0,284 0,575 0,628 1 5,3 100

Béryllium (Be) mg/l 0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001 s/o s/o s/o

Bismuth (Bi) mg/l 0,02 <0,02 <0,02 <0,02 <0,02 <0,02 <0,02 <0,02 <0,02 s/o s/o s/o

Bore (B) mg/l 0,01 <0,01 <0,01 <0,01 0,200 <0,01 0,040 <0,01 <0,01 s/o s/o 500

Cadmium (Cd) mg/l 0,003 <0,003 <0,003 <0,003 <0,003 <0,003 <0,003 <0,003 <0,003 0,005 0,00021 0,5

Calcium (Ca) mg/l 0,03 0,783 0,126 1850 729 624 956 1540 1550 s/o s/o s/o

Chrome (Cr) mg/l 0,003 <0,003 <0,003 <0,003 <0,003 0,013 <0,003 0,005 0,006 0,05 0,2 5,0

Cobalt (Co) mg/l 0,004 <0,004 <0,004 <0,004 <0,004 0,005 0,006 0,009 0,006 s/o 0,5 s/o

Cuivre (Cu) mg/l 0,003 0,007 0,009 <0,003 0,008 0,004 0,003 0,531 0,536 1* 0,0073 s/o

Fer (Fe) mg/l 0,006 0,081 0,060 17,8 5,17 8,52 0,217 13,4 12,6 s/o s/o s/o

Magnésium (Mg) mg/l 0,001 0,017 0,013 12,9 7,54 3,66 4,20 33,2 33,5 s/o s/o s/o

Manganèse (Mn) mg/l 0,002 0,003 <0,002 21,3 11,6 8,41 8,23 20,5 21,0 0,05* s/o s/o

Mercure (Hg) mg/l 0,00001 0,00007 <0,00001 0,00057 0,00033 0,00009 0,00008 <0,00001 <0,00001 0,001 0,00013 0,1

Molybdène (Mo) mg/l 0,009 0,012 0,011 <0,009 <0,009 <0,009 <0,009 <0,009 <0,009 0,07 2 s/o

Nickel (Ni) mg/l 0,004 <0,004 <0,004 <0,004 0,012 0,017 0,028 0,033 0,033 0,020 0,26 s/o

Plomb (Pb) mg/l 0,02 <0,02 <0,02 <0,02 <0,02 <0,02 <0,02 <0,02 <0,02 0,01 0,034 5,0

Potassium (K) mg/l 0,05 0,51 <0,05 54,4 21,7 28,7 7,70 25,8 26,1 s/o s/o s/o

Sélénium (Se) mg/l 0,1 <0,1 <0,1 <0,1 <0,1 <0,1 <0,1 <0,1 <0,1 0,01 0,02 1,0

Sodium2 (Na) mg/l 0,05 1280 0,430 1,70 1502 1414 1516 1,40 1,80 200 s/o s/o

Titane (Ti) mg/l 0,002 <0,002 <0,002 <0,002 <0,002 <0,002 <0,002 <0,002 <0,002 s/o s/o s/o

Zinc (Zn) mg/l 0,005 0,057 0,062 0,119 0,132 0,167 0,148 0,271 0,290 5* 0,0167 s/o

Sulfates (SO42-) mg/l 0,6 22,8 14,8 14,5 10,4 <0,6 7,90 14,0 15,2 s/o s/o s/o

s/o : sans objet dans la PPSRTC et/ou la Directive 019

*: Des objecti fs d’ordre esthétiques sont disponibles pour certa ins paramètres .1: Une contamination en Antimoine dans les solutions d'extraction nous empêche d'uti l i ser les résultats des analyses .

2: Le sodium entre dans la compos ition de la solution #1 pour les extractions TCLP.

Composite

Éponte 1+3

grossier

Composite

Éponte 6

grossier

Rejets

Flottation

(double)

Blanc

solution 1

Blanc

solution 2

Composite

Éponte 2+5

grossier

Composite

Éponte 4

grossier

Rejets

Flottation

Critères d'eau

souterraine

aux fins de

consommation

Critères de

résurgences

dans les eaux

de surface

Résidus

miniers à

risques élevés

(Directive 019)


Page 37

4. Conclusions – caractérisation environnementale

Les conclusions suivantes ne s’appliquent qu’aux échantillons reçus pour analyses et qu’aux conditions d’analyses utilisées dans l’étude. L’étude environnementale nous apporte les principales conclusions suivantes :

Les matériaux étudiés sont principalement composés de silicates.

Les matériaux contiennent des métaux potentiellement problématiques au niveau de l’environnement à l’état de traces mais non négligeables (dépassements de certains critères PPSRTC), soient : As, Ba, Cr, Cu, Mn, Mo, Ni.

Les matériaux sont peu sulfureux Stotal ≤ 0,557 % Stotal et ont donc des potentiels de génération d’acide faibles (PAM ≤ 30,5 kg CaCO3/t).

Les matériaux contiennent un potentiel de neutralisation (PN) provenant des carbonates (PN ≥ 44,9 kg CaCO3/t).

Seul le matériel composite Éponte 4 est considéré acidogène, car son rapport PN/PA est < 3 (PN/PA = 2,6).

Les essais de lixiviation MA.100-Lix.com.1.1 (TCLP) montrent que : Les rinçages dépassent les critères de résurgence (RESIE) ou de l’eau souterraine aux fins

de consommation (ESFC) de la Politique de protection des sols et réhabilitation des terrains contaminés (PPSRTC), notamment pour Ag (2 matériaux sur 5), As (1/5), Ba (1/5), Cu (1/5), Mn (5/5), Hg (2/5), Ni (2/5) et Zn (5/5) ;

Tous les matériaux testés doivent être considérés « Résidus miniers lixiviables »; Aucun des matériaux testés n’est considéré « résidus miniers à risques élevés ».

5. Recommandations

Nous avons vu plus haut qu’un des cinq matériaux testés, soit le composite Éponte 4, retournait un rapport PN/PA de 2,6 et devait donc, selon la Directive 019, être considéré acidogène (même si son PNN est de 28 kg CaCO3/t). Il est recommandé de tester ce matériel à l’aide d’un essai cinétique en cellule d’humidité ou en colonne afin d’avoir l’heure juste sur son caractère acidogène. Tel que mentionné précédemment, les conditions de l’essai standardisé MA.100-Lix.com.1.1 (TCLP) ne représentent pas des conditions naturelles d’altération des matériaux, mais servent généralement plus d’essais diagnostiques permettant une première évaluation de la mobilité des contaminants. Encore une fois, des essais cinétiques se rapprochant des conditions naturelles d’altération des matériaux seraient à préconiser si le client veut évaluer la mobilité plus naturelle des éléments Ag, As, Ba, Cu, Mn, Hg, Ni et Zn.


Page 38

6. Références

CENTRE D’EXPERTISE EN ANALYSE ENVIRONNEMENTALE DU QUÉBEC. Détermination des métaux : méthode par spectrométrie de masse à source ionisante au plasma d’argon. MA. 200 – Mét 1.2, Rév. 4, Ministère du Développement durable, de l’Environnement, de la Faune et des Parcs du Québec, 2013, 34 p.

Directive 019 sur l’industrie minière, Gouvernement du Québec, ministère du Développement durable, Environnement et Parcs, Mars 2012.

KANDJI E.B. (2014) Essais de lixiviation conçus pour les rejets industriels et municipaux en général : application au contexte minier. Rapport de synthèse environnemental présenté comme exigence partielle au doctorat en sciences de l’environnement. UQAT, 50.p

LAWRENCE, R.W. et WANG, Y. (1996). Determination of Neutralization Potential for Acid Rock Drainage Prediction, MEND report 1.16.3.

MILLER, S.D., JERRERY, J.J. ET WONG, J.W.C. (1991). Use and misuse of the acidbase account for "AMD" prediction. Proc. of the Second International Conference on the Abatement of Acidic Drainage. Montreal, Canada. 3,489-506

SOBEK, A.A., SCHULLER, W.A., FREEMAN, J.R. et SMITH, R.M. (1978). Field and Laboratory Methods Applicable to Overburdens and Minesoils. EPA-600/2-78-054.


Page 39

Partie 3 : Caractérisation minéralogique

1. Étude

L’objectif principal de l’étude, selon les renseignements obtenus du client, consiste à établir le statut de l’or dans un échantillon de concentré de sulfure. L’étude comprend une caractéri-sation chimique et minéralogique par microscopie optique et microsonde électronique.

2. Préparation de l’échantillon

L’échantillon fourni par Jean Lelièvre est sous forme de poudre, portant la référence « concentré de flottation ». Il a été numéroté selon la référence URSTM : U34289.

3. Caractérisation de l’échantillon

3.1 Caractérisation chimique

La caractérisation chimique de l’échantillon « concentré de flottation » a été réalisée par ICP-AES pour le dosage des éléments chimiques et, spécifiquement, ceux associés aux sulfures de base (S, Cu, Pb, Zn et As). La digestion de l’échantillon est totale et se fait par l’intermédiaire de plusieurs acides très corrosifs (HNO3/Br2/HF/HCl) qui solubilisent, sans exception, tous les minéraux de l’échantillon. L’or a été analysé par pyroanalyse auprès d’un laboratoire externe pour les métaux précieux.

3.2 Caractérisation minéralogique par microscopie optique et microsonde électronique

L’échantillon a fait l’objet d’une étude minéralogique par microscopie optique en lumière réfléchie. Avant les observations, l’échantillon (en poudre) a été monté en section polie qui consiste à l’imprégner dans une résine Epoxy, mélangée à un durcisseur. La section est ensuite polie à l’aide de poudres diamantées sur une polisseuse automatique. La préparation de la section polie a été réalisée selon une méthode spécialement développée à l’URSTM-UQAT pour les échantillons aurifères.

L’étude minéralogique au microscope optique a aussi consisté à préparer l’échantillon aux analyses chimiques ponctuelles, à la microsonde électronique (MSE). La limite de détection de cette dernière est d’environ 80 ppm, permettant de faire des micro-analyses pour le dosage de l’or structural dans les sulfures et d’en faire la quantification. L’or structural est l’or non métallique, qui est invisible et se trouve dans le réseau cristallin des minéraux sulfurés. Il est réfractaire à l’extraction au cyanure.

4. Résultats

4.1 Analyse chimique

L’analyse chimique par ICP-AES montre que l’échantillon contient 20,5 wt.% de soufre et celle obtenue par un four à induction est de 19 wt.%. C’est cette dernière qui est considérée dans ce rapport pour la fiabilité reconnue du four à induction pour le dosage du soufre en forte concentration. La teneur en arsenic dans l’échantillon est de 11,25 wt. % qui représente une moyenne de deux méthodes de digestion (digestion totale, tableau 21) et celle spécifique aux


Page 40

métaux lourds (tableau 22). Les teneurs en Zn, Cu et Pb sont très faibles, ce qui indique que le concentré de flottation ne contient pas de sulfures de Pb, Zn et Cu. L’analyse de l’or par pyroanalyse du « concentré de flottation » donne une teneur de 105,20 g/t.

Tableau 21 : Analyses chimiques totale par ICP-AES de l’échantillon «concentré de flottation»

Éléments

Concentré de flottation

(wt.%) (U34289)

Al

2,94

As

11,49

Ba

0,02

Ca

2,67

Cd

<5 ppm

Co

0,05

Cr

0,09

Cu

0,15

Fe

31,26

K

0,80

Mg

1,30

Mn

0,06

Mo

0,02

Na

0,79

Ni

0,11

Pb

0,005

Stotal

20,16

Sn

<5 ppm

Ti

0,85

Zn 0,06

Tableau 22 : Analyses chimiques des métaux lourds par ICP-AES de l’échantillon «concentré de flottation»

Éléments

(wt.%)

Concentré de flottation

(U34289)

As 11

Be <1 ppm

Bi 0,019

Sb 0,013

Se <3 ppm

Te <2 ppm


Page 41

4.2 Microscopie optique

L’étude de l’échantillon au microscope optique a permis d’identifier uniquement les minéraux métalliques. L’emphase a été mise sur l’observation de l’or libre et les sulfures qui sont susceptibles de contenir de l’or structural comme la pyrite et l’arsénopyrite. Ces observations ont permis l’identification de cinq grains d’or. La figure 18 montre que l’or se présente sous forme de grains libres (Figure 18-A), attaché (Figure 18-B) ou inclus dans l’arsénopyrite (Figure 18-C). Les observations microscopiques ont également montré que l’arsénopyrite et la pyrite sont les minéraux sulfurés les plus abondants, avec des traces de pyrrhotite et quelques grains de chalcopyrite.

Les observations microscopiques ont permis la sélection d’un grand nombre d’arsénopyrite et de pyrite, avec quelques pyrrhotite et chalcopyrites pour les microanalyses à la microsonde électronique (MSE). Ces minéraux sont susceptibles de contenir de l’or structural dans leur réseau cristallin.

Figure 18 : Photographies au microscope optique montrant les trois statuts de l’or

4.3 Microsonde électronique

L’étude minéralogique du « concentré de flottation » au microscope optique a été complétée par des analyses ponctuelles à la MSE pour les arsénopyrites, les pyrites et les pyrrhotites (présélectionnées). Toutes les analyses ponctuelles sont présentées au tableau de l’annexe 4. Les résultats d’analyses à la microsonde montrent que l’or structural est contenu uniquement dans l’arsénopyrite, dont les teneurs en or sont présentées graphiquement à la figure 19. Pour l’ensemble des microanalyses effectuées, la moyenne arithmétique (n=130) des teneurs en or dans l’arsénopyrite est de 178 ppm (ces valeurs ne prennent pas en considération les minéraux dont la teneur en Au est en-dessous de la limite de détection de la microsonde qui est de l’ordre de 85 ppm). Une seule pyrite montre une occurrence en or de l’ordre de 123 ppm. Le tableau 23 synthétise le nombre d’analyses effectuées à la MSE et montre que l’arsénopyrite représente 71 % des sulfures analysés, la pyrite représente 20 % des grains analysés et la pyrrhotite seulement 8 %. Tous les grains ont été photographiés et les photos sont présentées à l’annexe 5.

A B C


Page 42

0

50

100

150

200

250

300

350

400

450

500

0 20 40 60 80 100 120 140 160 180

Ten

eu

r e

n A

u (

pp

m)

Nombre de mesure

Limite de détection de la MSE

Figure 19 : Représentation graphique des teneurs en or dans l’arsénopyrite dans l’échantillon «concentré de flottation»

Tableau 23 : Résumé des observations au microscope optique des minéraux sulfurés de l’échantillon « concentré de flottation »

Nombre de minéraux analysés Concentré de

flottation

Arsenopyrite 109

Arsenopyrite aurifère 21

Pyrrhotite 15

Pyrite 36

Pyrite aurifère 1

Nombre total de minéraux sulfurés traités 182

Pourcentage relatif d'arsénopyrites par rapport aux sulfures 71,4 %

Pourcentage relatif d'arsénopyrites aurifères 19,3 %

Pourcentage de pyrrhotites par rapport aux sulfures 8,2 %

Pourcentage de pyrites par rapport aux sulfures 20,3 %


Page 43

4.4 Quantification de l’or associé à l’arsénopyrite dans l’échantillon « concentré de flottation »

La teneur en or associé aux arsénopyrites est calculée à l’aide de l’équation 1.

[éq.1]

Où « Au i » et « Aire i » sont respectivement la teneur en or et l’aire du minéral i.

Pour l’estimation de la teneur en or structural dans les sulfures, les hypothèses suivantes ont été formulées :

La teneur ponctuelle de l’or mesurée dans un minéral sulfuré, sur environ 1 µm3 de volume, représente une teneur moyenne pour tout le minéral;

Pour les grains soumis à plusieurs mesures ponctuelles, la moyenne des mesures représente la teneur moyenne de l’or dans le grain (cas de 26 analyses);

La teneur moyenne de l’or dans le minéral analysé est pondérée par rapport à sa surface, et la teneur moyenne en or de l’échantillon est établie par rapport à la somme des surfaces de tous les sulfures analysés (aurifères et non aurifères);

Une seule pyrite présente une occurrence positive en or et n’a pas été prise en considération dans les calculs, sachant que son influence est négligeable dans le résultat final

Ainsi, la teneur en or associée aux arsénopyrites dans l’échantillon est donnée par l’équation 2.

[éq.2] Où :

[éq.3]

Le pourcentage massique de l’arsénopyrite dans l’échantillon a été déterminé par calculs minéralogiques, en se basant sur la teneur en As obtenue par analyse chimique (ICP-AES). Le pourcentage massique calculé de l’arsénopyrite est de 25,76 wt. % (éq.3), ce qui a permis de calculer la quantité d’or dans l’échantillon, qui est de 6,04 g/t (éq.2). Cette teneur correspond à 6 % de l’or total obtenu par pyroanalyse et qui est estimé à 105 g. Les détails des calculs sont donnés au tableau 24.

Si on estime que l’erreur sur la mesure de la taille des arsénopyrites est d’environ 5 %, ceci donne une teneur en or dans l’échantillon comprise entre 5,75 (+ 5 % erreur) et 6,35 g/t (- 5 % erreur). À ces erreurs d’estimation de surface, d’autres peuvent être liées à la pyroanalyse elle-même, à la microsonde ou à la statistique (nombre de grains analysés) et peuvent affecter l’estimation de l’or structural par calculs minéralogiques. Toutefois, ces erreurs ne doivent pas être si grandes pour affecter significativement la teneur calculée en or structural. On peut donc estimer que l’Au structural représente environ 6 % de l’Au total de l’échantillon.


Page 44

Tableau 24 : Récapitulatif des calculs pour l’estimation de l’or structural associé à l’arsénopyrite dans l’échantillon «concentré de flottation»

Au associé à l’arsénopyrite

Aire total des arsénopyrites (µm²) 396 757

Au pondéré sur l'aire total des arsénopyrites (ppm) 9 296 773

Au lié aux arsénopyrites dans l'échantillon (ppm) 23,43

Teneur en arsénopyrite dans l'échantillon (wt.%) 25,76

Au structural lié aux pyrites dans l’échantillon (g/t) 6,04

Au par pyroanalyse (g/t) 105

5. Conclusion – caractérisation minéralogique

Les observations minéralogiques par microscopie optique et les microanalyses à l’aide de la microsonde électronique, ainsi que les calculs minéralogiques sur l’échantillon « concentré de flottation » nous ont permis de faire les conclusions suivantes :

L’or structural est contenu uniquement dans l’arsénopyrite;

20 % des arsénopyrites analysées sont aurifères;

La teneur moyenne de l’or dans l’arsénopyrite est de 178 ppm;

La teneur de l’or structural associé aux arsénopyrites est de 6,04 g/t, ce qui représente environ 6 % de l’or total obtenu par pyroanalyse;

94 % de l’or de l’échantillon serait sous forme d’or libre ou inclus dans les minéraux sulfurés, tel qu’observé au microscope optique.

Jean Lelièvre, ing. Mathieu Villeneuve Hassan Bouzahzah, Ph.D. Mélinda Gervais

Annexe 1

Essais métallurgiques détaillés

(sur CD-rom)

Annexe 2

Protocole de cyanuration

(sur CD-rom)

Annexe 3

Certificats d’analyses chimiques

(sur CD-rom)

Annexe 4

Compositions chimiques élémentaires des

pyrites, chalcopyrites et sphalérites par microsonde électronique

(sur CD-rom)

Annexe 5

Photographies au microscope optique de tous les minéraux sulfurés analysés

à la microsonde électronique

(sur CD-rom)

152, avenue Murdoch ~ Rouyn-Noranda (Québec) CANADA J9X 1E2 Tél. : 819 797-3222 ~ Fax : 819 762-6640 ~ www.genivar.com

Caractérisation physicochimique du minerai et des stériles à la propriété

O’Brien

Ressources Minières Radisson Inc. Rouyn-Noranda (Cadillac), Québec

Référence à citer : GENIVAR. 2012. Caractérisation physicochimique du minerai et des stériles à la propriété O’Brien. Rapport réalisé pour Ressources Minières Radisson Inc., 11 pages et annexes.

Ressources Minières Radisson Inc. GENIVAR 121-13415-00 Juillet 2012

i

ÉQUIPE DE RÉALISATION

Ressources Minières Radisson inc.

Eugène Gauthier, ing. Directeur Exploration

GENIVAR

René Fontaine, ing. Directeur Environnement ATNQ

Éric Gingras, M.Sc., EESA® Chargé de projet

Marie-Élise Viger, ing. jr., M.Sc. A. Responsable de l’analyse et de la rédaction

Rénald Lemieux, ing. M. Sc. Env. Collaborateur

Dominic Paiement-Lamothe Collaborateur

Line Poulin Correction et mise en page


iii

TABLE DES MATIÈRES

1 INTRODUCTION .................................................................................................................... 1

1.1 Identification du requérant et des personnes-ressources .............................................. 1

1.2 Description et localisation du terrain visé ...................................................................... 2

2 PROTOCOLE D’ÉCHANTILLONNAGE ................................................................................ 3

3 CARACTÉRISATION PHYSICO-CHIMIQUE DU MINERAI ET DES STÉRILES PRÉSENTS À LA PROPRIÉTÉ RADISSON .......................................................................... 5

3.1 Programme de caractérisation ...................................................................................... 5

3.2 Résultats d’analyse ....................................................................................................... 5

3.2.1 Caractérisation géochimique .............................................................................. 5

3.2.2 Potentiel de génération d’acide .......................................................................... 6

3.2.3 Essais de lixiviation ............................................................................................ 7

3.3 Recommandations d’entreposage ................................................................................. 8

3.3.1 Entreposage du minerai ..................................................................................... 8

3.3.2 Entreposage des stériles .................................................................................... 8

4 CONCLUSION ET RECOMMANDATIONS............................................................................ 9

5 RÉFÉRENCES ..................................................................................................................... 11

ANNEXES

Annexe A Tableaux de résultats

Annexe B Certificats d’analyse du laboratoire

Annexe C Exigences de rejet – Directive 019


1

1 INTRODUCTION

La compagnie Ressources Minières Radisson Inc. désire réaliser un programme d’exploration minière avancée à la propriété O'Brien en aménagement une rampe d’exploration souterraine.

Le programme d’exploration prévoit un échantillonnage en vrac de 50 000 tonnes métriques de minerai et estime l’enlèvement de 50 000 à 100 000 tonnes de stériles.

1.1 Identification du requérant et des personnes-ressources

Requérant et personne responsable :

Ressources Minières Radisson Inc. 153-A, Rue Perreault Val-d'Or, Québec J9P 2H1

Personnes responsables :

M. Eugène Gauthier, ing. Directeur Exploration Téléphone : (819) 874-0030 Télécopieur : (819) 825-1199 Courriel : [email protected]

Consultant et personne responsable :

GENIVAR INC. 152, avenue Murdoch Rouyn-Noranda (Québec) J9X 1E2 Téléphone : (819) 797-3222 Télécopieur : (819) 762-6640

Personnes responsables :

M. René Fontaine, ing. Directeur Env. ATNQ Courriel : [email protected]

M. Éric Gingras, M.Sc., EESA® Chargé de projet Courriel : [email protected]




GENIVAR Ressources Minières Radisson inc. Juillet 2012 121-13415-00

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1.2 Description et localisation du terrain visé

Le projet d’exploration minière présenté par Ressources Minières Radisson Inc. consiste en une phase avancée d’exploration ayant pour but de prélever un échantillonnage en vrac de 50 000 tonnes métriques à des fins d’évaluation métallurgique.

Le secteur directement visé par le projet minier de la propriété O'Brien, appartenant à 100 % à la compagnie Ressources Minières Radisson Inc., est situé au sud du 49e parallèle dans la région administrative de l’Abitibi-Témiscamingue. La propriété est localisée dans la partie ouest du canton de Cadillac, dans le cœur de la ceinture aurifère de l’Abitibi, soit à mi-chemin entre les villes de Rouyn-Noranda et de Val-d’Or. Elle se situe sur les terres du domaine public, l’affectation municipale (zonage) de la propriété s’avère l’exploitation des ressources (minière ou forestière).

Les coordonnées géographiques du gisement du projet minier de la propriété O'Brien sont :

● Latitude nord (NAD 83) : 48°14’32";

● Longitude ouest (NAD 83) : 78°23’20".

La propriété est facilement accessible, étant située directement au nord du quartier Cadillac, avec la route 117 passant à sa limite sud. Une voie ferroviaire est située à proximité de la propriété.

Ressources Minières Radisson Inc. GENIVAR 121-13415-00 Juillet 2012 Août 2010

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2 PROTOCOLE D’ÉCHANTILLONNAGE

La caractérisation de chaque matériau extrait (minerai et stérile) est nécessaire pour connaître les implications environnementales sur les différents milieux récepteurs du site minier. Par conséquent, des analyses géochimiques : le contenu en métaux, la détermination du potentiel de génération d’acide, la détermination du potentiel de neutralisation et la détermination du potentiel de lixiviation ont été effectuées. Ces analyses ont comme but de caractériser conformément les matériaux selon les standards environnementaux de l'industrie minière.

Le nombre d'échantillons représentatifs soumis aux différents essais de laboratoire doit correspondre aux recommandations du Guide de caractérisation des résidus miniers et du minerai, version préliminaire 2003, publié par le Ministère de l'Environnement du Québec, tel qu’indiqué dans le tableau 2-1 ci-dessous.

Compte tenu que l’exploration minière avancée implique l’extraction de 50 000 tonnes métriques de minerai et de 50 000 à 100 000 tonnes métriques de stériles, la deuxième catégorie devra être respectée pour le minerai et le stérile. Par conséquent, huit (8) échantillons de minerai et huit (8) échantillons de stérile minier ont été analysés afin de déterminer la composition géochimique totale et le potentiel de génération d’acide.

En conclusion, le nombre d’échantillons et d’essais de laboratoire effectués sur le minerai et les stériles respecte la deuxième catégorie du tableau 2-1 ci-dessous.

Tableau 2-1 : Nombre d’échantillons requis pour un programme adéquat de caractérisation selon le GCRMM

Catégorie Masse de l'unité géologique qui fera l'objet d'une extraction du

minerai (tonnes)

Nombre minimum d'échantillons requis aux fins d'analyses

1 10 000 3

2 10 000 et 100 000 entre 3 et 8

3 100 000 et 1 000 000 entre 8 et 26

4 1 000 000 et 10 000 000 entre 26 et 80

5 10 000 000 144

Note : a Données basées sur une relation mathématique.


5

3 CARACTÉRISATION PHYSICO-CHIMIQUE DU MINERAI ET

DES STÉRILES PRÉSENTS À LA PROPRIÉTÉ RADISSON

3.1 Programme de caractérisation

En juin 2012, seize (16) échantillons, (8 échantillons de minerai et 8 échantillons de stérile) ont été prélevés puis analysés.

Ces échantillons provenant de différentes lithologies et zones ont été prélevés par un géologue de Ressources Minières Radisson Inc. et sont représentatifs du gisement anticipé.

Les différents essais réalisés consistent en une description des caractéristiques géochimiques, le potentiel de génération d’acide et des analyses chimiques en condition acide. Ces différents essais permettent de caractériser le minerai et les stériles selon les définitions de l’annexe II de la Directive 019.

Les résultats d’analyses sont présentés aux tableaux 1 à 3 (annexe A) tandis que les certificats d’analyses chimiques sont consignés à l’annexe B. Toutes les analyses ont été effectuées par le laboratoire Multilab Direct de Rouyn-Noranda.

Les essais suivants ont été réalisés sur tous les échantillons prélevés et sont présentés dans les tableaux 1 à 3 :

analyse de dix-neuf (19) éléments chimiques (paramètres);

potentiel de génération d’acide selon des essais statiques réalisés par la méthode prescrite par le Centre d’expertise en analyse environnementale du Québec;

essai de lixiviation par la méthode TCLP 1311.

Les résultats obtenus furent comparés aux critères génériques pour les sols et les eaux souterraines selon la Politique de protection des sols et de réhabilitation des terrains contaminés (Politique) ainsi qu’à la Directive 019.

3.2 Résultats d’analyse

3.2.1 Caractérisation géochimique Des 19 métaux et autres composés inorganiques analysés, 15 font partie de la Grille des critères génériques pour les sols de la Politique du MDDEP.

Le tableau 1 de l’annexe A présente le contenu disponible en éléments provenant des échantillons du minerai et des stériles, de même que les critères de la Politique de protection des sols et de réhabilitation de terrains contaminés. Le critère « A » a été déterminé à partir des données suggérées pour la province géologique du Supérieur.


6

Des dépassements du critère « A » sont observés pour tous les échantillons analysés pour les paramètres arsenic, baryum, cadmium, chrome, cobalt, cuivre, manganèse, molybdène, nickel, sélénium et soufre. Ainsi le minerai et les stériles ne peuvent pas être considérées à faible risque.

Pour le minerai, les huit échantillons analysés présentent des concentrations supérieures au critère « C » de la Politique pour les paramètres arsenic et soufre. De plus, quatre échantillons, soit 50%, dépassent le critère « B » de la Politique pour le cadmium dont un dépassant le critère « C ». Finalement, un échantillon présente un dépassement du critère « B » pour le sélénium.

Pour les stériles, trois échantillons, soit 37,5% dépassent le critère « C » pour l’arsenic. Pour les paramètres de cadmium, cobalt et cuivre, un dépassement du critère « B » est observé. Deux échantillons présentent des concentrations supérieures au critère « B » dont une supérieure au critère « C » pour le chrome et le nickel. Trois échantillons montrent des concentrations supérieures au critère « B » de la Politique pour le manganèse. Finalement, tous les échantillons présentent des concentrations en soufre supérieure au critère « B » dont cinq supérieures au critère « C » de la Politique.

3.2.2 Potentiel de génération d’acide Le potentiel de génération d’acide représente la quantité d’ions H+ qui sont générés par l’oxydation du matériel, principalement par l’oxydation de minéraux sulfureux, et qui ne sont pas neutralisés par des formations telles que des carbonates.

Selon l’annexe II de la Directive 019, les résidus miniers sont considérés acidogènes s’ils contiennent « du soufre (Stotal) en quantité supérieure à 0,3 % et dont le potentiel de génération acide a été confirmé par des essais de prévision statiques ».

Des résidus miniers sont considérés potentiellement générateurs d’acide, selon les critères suivants :

Le potentiel net de neutralisation (PNN) d’acide est inférieur à 20 kg CaCO3/t;

Le rapport du potentiel de neutralisation d’acide sur le potentiel de génération d’acide (PN/PA) est inférieur à 3.

Le tableau 2 de l’annexe A présentent les résultats des essais réalisés pour déterminer le potentiel de génération d’acide sur les échantillons de minerai et de stériles.

Les huit échantillons de minerai ont présenté un contenu en soufre supérieur à 0,3 %. Cinq échantillons ont un ratio PN/PA inférieur à 3. Par contre, ces échantillons ont tous un PNN supérieur à 23 kg CaCO3/tonne. Il n’est donc pas attendu que le minerai produise de l’acide lors de son entreposage.


7

Pour ce qui est des stériles, trois échantillons ont présenté un contenu en soufre supérieur à 0,3 %. Toutefois, seul un échantillon (Stérile #3 - 136104) a présenté à la fois :

Un PNN de 9,9 (inférieur à 20 kg CaCO3/ton);

Un ratio PN/PA de 1,5 (inférieur à 3).

Bien que cet échantillon neutralise une fois l’acide généré, il est considéré comme potentiellement générateur d’acide. Cet échantillon représente 12,5% des échantillons de stériles analysés. Les deux autres échantillons présentant un contenu en soufre >0,3% montrent des PNN supérieurs à 139 kg CaCO3/ton et des ratios PN/PA supérieurs 14,5.

L’acide potentiellement généré sera neutralisé par le fort potentiel neutralisant des autres stériles. Il n’est donc pas attendu que les stériles produisent de l’acide lors de leur entreposage.

3.2.3 Essais de lixiviation Les essais de lixiviation permettent d’évaluer la mobilité des espèces inorganiques sous conditions acides. Le test TCLP (Toxicity Characteristic Leaching Test) permet de déterminer si un résidu est lixiviable ou non. De plus, il permet d’évaluer le type de contamination pouvant se retrouver dans l’eau de ruissellement.

Les tableaux 3 de l’annexe A présentent les résultats des essais de lixiviation par la méthode TCLP 1311 effectués sur les échantillons de minerai et de stériles.

Selon les essais de lixiviation, tous les échantillons de minerai ainsi que 50% des échantillons de stériles sont classifiés comme lixiviables : les teneurs obtenues pour les différents échantillons au site minier O’Brien de Radisson sont toutes inférieures aux valeurs limites des concentrations maximales selon le tableau 1 de l’annexe II de la Directive 019. Toutefois, certaines valeurs sont supérieures aux critères de Résurgence dans les eaux de surface ou infiltration dans les égouts (RÉSIE) selon la Politique de protection des sols et de réhabilitation des terrains contaminés.

Les paramètres dépassant les critères de RESIE sont l’aluminium (pour un échantillon de minerai et un de stérile), l’arsenic (pour six échantillons de minerai), le chrome (pour sept échantillons de minerai et quatre de stérile), le cuivre (pour deux échantillons de minerai), le nickel (pour quatre échantillons de minerai et un de stérile) et le plomb (pour un échantillon de minerai).

Les échantillons de minerai et de stériles ne présentent pas de potentiel de génération d’acide. De plus, depuis les années 2000, la construction de l’usine d’acide sulfurique à la Fonderie Horne permet de récupérer et d’éliminer 90% du dioxyde de soufre (SO2 – responsable de l’acidification des pluies). Il est donc peu probable de reproduire naturellement les conditions des essais TCLP sur les haldes d’entreposage.


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3.3 Recommandations d’entreposage

Selon les résultats d’analyse, le minerai ainsi que les stériles de la propriété O’Brien sont classés lixiviables.

3.3.1 Entreposage du minerai L’entreposage du minerai peut être fait à ciel ouvert. Par contre, puisque celui-ci est lixiviable, il est recommandé de l’entreposer sur une surface étanche.

Les eaux de ruissellement doivent être captées et traitées au besoin afin qu’elles respectent les critères de la colonne II du tableau 2.1 de la section 2.1.1.1 de la Directive 019. Le tableau est présenté à l’annexe C.

Une fois par trimestre, les eaux de ruissellement provenant de l’entreposage doivent être analysées pour les paramètres du tableau ainsi que le pH et le débit. Ces résultats doivent être transmis au MDDEP. Selon les résultats de lixiviation, le chrome peut aussi être lixiviable (7 échantillons de minerai ont présenté ce paramètre), il est donc recommandé d’inclure ce paramètre dans l’analyse des eaux de ruissellement.

Il est fort probable que le minerai enrichi ou le concentré soit lui aussi lixiviable. Si tel est le cas, celui-ci devra être entreposé sous un abri et sur une surface étanche et équipée d’un système de récupération des eaux de lixiviation.

3.3.2 Entreposage des stériles De manière générale, l’aire d’accumulation des stériles doit être située à au moins 60 mètres de la ligne des hautes eaux d’un cours d’eau à débit régulier ou intermittent visé par l’application de la Politique de protection des rives, du littoral et des plaines inondables.

L’aire d’accumulation des stériles miniers lixiviables doit être conçue de manière à empêcher le transport de contaminants vers les eaux souterraines. Cela dit, un réseau de captage de l’eau de percolation (fossés de drainage autour des haldes) doit être installé pour acheminer l’eau à un système de traitement.

Comme les eaux de ruissellement du minerai, les eaux de ruissellement des stériles doivent respectés les critères de la colonne II du tableau 2.1 de la section 2.1.1.1 de la Directive 019. L’analyse pour le chrome est aussi recommandée.


9

4 CONCLUSION ET RECOMMANDATIONS

Les échantillons de la propriété O’Brien furent soumis aux analyses pour le contenu géochimique, le potentiel générateur d’acide (PGA tel que décrit par le CEAEQ) et l’analyse chimique sous des conditions acides (TCLP 1311).

Suite aux différentes analyses réalisées sur les échantillons de minerai (8) et de stériles (8) de la propriété minière O’Brien, les constats et recommandations sont les suivants :

La roche présente sur le site contient naturellement de l’arsenic et des sulfures en concentration supérieure au critère « C » de la Politique. Ainsi, le minerai et les stériles ne peuvent pas êtres considérés à faible risque. Des concentrations supérieures au critère « C » ont aussi été observés pour les paramètres cadmium, chrome et nickel.

Malgré le contenu élevé en soufre, tous les échantillons de minerai présentent un PNN supérieur à 20 kg CaCO3/tonne. Il n’est donc pas attendu que celui-ci génère de l’acide. Pour les stériles, un échantillon a présenté un potentiel de génération d’acide (PNN de 9,9 kg CaCO3/tonne et un ratio PN/PA de 1,5). Par contre, tous les autres échantillons ne sont pas potentiellement générateur d’acide et possèdent un fort potentiel de neutralisation (tous supérieurs à 139 kg CaCO3/tonne). Il n’est donc pas attendu qu’un drainage acide survienne sur la halde à stérile.

Selon les essais de lixiviation TCLP 1311, tous les échantillons de minerai ainsi que 50% des échantillons de stériles sont classifiés comme étant lixiviables. Les paramètres lixiviables sont l’aluminium, l’arsenic, le chrome, le cuivre, le nickel et le plomb.

Par conséquent, le minerai et les stériles peuvent être entreposés sur des haldes à ciel ouvert comportant des fossés de drainage afin de récupérer les eaux de ruissellement. Ces eaux devront être analysées une fois par trimestre pour les paramètres identifiés au tableau 2.1 de la section 2.1.1.1 de la Directive 019 en plus du paramètre chrome. Les eaux peuvent nécessités un traitement si elles ne rencontrent pas les exigences de rejet de la colonne II du tableau 2.1.

Ressources Minières Radisson inc. GENIVAR 121-13415-00 Juillet 2012

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5 RÉFÉRENCES

CENTRE D’EXPERTISE ET ANALYSE ENVIRONNEMENTALE DU QUÉBEC (CEAEQ). 2005. Protocole de lixiviation pour les espèces inorganiques, MA. 100 – Lix.com. 1.0. 17 p.

MINISTÈRE DE L'ENVIRONNEMENT DU QUÉBEC, DIRECTION DES POLITIQUES DU SECTEUR INDUSTRIEL. 2003. Guide de caractérisation des résidus miniers et du minerai, version préliminaire.

MINISTÈRE DES RESSOURCES NATURELLES ET DE LA FAUNE. 1997. La restauration des sites miniers : Guide et modalités de préparation du plan et exigences générales en matière de restauration des sites miniers au Québec.

MINISTÈRE DU DÉVELOPPEMENT DURABLE, DE L’ENVIRONNEMENT ET DES PARCS. 2012. Directive 019 sur l’industrie minière.

MINISTÈRE DU DÉVELOPPEMENT DURABLE, DE L’ENVIRONNEMENT ET DES PARCS. 1999. Politique de protection des sols et de réhabilitation des terrains contaminés. 74 pages + annexes.

MINISTÈRE DU DÉVELOPPEMENT DURABLE, DE L’ENVIRONNEMENT ET DES PARCS. 2008. Politique de protection des rives, du littoral et des plaines inondables.

Annexe A Tableaux de résultats

Tableau 1 (1 de 2)Caractérisation géochimique des échantillons de minerai et de stérilesRouyn-Noranda (Cadillac), QcN/réf: 121-13415-00

Minerai #1 Minerai #2 Minerai #3 Minerai #4 Minerai #5 Minerai #6 Minerai #7 Minerai #8A B C 136099 136100 136101 136110 136111 136112 136113 136114

Aluminium Al mg/kg --- --- --- 0,6 14424 13916 22307 15252 19168 18404 14402 16620 16812Antimoine Sb mg/kg --- --- --- 0,1 7,9 7,5 7,7 1,8 1 1,4 0,5 2 3,7Argent Ag mg/kg 0,5 20 40 2 <2 <2 <2 <2 <2 <2 <2 2 <2Arsenic As mg/kg 5 30 50 0,05 3226 4382 2050 18144 17737 13615 13833 11659 10581Baryum Ba mg/kg 200 500 2000 0,01 119 99 139 69,4 188 128 100 162 126Bore B mg/kg --- --- --- 0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01Cadmium Cd mg/kg 0,9 5 20 0,005 17,160 21,300 11,610 0,182 9,900 0,122 0,145 0,127 7,568Chrome Cr mg/kg 85 250 800 0,05 142 157 138 82 116 81 75 136 116Cobalt Co mg/kg 20 50 300 0,05 18,3 16,5 32,5 22,2 25,1 6,6 4,5 14,2 17,5Cuivre Cu mg/kg 50 100 500 5 54 35 68 97 52 52 60 42 58Fer Fe mg/kg --- --- --- 0,5 32990 29893 62208 68876 69851 47200 47263 55753 51754Manganèse Mn mg/kg 1000 1000 2200 0,05 604 508 1248 1384 1376 802 987 1038 993Mercure Hg mg/kg 0,3 2 10 0,1 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01Molybdène Mo mg/kg 6 10 40 0,05 6,9 6,5 2,9 0,92 1,0 1,9 1,0 1,0 2,8Nickel Ni mg/kg 50 100 500 0,05 54,9 50,2 39,8 46,2 42,9 37,2 34 44,2 43,7Plomb Pb mg/kg 40 500 1000 0,05 16,6 17,4 5,8 <0,05 <0,05 <0,05 <0,05 <0,05 5,0Sélénium Se mg/kg 3 3 10 0,05 0,06 0,13 <0,05 <0,05 <0,05 5,6 <0,05 <0,05 0,8Soufre S % 0,04 0,1 0,2 0,1 1,30 1,30 2,30 2,20 2,10 1,60 1,60 1,80 1,78Zinc Zn mg/kg 110 500 1500 0,05 41,9 47,7 8,5 65 <0,05 38 33,0 8 34,6

NOTE:(1) : Limite de détection rapportée par le laboratoire.(2) : Pour une valeur inférieure à la limite de détection rapportée (LDR), la concentration utilisée correspond à |LDR|.

LÉGENDE:100 : Concentration < A100 : Concentration = A100 : Concentration > A et < B100 : Concentration > B et < C100 : Concentration > C

Moyenne minerai(2)Éléments

Échantillon CritèresLDR(1)

Tableau 1 (2 de 2)Caractérisation géochimique des échantillons de minerai et de stérilesRouyn-Noranda (Cadillac), QcN/réf: 121-13415-00

Stérile #1 Stérile #2 Stérile #3 Stérile #4 Stérile #5 Stérile #6 Stérile #7 Stérile #8A B C 136102 136103 136104 136105 136106 136107 136108 136109

Aluminium Al mg/kg --- --- --- 0,6 24659 31985 25495 22364 40183 21085 16831 16955 28937Antimoine Sb mg/kg --- --- --- 0,1 10,6 23,6 0,4 9,7 10,1 0,2 <0,1 2 7,1Argent Ag mg/kg 0,5 20 40 2 <2 <2 <2 <2 <2 <2 <2 <2 <2Arsenic As mg/kg 5 30 50 0,05 28,40 667,00 458,0 199,00 16,20 <0,05 15,2 29,5 176,67Baryum Ba mg/kg 200 500 2000 0,01 177 4 191 301 9 35 <0,01 122 105Bore B mg/kg --- --- --- 0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01Cadmium Cd mg/kg 0,9 5 20 0,005 <0,005 2,170 1,970 0,341 <0,005 8,07 0,268 0,158 2,163Chrome Cr mg/kg 85 250 800 0,05 247 1954 220 387 173 125 81,2 170 510,4Cobalt Co mg/kg 20 50 300 0,05 24,9 67,0 25,0 22,4 45,7 9,3 33 5 29,0Cuivre Cu mg/kg 50 100 500 5 48 64 61 33 91 23 146 19 59Fer Fe mg/kg --- --- --- 0,5 37116 58712 47190 37175 74008 47093 60180 37423 50840Manganèse Mn mg/kg 1000 1000 2200 0,05 360 1259 663 643 1124 473 1373 385 810Mercure Hg mg/kg 0,3 2 10 0,1 <0,1 0,2 <0,1 <0,1 <0,1 <0,1 <0,1 <0,1 0,1Molybdène Mo mg/kg 6 10 40 0,05 6,6 4,3 6,2 2,4 7,6 2,6 1 1,5 4,0Nickel Ni mg/kg 50 100 500 0,05 86 685 63 112 56 89 53 84 153Plomb Pb mg/kg 40 500 1000 0,05 12,2 2,4 8,1 6,3 7,6 <0,05 <0,05 <0,05 4,6Sélénium Se mg/kg 3 3 10 0,05 <0,05 <0,05 <0,05 <0,05 <0,05 <0,05 <0,05 <0,05 <0,05Soufre S % 0,04 0,1 0,2 0,1 0,27 0,11 0,61 0,30 0,33 0,17 0,61 0,19 0,32Zinc Zn mg/kg 110 500 1500 0,05 63 28 59 59 106 58 10 71 57

NOTE:(1) : Limite de détection rapportée par le laboratoire.(2) : Pour une valeur inférieure à la limite de détection rapportée (LDR), la concentration utilisée correspond à |LDR|.

LÉGENDE:100 : Concentration < A100 : Concentration = A100 : Concentration > A et < B100 : Concentration > B et < C100 : Concentration > C

ÉchantillonÉléments LDR(1) Moyenne

stérile(2)Critères

Tableau 2 (1 de 2)Résultats du potentiel générateur d'acide pour les échantillons de minerai et de stérilesRouyn-Noranda (Cadillac), QcN/réf: 121-13415-00

Minerai #1 Minerai #2 Minerai #3 Minerai #4 Minerai #5 Minerai #6 Minerai #7 Minerai #8136099 136100 136101 136110 136111 136112 136113 136114

PN kg CaCO3/ t 63 89 236 157 207 83 126,0 177 142,25PA kg CaCO3/ t 40,0 39,4 71,9 69,7 66,3 49,3 51,1 57,5 55,65PNN kg CaCO3/ t 23,0 49,6 164,0 87,3 141,0 33,7 74,9 120 86,69PN/PA ratio 1,6 2,3 3,3 2,3 3,1 1,7 2,5 3,1 2,46S % 1,3 1,3 2,3 2,2 2,1 1,6 1,6 1,8 1,78Potentiel --- NPGA NPGA NPGA NPGA NPGA NPGA NPGA NPGA NPGA

NOTE:(1) : Pour une valeur inférieure à la limite de détection rapportée (LDR), la concentration utilisée correspond à |LDR|.

LÉGENDE:PN: Potentiel de neutralisation d'acidePA: Potentiel d'acidité des sulfuresPNN: Potentiel de neutralisation net, calculé comme PN-PA

0,1 : PN/PA < 3.015 : PNN < 20 kg CaCO3/ton5 : sulfure > 0,3 %

NPGA : non potentiel générateur d'acideZG : Zone grise

PGA : Générateur d'acide

Un échantillon potentiellement générateur d'acide est défini comme suit:"Un échantillon est considéré comme potentiellement générateur d'acide si le pourcentage de soufre est supérieur à 0,3% et le potentiel net de neutralisation inférieur ou égale à 20 kg/tonne de CaCO3."

Tiré de : CENTRE D'EXPERTISE EN ANALYSE ENVIRONNEMENTALE DU QUÉBEC, Détermination du potentiel de génération d'acide: méthode par titrage avec de l'acide sulfurique. MA. 110 - PGA 1.0, Rév. 3,Ministère du Développement durable, de l'Environnement et des Parcs du Québec, 2006, 10 p.

Échantillon Moyenne minerai(1)PGA

Tableau 2 (2 de 2)Résultats du potentiel générateur d'acide pour les échantillons de minerai et de stérilesRouyn-Noranda (Cadillac), QcN/réf: 121-13415-00

Stérile #1 Stérile #2 Stérile #3 Stérile #4 Stérile #5 Stérile #6 Stérile #7 Stérile #8136102 136103 136104 136105 136106 136107 136108 136109

PN kg CaCO3/ t 46 259 29 158 149 34 314 24 126,63PA kg CaCO3/ t 8,40 3,40 19,10 9,40 10,30 5,30 19,0 6,0 10,11PNN kg CaCO3/ t 37,6 256,0 9,9 149,0 139,0 28,7 285,0 18,0 115,40PN/PA ratio 5,5 76,2 1,5 16,8 14,5 6,4 16,5 4,0 17,67S % 0,27 0,11 0,61 0,30 0,33 0,17 0,61 0,19 0,32Potentiel --- NPGA NPGA PGA NPGA NPGA NPGA NPGA NPGA NPGA

NOTE:(1) : Pour une valeur inférieure à la limite de détection rapportée (LDR), la concentration utilisée correspond à |LDR|.

LÉGENDE:PN: Potentiel de neutralisation d'acidePA: Potentiel d'acidité des sulfuresPNN: Potentiel de neutralisation net, calculé comme PN-PA

0,1 : PN/PA < 3.015 : PNN < 20 kg CaCO3/ton5 : sulfure > 0,3 %

NPGA : non potentiel générateur d'acideZG : Zone grise

PGA : Générateur d'acide

Un échantillon potentiellement générateur d'acide est défini comme suit:"Un échantillon est considéré comme potentiellement générateur d'acide si le pourcentage de soufre est supérieur à 0,3% et le potentiel net de neutralisation inférieur ou égale à 20 kg/tonne de CaCO3."

Tiré de : CENTRE D'EXPERTISE EN ANALYSE ENVIRONNEMENTALE DU QUÉBEC, Détermination du potentiel de génération d'acide: méthode par titrage avec de l'acide sulfurique. MA. 110 - PGA 1.0, Rév. 3,Ministère du Développement durable, de l'Environnement et des Parcs du Québec, 2006, 10 p.

Moyenne stérile(2)

ÉchantillonPGA

Tableau 3 (1 de 2)Résultats des analyses chimiques (essai de lixiviation TCLP 1311) des échantillons de minerai et de stérilesRouyn-Noranda (Cadillac), QcN/réf: 121-13415-00

Minerai #1 Minerai #2 Minerai #3 Minerai #4 Minerai #5 Minerai #6 Minerai #7 Minerai #8136099 136100 136101 136110 136111 136112 136113 136114

pH 6,5> <8,5 --- --- --- 5,53 6,35 5,67 - - - - - 5,850Aluminium Al mg/L --- 0,75 --- 0,006 0,43 0,03 <0,006 0,219 0,038 <0,006 1,66 0,189 0,322Antimoine Sb mg/L 0,006 -- --- 0,0001 0,0030 0,0076 <0,0001 0,0072 0,0074 0,0040 0,0063 0,0081 0,005Argent Ag mg/L 0,1 0,00062 --- 0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005Arsenic As mg/L 0,025 0,34 5 0,0005 0,5905 0,1070 0,1591 1,504 0,9565 0,3325 3,9560 1,0860 1,086Baryum Ba mg/L 1 5,3 100 0,0005 0,130 0,129 0,335 0,190 0,386 0,232 0,277 0,348 0,253Bore B mg/L --- --- --- 0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01Cadmium Cd mg/L 0,005 0,0021 0,5 0,0001 0,0003 0,0001 <0,0001 0,0007 0,0008 0,0005 0,0013 0,0005 0,0006Chrome Cr mg/L 0,05 0,0016 5 0,0006 0,0225 0,0066 <0,0006 0,0242 0,0296 0,0289 0,0646 0,0372 0,0268Cobalt Co mg/L --- 0,5 --- 0,001 0,004 0,003 <0,001 0,025 0,022 0,009 0,027 0,024 0,014Cuivre Cu mg/L 1 0,0073 --- 0,0005 0,0023 0,0060 0,0196 0,0019 0,0012 0,0226 0,0017 0,0012 0,0071Fer Fe mg/L --- --- --- 0,01 6,4 2,0 19,3 13,3 12,7 1,1 23,1 13,5 11,425Manganèse Mn mg/L 0,05 --- --- 0,0005 8,3870 7,7020 18,8500 17,7000 19,98 8,8910 21,3500 17,1400 15,0000Mercure Hg mg/L 0,001 0,00013 0,1 0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002Molybdène Mo mg/L 0,07 2 --- 0,0005 0,0029 0,0025 <0,0005 <0,0005 <0,0005 0,0021 0,0009 <0,0005 0,0021Nickel Ni mg/L 0,02 0,26 --- 0,0005 0,0638 0,0776 0,0285 0,4904 0,4821 0,1917 0,4249 0,4978 0,2821Plomb Pb mg/L 0,01 0,034 5 0,0005 0,0353 <0,0005 0,0005 0,0152 0,0073 0,0035 0,0281 0,0110 0,0144Sélénium Se mg/L 0,01 0,02 1 0,001 0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001Zinc Zn mg/L 5 0,067 --- 0,001 0,039 0,030 <0,001 0,034 0,015 0,032 0,037 0,009 0,026

NOTES:(1) : Critères d'eau souterraine aux fins de consommation de la Politique de protection des sols et de réhabilitation des terrains contaminés(2) : Critères de résurgence dans les eaux de surfaces ou infiltration dans les égouts de la Politique de protection des sols et de réhabilitation des terrains contaminés(3) : Concentrations maximales dans un liquide ou un lixiviat d'une matière solide, tiré du tableau 1 de l'annexe II de la Directives 019(4) : Limite de détection rapportée par le laboratoire.(5) : Pour une valeur inférieure à la limite de détection rapportée (LDR), la concentration utilisée correspond à |LDR|.

LÉGENDE:- : Non défini ou non analysé

100 : Concentration supérieure au critère RÉSIE100 : Concentration supérieure au critère de la directive 019

LDR(4) Moyenne(5)EC(1) RÉSIE(2) Directive 019(3)Paramètres

Échantillons

Tableau 3 (2 de 2)Résultats des analyses chimiques (essai de lixiviation TCLP 1311) des échantillons de minerai et de stérilesRouyn-Noranda (Cadillac), QcN/réf: 121-13415-00

Stérile #1 Stérile #2 Stérile #3 Stérile #4 Stérile #5 Stérile #6 Stérile #7 Stérile #8136102 136103 136104 136105 136106 136107 136108 136109

pH 6,5> <8,5 --- --- --- 5,47 5,02 5,26 6,2 5,73 - - - 5,54Aluminium Al mg/L --- 0,75 --- 0,006 <0,006 <0,006 <0,006 <0,006 0,34 0,703 0,03 1,95 0,381Antimoine Sb mg/L 0,006 -- --- 0,0001 <0,0001 <0,0001 <0,0001 <0,0001 0,0011 0,0025 0,0003 0,0014 0,0057Argent Ag mg/L 0,1 0,00062 --- 0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005Arsenic As mg/L 0,025 0,34 5 0,0005 <0,0005 0,0501 0,2080 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 0,033Baryum Ba mg/L 1 5,3 100 0,0005 0,207 0,037 0,277 0,424 0,019 0,068 0,0275 0,148 0,151Bore B mg/L --- --- --- 0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01 <0,01Cadmium Cd mg/L 0,005 0,0021 0,5 0,0001 <0,0001 <0,0001 <0,0001 <0,0001 0,0010 <0,0001 0,0012 <0,0001 0,0004Chrome Cr mg/L 0,05 0,0016 5 0,0006 <0,0006 <0,0006 <0,0006 <0,0006 0,0238 0,036 0,0293 0,0421 0,0167Cobalt Co mg/L --- 0,5 --- 0,001 <0,001 0,009 <0,001 <0,001 0,008 0,006 0,026 0,004 0,007Cuivre Cu mg/L 1 0,0073 --- 0,0005 0,0196 0,0070 0,0197 0,0202 0,0037 0,0239 0,0014 0,0254 0,015Fer Fe mg/L --- --- --- 0,01 2,4 40,9 4,3 4,4 13,8 7,3 11,1 7,5 11,463Manganèse Mn mg/L 0,05 --- --- 0,0005 1,9580 12,7200 3,1400 4,8650 16,0100 0,9857 12,30 0,7625 6,5927Mercure Hg mg/L 0,001 0,00013 0,1 0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002 <0,00002Molybdène Mo mg/L 0,07 2 --- 0,0005 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 0,0021 0,0007 0,0022 0,0017Nickel Ni mg/L 0,02 0,26 --- 0,0005 0,0148 0,1271 0,0163 0,0029 0,0660 0,1142 0,4630 0,0573 0,1077Plomb Pb mg/L 0,01 0,034 5 0,001 <0,0005 <0,0005 <0,0005 <0,0005 <0,0005 0,0118 0,0013 0,0013 0,0048Sélénium Se mg/L 0,01 0,02 1 0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001 <0,001Zinc Zn mg/L 5 0,067 --- 0,001 <0,001 <0,001 <0,001 <0,001 0,017 <0,001 0,040 0,020 0,010

NOTES:(1) : Critères d'eau souterraine aux fins de consommation de la Politique de protection des sols et de réhabilitation des terrains contaminés(2) : Critères de résurgence dans les eaux de surfaces ou infiltration dans les égouts de la Politique de protection des sols et de réhabilitation des terrains contaminés(3) : Concentrations maximales dans un liquide ou un lixiviat d'une matière solide, tiré du tableau 1 de l'annexe II de la Directives 019(4) : Limite de détection rapportée par le laboratoire.(5) : Pour une valeur inférieure à la limite de détection rapportée (LDR), la concentration utilisée correspond à |LDR|.

LÉGENDE:- : Non défini ou non analysé

100 : Concentration supérieure au critère RÉSIE100 : Concentration supérieure au critère de la directive 019

Moyenne stérile(2)EC(1) RÉSIE(2) Directive

019(3) LDR(4)ÉchantillonsParamètres

Annexe B Certificats d’analyse du laboratoire

Client : Genivar Inc

Responsable : Mme Marie-Élise Viger

Adresse : 152, avenue Murdoch

Rouyn-Noranda Québec J9X 1E1

tél.: (819) 797-3222 (298)

fax.: (819) 762-6640

Numéro de projet : C-113484

Lieu de prélèvement : Radisson Date de prélèvement : 17 avril 2012

Échantillon : 136099 Heure de prélèvement : N/D

Nom du préleveur : Eugène Gauthier Date de réception : 18 mai 2012

Type d'échantillon : Minerai

Réseau: 121-131415-00

Date d'émission : 20 juin 2012

Les résultats ne se rapportent qu'aux échantillons soumis pour analyse.

Les échantillons seront conservés pendant 30 jours à partir de la date du rapport à moins d'avis écrit du client.

Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.

Certificat d'analyse

Sauf indication contraire, tous les échantillons ont été reçus en bon état.

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Version 3ième: 26/10/2005

Numéro de projet : C-113484Échantillon : 136099 Date de prélèvement :

Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres Résultats Méthode d'analyse Date d'analyseAluminium (Al) 14424 mg/Kg M-MET-3.0 30 mai 2012Antimoine (Sb) 7.9 mg/Kg M-MET-3.0 30 mai 2012Arsenic (As) 3226(> C) mg/Kg M-MET-3.0 05 juin 2012Baryum (Ba) 119(< A) mg/Kg M-MET-3.0 30 mai 2012Bore (B) <0.01 mg/Kg M-MET-3.0 30 mai 2012Cadmium (Cd) 17.16(B-C) mg/Kg M-MET-3.0 30 mai 2012Chrome (Cr) 142(A-B) mg/Kg M-MET-3.0 30 mai 2012Cobalt (Co) 18.3(< A) mg/Kg M-MET-3.0 30 mai 2012Cuivre (Cu) 54(A-B) mg/Kg M-MET-3.0 30 mai 2012Fer (Fe) 32990 mg/Kg M-MET-3.0 30 mai 2012Manganèse (Mn) 604(< A) mg/Kg M-MET-3.0 30 mai 2012Molybdene (Mo) 6.9(A-B) mg/Kg M-MET-3.0 30 mai 2012Nickel (Ni) 54.9(A-B) mg/Kg M-MET-3.0 30 mai 2012Plomb (Pb) 16.6(< A) mg/Kg M-MET-3.0 30 mai 2012Préparation d'échantillon Sous-traitance\Laboratoire Expert Inc.

Sélénium (Se) 0.06(< A) mg/Kg M-MET-3.0 05 juin 2012Zinc (Zn) 41.9(< A) mg/Kg M-MET-3.0 30 mai 2012Potentiel générateur acide Sous-traitance\Maxxam Analytics Inc

Soufre 1.3 % Sous-traitance\Maxxam Analytics Inc 12 juin 2012Potentiel d'acidité maximal (PA) 40.0 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisation brut (PN) 63 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisaton net (PNN) 23.0 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012% Humidité <0.1 % M-HUM-1.0 29 mai 2012Mercure (Hg) <0.1 mg/Kg M-HG-2.0 05 juin 2012Argent (Ag) <2 mg/Kg M-MET-4.0 30 mai 2012




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Lieu de prélèvement : Radisson Heure de prélèvement :Paramètre Valeur Unité Méthode AccréditationAluminium (Al) 0.6 mg/Kg M-MET-3.0 OuiAntimoine (Sb) 0.1 mg/Kg M-MET-3.0Arsenic (As) 0.05 mg/Kg M-MET-3.0 OuiBaryum (Ba) 0.01 mg/Kg M-MET-3.0 OuiBore (B) 0.01 mg/Kg M-MET-3.0Cadmium (Cd) 0.005 mg/Kg M-MET-3.0 OuiChrome (Cr) 0.05 mg/Kg M-MET-3.0 OuiCobalt (Co) 0.05 mg/Kg M-MET-3.0 OuiCuivre (Cu) 5 mg/Kg M-MET-3.0 OuiFer (Fe) 0.5 mg/Kg M-MET-3.0Manganèse (Mn) 0.05 mg/Kg M-MET-3.0 OuiMolybdene (Mo) 0.05 mg/Kg M-MET-3.0 OuiNickel (Ni) 0.05 mg/Kg M-MET-3.0 OuiPlomb (Pb) 0.05 mg/Kg M-MET-3.0 OuiSélénium (Se) 0.05 mg/Kg M-MET-3.0 OuiZinc (Zn) 0.05 mg/Kg M-MET-3.0 Oui% Humidité 0.1 % M-HUM-1.0Mercure (Hg) 0.1 mg/Kg M-HG-2.0 OuiArgent (Ag) 2 mg/Kg M-MET-4.0 Oui

Limite de détection rapportée

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Lieu de prélèvement : Radisson Heure de prélèvement :ParamètresAluminium (Al) mg/Kg Blanc <0.6

Nom Standard DMR.0231-2012-1Valeur obtenue 36907

Justesse 97.1%Intervalle 32300 - 43700

Antimoine (Sb) mg/Kg Blanc <0.1Argent (Ag) mg/Kg Blanc <2

Nom Standard D-076-540Valeur obtenue 33.0

Justesse 95.9%Intervalle 29.2 - 39.6

Arsenic (As) mg/Kg Blanc <0.05Nom Standard D-076-540

Valeur obtenue 75.6Justesse 80%Intervalle 73.2 - 115.8

Baryum (Ba) mg/Kg Blanc <0.01Nom Standard D-076-540

Valeur obtenue 189Justesse 86.8%Intervalle 142 - 192

Bore (B) mg/Kg Blanc <0.01Nom Standard D-076-540


Cadmium (Cd) mg/Kg Blanc <0.005Nom Standard D-076-540

Valeur obtenue 63.1Justesse 95.7%Intervalle 46.9 - 74.1

Chrome (Cr) mg/Kg Blanc <0.05Nom Standard D-076-540


Cobalt (Co) mg/Kg Blanc <0.05Nom Standard D-076-540

Valeur obtenue 111Justesse 91.2%

Certificat contrôle qualité

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Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres

Intervalle 87 - 117Cuivre (Cu) mg/Kg Blanc <5



Fer (Fe) mg/Kg Blanc <0.5Manganèse (Mn) mg/Kg Blanc <0.05

Nom Standard D-076-540Valeur obtenue 328


Mercure (Hg) mg/Kg Blanc <0.1Nom Standard DOLT-4


Mercure (Hg) mg/Kg Nom Standard DORM-3Valeur obtenue 0.40


Molybdene (Mo) mg/Kg Blanc <0.05Nom Standard D-076-540


Nickel (Ni) mg/Kg Blanc <0.05Nom Standard D-076-540


Plomb (Pb) mg/Kg Blanc <0.05Nom Standard D-076-540


Sélénium (Se) mg/Kg Blanc <0.05Nom Standard D-076-540

Valeur obtenue 77.4Justesse 89.6%


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Intervalle 67.0 - 105.8Zinc (Zn) mg/Kg Blanc <0.05




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Lieu de prélèvement : Radisson Heure de prélèvement :

Méthode laboratoire Méthode de référence

M-MET-3.0 MA.200-Mét. 1.2

M-HG-2.0 MA.207-Hg 2.0

M-MET-4.0 EPA-3050b


Informations supplémentaires

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tél.: (819) 797-3222 (298)

fax.: (819) 762-6640






Réseau: 121-131415-00







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Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres Résultats Méthode d'analyse Date d'analyseAluminium (Al) 13916 mg/Kg M-MET-3.0 30 mai 2012Antimoine (Sb) 7.5 mg/Kg M-MET-3.0 30 mai 2012Arsenic (As) 4382(> C) mg/Kg M-MET-3.0 05 juin 2012Baryum (Ba) 99.4(< A) mg/Kg M-MET-3.0 30 mai 2012Bore (B) <0.01 mg/Kg M-MET-3.0 30 mai 2012Cadmium (Cd) 21.30(> C) mg/Kg M-MET-3.0 30 mai 2012Chrome (Cr) 157(A-B) mg/Kg M-MET-3.0 30 mai 2012Cobalt (Co) 16.5(< A) mg/Kg M-MET-3.0 30 mai 2012Cuivre (Cu) 35(< A) mg/Kg M-MET-3.0 30 mai 2012Fer (Fe) 29893 mg/Kg M-MET-3.0 30 mai 2012Manganèse (Mn) 508(< A) mg/Kg M-MET-3.0 30 mai 2012Mercure (Hg) <0.1 mg/Kg M-HG-2.0 05 juin 2012Molybdene (Mo) 6.5(A-B) mg/Kg M-MET-3.0 30 mai 2012Nickel (Ni) 50.2(A-B) mg/Kg M-MET-3.0 30 mai 2012Plomb (Pb) 17.4(< A) mg/Kg M-MET-3.0 30 mai 2012Préparation d'échantillon Sous-traitance\Laboratoire Expert Inc.

Sélénium (Se) 0.13(< A) mg/Kg M-MET-3.0 05 juin 2012Zinc (Zn) 47.7(< A) mg/Kg M-MET-3.0 30 mai 2012Potentiel générateur acide Sous-traitance\Maxxam Analytics Inc

Soufre 1.3 % Sous-traitance\Maxxam Analytics Inc 12 juin 2012Potentiel d'acidité maximal (PA) 39.4 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisation brut (PN) 89 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisaton net (PNN) 49.6 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012% Humidité <0.1 % M-HUM-1.0 29 mai 2012Argent (Ag) <2 mg/Kg M-MET-4.0 30 mai 2012




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Lieu de prélèvement : Radisson Heure de prélèvement :Paramètre Valeur Unité Méthode AccréditationAluminium (Al) 0.6 mg/Kg M-MET-3.0 OuiAntimoine (Sb) 0.1 mg/Kg M-MET-3.0Arsenic (As) 0.05 mg/Kg M-MET-3.0 OuiBaryum (Ba) 0.01 mg/Kg M-MET-3.0 OuiBore (B) 0.01 mg/Kg M-MET-3.0Cadmium (Cd) 0.005 mg/Kg M-MET-3.0 OuiChrome (Cr) 0.05 mg/Kg M-MET-3.0 OuiCobalt (Co) 0.05 mg/Kg M-MET-3.0 OuiCuivre (Cu) 5 mg/Kg M-MET-3.0 OuiFer (Fe) 0.5 mg/Kg M-MET-3.0Manganèse (Mn) 0.05 mg/Kg M-MET-3.0 OuiMercure (Hg) 0.1 mg/Kg M-HG-2.0 OuiMolybdene (Mo) 0.05 mg/Kg M-MET-3.0 OuiNickel (Ni) 0.05 mg/Kg M-MET-3.0 OuiPlomb (Pb) 0.05 mg/Kg M-MET-3.0 OuiSélénium (Se) 0.05 mg/Kg M-MET-3.0 OuiZinc (Zn) 0.05 mg/Kg M-MET-3.0 Oui% Humidité 0.1 % M-HUM-1.0Argent (Ag) 2 mg/Kg M-MET-4.0 Oui


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Mercure (Hg) mg/Kg Blanc <0.1Nom Standard DOLT-4


Mercure (Hg) mg/Kg Nom Standard DORM-3Valeur obtenue 0.40









Valeur obtenue 77.4Justesse 89.6%


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Intervalle 67.0 - 105.8Zinc (Zn) mg/Kg Blanc <0.05




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M-MET-3.0 MA.200-Mét. 1.2

M-HG-2.0 MA.207-Hg 2.0

M-MET-4.0 EPA-3050b



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tél.: (819) 797-3222 (298)

fax.: (819) 762-6640






Réseau: 121-131415-00







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Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres Résultats Méthode d'analyse Date d'analyseAluminium (Al) 22307 mg/Kg M-MET-3.0 30 mai 2012Antimoine (Sb) 7.7 mg/Kg M-MET-3.0 30 mai 2012Arsenic (As) 2050(> C) mg/Kg M-MET-3.0 05 juin 2012Baryum (Ba) 139(< A) mg/Kg M-MET-3.0 30 mai 2012Bore (B) <0.01 mg/Kg M-MET-3.0 30 mai 2012Cadmium (Cd) 11.61(B-C) mg/Kg M-MET-3.0 30 mai 2012Chrome (Cr) 138(A-B) mg/Kg M-MET-3.0 30 mai 2012Cobalt (Co) 32.5(A-B) mg/Kg M-MET-3.0 30 mai 2012Cuivre (Cu) 68(A-B) mg/Kg M-MET-3.0 30 mai 2012Fer (Fe) 62208 mg/Kg M-MET-3.0 30 mai 2012Manganèse (Mn) 1248(B-C) mg/Kg M-MET-3.0 30 mai 2012Molybdene (Mo) 2.9(< A) mg/Kg M-MET-3.0 30 mai 2012Nickel (Ni) 39.8(< A) mg/Kg M-MET-3.0 30 mai 2012Plomb (Pb) 5.8(< A) mg/Kg M-MET-3.0 30 mai 2012Préparation d'échantillon Sous-traitance\Laboratoire Expert Inc.

Sélénium (Se) <0.05(< A) mg/Kg M-MET-3.0 05 juin 2012Zinc (Zn) 8.5(< A) mg/Kg M-MET-3.0 30 mai 2012Potentiel générateur acide Sous-traitance\Maxxam Analytics Inc

Soufre 2.3 % Sous-traitance\Maxxam Analytics Inc 12 juin 2012Potentiel d'acidité maximal (PA) 71.9 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisation brut (PN) 236 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisaton net (PNN) 164 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012% Humidité <0.1 % M-HUM-1.0 29 mai 2012Argent (Ag) <2 mg/Kg M-MET-4.0 30 mai 2012Mercure (Hg) <0.1 mg/Kg M-HG-2.0 05 juin 2012




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Lieu de prélèvement : Radisson Heure de prélèvement :Paramètre Valeur Unité Méthode AccréditationAluminium (Al) 0.6 mg/Kg M-MET-3.0 OuiAntimoine (Sb) 0.1 mg/Kg M-MET-3.0Arsenic (As) 0.05 mg/Kg M-MET-3.0 OuiBaryum (Ba) 0.01 mg/Kg M-MET-3.0 OuiBore (B) 0.01 mg/Kg M-MET-3.0Cadmium (Cd) 0.005 mg/Kg M-MET-3.0 OuiChrome (Cr) 0.05 mg/Kg M-MET-3.0 OuiCobalt (Co) 0.05 mg/Kg M-MET-3.0 OuiCuivre (Cu) 5 mg/Kg M-MET-3.0 OuiFer (Fe) 0.5 mg/Kg M-MET-3.0Manganèse (Mn) 0.05 mg/Kg M-MET-3.0 OuiMolybdene (Mo) 0.05 mg/Kg M-MET-3.0 OuiNickel (Ni) 0.05 mg/Kg M-MET-3.0 OuiPlomb (Pb) 0.05 mg/Kg M-MET-3.0 OuiSélénium (Se) 0.05 mg/Kg M-MET-3.0 OuiZinc (Zn) 0.05 mg/Kg M-MET-3.0 Oui% Humidité 0.1 % M-HUM-1.0Argent (Ag) 2 mg/Kg M-MET-4.0 OuiMercure (Hg) 0.1 mg/Kg M-HG-2.0 Oui


17 avril 2012N/D



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Justesse 97.1%Intervalle 32300 - 43700Duplicata 22307-24289

Antimoine (Sb) mg/Kg Blanc <0.1Duplicata 7.7-12.9

Argent (Ag) mg/Kg Blanc <2Nom Standard D-076-540

Valeur obtenue 33.0Justesse 95.9%Intervalle 29.2 - 39.6Duplicata <2-<2


Valeur obtenue 75.6Justesse 80%Intervalle 73.2 - 115.8Duplicata 2050-2101


Valeur obtenue 189Justesse 86.8%Intervalle 142 - 192Duplicata 139-149


Valeur obtenue 98Justesse 92.5%Intervalle 82 - 130Duplicata <0.01-<0.01


Valeur obtenue 63.1Justesse 95.7%Intervalle 46.9 - 74.1Duplicata 11.61-11.02



17 avril 2012N/D



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Valeur obtenue 63.8Justesse 90.6%Intervalle 59.8 - 81.0Duplicata 138-147


Valeur obtenue 111Justesse 91.2%Intervalle 87 - 117Duplicata 32.5-33.2

Cuivre (Cu) mg/Kg Blanc <5Nom Standard D-076-540

Valeur obtenue 85.0Justesse 93.2%Intervalle 67.7 - 91.5Duplicata 68-74

Fer (Fe) mg/Kg Blanc <0.5Duplicata 62208-67454

Manganèse (Mn) mg/Kg Blanc <0.05Nom Standard D-076-540

Valeur obtenue 328Justesse 84.1%Intervalle 219 - 347Duplicata 1248-1309

Mercure (Hg) mg/Kg Blanc <0.1Nom Standard DORM-3



Valeur obtenue 50.0Justesse 80.1%Intervalle 32.3 - 51.1Duplicata 2.9-2.4




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Duplicata 39.8-40.7Plomb (Pb) mg/Kg Blanc <0.05


Justesse 91.1%Intervalle 71.1 - 112.5Duplicata 5.8-4.9


Valeur obtenue 77.4Justesse 89.6%Intervalle 67.0 - 105.8Duplicata <0.05-<0.05

Zinc (Zn) mg/Kg Blanc <0.05Nom Standard D-076-540

Valeur obtenue 152Justesse 91.4%Intervalle 119 - 161Duplicata 8.5-9.2


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M-MET-3.0 MA.200-Mét. 1.2

M-MET-4.0 EPA-3050b

M-HG-2.0 MA.207-Hg 2.0



17 avril 2012N/D


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F-02-06






tél.: (819) 797-3222 (298)

fax.: (819) 762-6640






Réseau: 121-131415-00







Page 1 de 7

F-02-06



Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres Résultats Méthode d'analyse Date d'analyseAluminium (Al) 24659 mg/Kg M-MET-3.0 30 mai 2012Antimoine (Sb) 10.6 mg/Kg M-MET-3.0 30 mai 2012Arsenic (As) 28.4(A-B) mg/Kg M-MET-3.0 05 juin 2012Baryum (Ba) 177(< A) mg/Kg M-MET-3.0 30 mai 2012Bore (B) <0.01 mg/Kg M-MET-3.0 30 mai 2012Cadmium (Cd) <0.005(< A) mg/Kg M-MET-3.0 30 mai 2012Chrome (Cr) 247(A-B) mg/Kg M-MET-3.0 30 mai 2012Cobalt (Co) 24.9(A-B) mg/Kg M-MET-3.0 30 mai 2012Cuivre (Cu) 48(< A) mg/Kg M-MET-3.0 30 mai 2012Fer (Fe) 37116 mg/Kg M-MET-3.0 30 mai 2012Manganèse (Mn) 360(< A) mg/Kg M-MET-3.0 30 mai 2012Molybdene (Mo) 6.6(A-B) mg/Kg M-MET-3.0 30 mai 2012Nickel (Ni) 86.4(A-B) mg/Kg M-MET-3.0 30 mai 2012Plomb (Pb) 12.2(< A) mg/Kg M-MET-3.0 30 mai 2012Préparation d'échantillon Sous-traitance\Laboratoire Expert Inc.


Soufre 0.27 % Sous-traitance\Maxxam Analytics Inc 12 juin 2012Potentiel d'acidité maximal (PA) 8.4 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisation brut (PN) 46 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisaton net (PNN) 37.6 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012% Humidité <0.1 % M-HUM-1.0 29 mai 2012Argent (Ag) <2 mg/Kg M-MET-4.0 30 mai 2012Mercure (Hg) <0.1 mg/Kg M-HG-2.0 05 juin 2012




17 avril 2012N/D

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Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres% Humidité % Duplicata <0.1-<0.1Aluminium (Al) mg/Kg Blanc <0.6

















Valeur obtenue 111


17 avril 2012N/D



Page 4 de 7

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Cuivre (Cu) mg/Kg Blanc <5Nom Standard D-076-540

















17 avril 2012N/D



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17 avril 2012N/D



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M-MET-3.0 MA.200-Mét. 1.2

M-MET-4.0 EPA-3050b

M-HG-2.0 MA.207-Hg 2.0



17 avril 2012N/D


Page 7 de 7

F-02-06






tél.: (819) 797-3222 (298)

fax.: (819) 762-6640






Réseau: 121-131415-00







Page 1 de 7

F-02-06



Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres Résultats Méthode d'analyse Date d'analyseAluminium (Al) 31985 mg/Kg M-MET-3.0 30 mai 2012Antimoine (Sb) 23.6 mg/Kg M-MET-3.0 30 mai 2012Arsenic (As) 667(> C) mg/Kg M-MET-3.0 05 juin 2012Baryum (Ba) 4.2(< A) mg/Kg M-MET-3.0 30 mai 2012Bore (B) <0.01 mg/Kg M-MET-3.0 30 mai 2012Cadmium (Cd) 2.17(A-B) mg/Kg M-MET-3.0 30 mai 2012Chrome (Cr) 1954(> C) mg/Kg M-MET-3.0 30 mai 2012Cobalt (Co) 67.0(B-C) mg/Kg M-MET-3.0 30 mai 2012Cuivre (Cu) 64(A-B) mg/Kg M-MET-3.0 30 mai 2012Fer (Fe) 58712 mg/Kg M-MET-3.0 30 mai 2012Manganèse (Mn) 1259(B-C) mg/Kg M-MET-3.0 30 mai 2012Molybdene (Mo) 4.3(< A) mg/Kg M-MET-3.0 30 mai 2012Nickel (Ni) 685(> C) mg/Kg M-MET-3.0 30 mai 2012Plomb (Pb) 2.4(< A) mg/Kg M-MET-3.0 30 mai 2012Préparation d'échantillon Sous-traitance\Laboratoire Expert Inc.


Soufre 0.11 % Sous-traitance\Maxxam Analytics Inc 12 juin 2012Potentiel d'acidité maximal (PA) 3.4 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisation brut (PN) 259 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisaton net (PNN) 256 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012% Humidité <0.1 % M-HUM-1.0 29 mai 2012Argent (Ag) <2 mg/Kg M-MET-4.0 30 mai 2012Mercure (Hg) 0.2 mg/Kg M-HG-2.0 05 juin 2012




17 avril 2012N/D

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17 avril 2012N/D



Page 3 de 7

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17 avril 2012N/D



Page 4 de 7

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Valeur obtenue 152


17 avril 2012N/D



Page 5 de 7

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Page 6 de 7

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M-MET-3.0 MA.200-Mét. 1.2

M-MET-4.0 EPA-3050b

M-HG-2.0 MA.207-Hg 2.0



17 avril 2012N/D


Page 7 de 7

F-02-06






tél.: (819) 797-3222 (298)

fax.: (819) 762-6640






Réseau: 121-131415-00







Page 1 de 7

F-02-06



Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres Résultats Méthode d'analyse Date d'analyseAluminium (Al) 25495 mg/Kg M-MET-3.0 30 mai 2012Antimoine (Sb) 0.4 mg/Kg M-MET-3.0 30 mai 2012Arsenic (As) 458(> C) mg/Kg M-MET-3.0 05 juin 2012Baryum (Ba) 191(< A) mg/Kg M-MET-3.0 30 mai 2012Bore (B) <0.01 mg/Kg M-MET-3.0 30 mai 2012Cadmium (Cd) 1.97(A-B) mg/Kg M-MET-3.0 30 mai 2012Chrome (Cr) 220(A-B) mg/Kg M-MET-3.0 30 mai 2012Cobalt (Co) 25.0(A-B) mg/Kg M-MET-3.0 30 mai 2012Cuivre (Cu) 61(A-B) mg/Kg M-MET-3.0 30 mai 2012Fer (Fe) 47190 mg/Kg M-MET-3.0 30 mai 2012Manganèse (Mn) 663(< A) mg/Kg M-MET-3.0 30 mai 2012Molybdene (Mo) 6.2(A-B) mg/Kg M-MET-3.0 30 mai 2012Nickel (Ni) 62.5(A-B) mg/Kg M-MET-3.0 30 mai 2012Plomb (Pb) 8.1(< A) mg/Kg M-MET-3.0 30 mai 2012Préparation d'échantillon Sous-traitance\Laboratoire Expert Inc.


Soufre 0.61 % Sous-traitance\Maxxam Analytics Inc 12 juin 2012Potentiel d'acidité maximal (PA) 19.1 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisation brut (PN) 29 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012Potentiel neutralisaton net (PNN) 9.90 kg CaCO3/t Sous-traitance\Maxxam Analytics Inc 08 juin 2012% Humidité <0.1 % M-HUM-1.0 29 mai 2012Argent (Ag) <2 mg/Kg M-MET-4.0 30 mai 2012Mercure (Hg) <0.1 mg/Kg M-HG-2.0 05 juin 2012




17 avril 2012N/D

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17 avril 2012N/D



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17 avril 2012N/D



Page 4 de 7

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Valeur obtenue 152


17 avril 2012N/D



Page 5 de 7

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17 avril 2012N/D



Page 6 de 7

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M-MET-3.0 MA.200-Mét. 1.2

M-MET-4.0 EPA-3050b

M-HG-2.0 MA.207-Hg 2.0



17 avril 2012N/D


Page 7 de 7

F-02-06






tél.: (819) 797-3222 (298)

fax.: (819) 762-6640






Réseau: 121-131415-00







Page 1 de 7

F-02-06



Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres Résultats Méthode d'analyse Date d'analyseAluminium (Al) 22364 mg/Kg M-MET-3.0 30 mai 2012Antimoine (Sb) 9.7 mg/Kg M-MET-3.0 30 mai 2012Arsenic (As) 199(> C) mg/Kg M-MET-3.0 05 juin 2012Baryum (Ba) 301(A-B) mg/Kg M-MET-3.0 30 mai 2012Bore (B) <0.01 mg/Kg M-MET-3.0 30 mai 2012Cadmium (Cd) 0.341(< A) mg/Kg M-MET-3.0 30 mai 2012Chrome (Cr) 387(B-C) mg/Kg M-MET-3.0 30 mai 2012Cobalt (Co) 22.4(A-B) mg/Kg M-MET-3.0 30 mai 2012Cuivre (Cu) 33(< A) mg/Kg M-MET-3.0 30 mai 2012Fer (Fe) 37175 mg/Kg M-MET-3.0 30 mai 2012Manganèse (Mn) 643(< A) mg/Kg M-MET-3.0 30 mai 2012Molybdene (Mo) 2.4(< A) mg/Kg M-MET-3.0 30 mai 2012Nickel (Ni) 112(B-C) mg/Kg M-MET-3.0 30 mai 2012Plomb (Pb) 6.3(< A) mg/Kg M-MET-3.0 30 mai 2012Préparation d'échantillon Sous-traitance\Laboratoire Expert Inc.






17 avril 2012N/D

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17 avril 2012N/D



Page 3 de 7

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17 avril 2012N/D



Page 4 de 7

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Valeur obtenue 152


17 avril 2012N/D



Page 5 de 7

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17 avril 2012N/D



Page 6 de 7

F-02-06





M-MET-3.0 MA.200-Mét. 1.2

M-MET-4.0 EPA-3050b

M-HG-2.0 MA.207-Hg 2.0



17 avril 2012N/D


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F-02-06






tél.: (819) 797-3222 (298)

fax.: (819) 762-6640


Lieu de prélèvement : Radisson Date de prélèvement : 17 mai 2012




Réseau: 121-131415-00







Page 1 de 7

F-02-06



Lieu de prélèvement : Radisson Heure de prélèvement :Paramètres Résultats Méthode d'analyse Date d'analyseAluminium (Al) 40183 mg/Kg M-MET-3.0 30 mai 2012Antimoine (Sb) 10.1 mg/Kg M-MET-3.0 30 mai 2012Arsenic (As) 16.2(A-B) mg/Kg M-MET-3.0 05 juin 2012Baryum (Ba) 8.6(< A) mg/Kg M-MET-3.0 30 mai 2012Bore (B) <0.01 mg/Kg M-MET-3.0 30 mai 2012Cadmium (Cd) <0.005(< A) mg/Kg M-MET-3.0 30 mai 2012Chrome (Cr) 173(A-B) mg/Kg M-MET-3.0 30 mai 2012Cobalt (Co) 45.7(A-B) mg/Kg M-MET-3.0 30 mai 2012Cuivre (Cu) 91(A-B) mg/Kg M-MET-3.0 30 mai 2012Fer (Fe) 74008 mg/Kg M-MET-3.0 30 mai 2012Manganèse (Mn) 1124(B-C) mg/Kg M-MET-3.0 30 mai 2012Molybdene (Mo) 7.6(A-B) mg/Kg M-MET-3.0 30 mai 2012Nickel (Ni) 55.9(A-B) mg/Kg M-MET-3.0 30 mai 2012Plomb (Pb) 7.6(< A) mg/Kg M-MET-3.0 30 mai 2012Préparation d'échantillon Sous-traitance\Laboratoire Expert Inc.

Sélénium (Se) <0.05(< A) mg/Kg M-MET-3.0 05 juin 2012Zinc (Zn) 106(< A) mg/Kg M-MET-3.0 30 mai 2012Potentiel générateur acide Sous-traitance\Maxxam Analytics Inc





17 mai 2012N/D

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17 mai 2012N/D



Page 3 de 7

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17 mai 2012N/D



Page 4 de 7

F-02-06





















Valeur obtenue 152


17 mai 2012N/D



Page 5 de 7

F-02-06






17 mai 2012N/D



Page 6 de 7

F-02-06





M-MET-3.0 MA.200-Mét. 1.2

M-MET-4.0 EPA-3050b

M-HG-2.0 MA.207-Hg 2.0



17 mai 2012N/D


Page 7 de 7

F-02-06


Sommaire des résultatsClient : Genivar Inc

Responsable : Mme Marie-Élise Viger Date de réception : 18 mai 2012Adresse : 152, avenue Murdoch Nom du préleveur : Eugène Gauthier

Rouyn-Noranda Québec J9X 1E1 Type d'échantillon : Mineraitél.: (819) 797-3222 (298)

fax.: (819) 762-6640

No Multilab Direct 113485 113487 113489 113491 113493 113495 113497 113499Échantillon 136099 136100 136101 136102 136103 136104 136105 136106Date prélèvement 17-04-2012 17-04-2012 17-04-2012 17-04-2012 17-04-2012 17-04-2012 17-04-2012 17-05-2012Aluminium (Al) mg/L 0.425 0.032 <0.006 <0.006 <0.006 <0.006 <0.006 0.344 Antimoine (Sb) mg/L 0.0030 0.0076 <0.0001 <0.0001 <0.0001 <0.0001 <0.0001 0.0011 Argent (Ag) mg/L <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 Arsenic (As) mg/L 0.5905 0.1070 0.1591 <0.0005 0.0501 0.2080 <0.0005 <0.0005 Baryum (Ba) mg/L 0.1302 0.1285 0.3346 0.2065 0.0365 0.2770 0.4235 0.0190 Bore (lixiviation) mg/L <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 Cadmium (Cd) mg/L 0.0003 0.0001 <0.0001 <0.0001 <0.0001 <0.0001 <0.0001 0.0010 Chrome (Cr) mg/L 0.0225 0.0066 <0.0006 <0.0006 <0.0006 <0.0006 <0.0006 0.0238 Cobalt (Co) mg/L 0.004 0.003 <0.001 <0.001 0.009 <0.001 <0.001 0.008 Cuivre (Cu) mg/L 0.0023 0.0060 0.0019 0.0196 0.0070 0.0197 0.0202 0.0037 Fer (Fe) mg/L 6.4 2.0 19.3 2.4 40.9 4.3 4.4 13.8 Lixiviation (TCLP) Manganèse (Mn) mg/L 8.387 7.702 18.85 1.958 12.72 3.140 4.865 16.01 Mercure (Hg) mg/L <0.00002 <0.00002 <0.00002 <0.00002 <0.00002 <0.00002 <0.00002 <0.00002 Molybdene (Mo) mg/L 0.0029 0.0025 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 Nickel (Ni) mg/L 0.0638 0.0776 0.0285 0.0148 0.1271 0.0163 0.0029 0.0660 Plomb (Pb) mg/L 0.0353 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 Sélénium (Se) mg/L 0.001 <0.001 <0.001 <0.001 <0.001 <0.001 <0.001 <0.001 Zinc (Zn) mg/L 0.039 0.030 <0.001 <0.001 <0.001 <0.001 <0.001 0.017 % Humidité % <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1


Ces résultats se rapportent à ceux inscrits sur le(s) Certificat(s) d'analyse correspondant(s) au numéro de projet.

En cas de différence entre ces documents, les résultats du(des) Certificat(s) d'analyse, dûment signé(s),

ont préséance sur ceux de ce sommaire des résultats.

Page 1 de 4

F-02-13Version 3ième: 30/03/2011




fax.: (819) 762-6640

No Multilab Direct 113485 113487 113489 113491 113493 113495 113497 113499Échantillon 136099 136100 136101 136102 136103 136104 136105 136106Date prélèvement 17-04-2012 17-04-2012 17-04-2012 17-04-2012 17-04-2012 17-04-2012 17-04-2012 17-05-2012pH 5.53 6.35 5.67 5.47 5.02 5.26 6.2 5.73





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F-02-13Version 3ième: 30/03/2011




fax.: (819) 762-6640

ParamètresAluminium (Al) 0.006 mg/L M-LIX-1.0Antimoine (Sb) 0.0001 mg/L M-LIX-1.0Argent (Ag) 0.0005 mg/L M-LIX-1.0Arsenic (As) 0.0005 mg/L M-LIX-1.0Baryum (Ba) 0.0005 mg/L M-LIX-1.0Bore (lixiviation) 0.01 mg/L M-LIX-1.0Cadmium (Cd) 0.0001 mg/L M-LIX-1.0Chrome (Cr) 0.0006 mg/L M-LIX-1.0Cobalt (Co) 0.001 mg/L M-LIX-1.0Cuivre (Cu) 0.0005 mg/L M-LIX-1.0Fer (Fe) 0.01 mg/L M-LIX-1.0Lixiviation (TCLP) N.D. M-LIX-1.0Manganèse (Mn) 0.0005 mg/L M-LIX-1.0Mercure (Hg) 0.00002 mg/L M-LIX-1.0Molybdene (Mo) 0.0005 mg/L M-LIX-1.0Nickel (Ni) 0.0005 mg/L M-LIX-1.0Plomb (Pb) 0.0005 mg/L M-LIX-1.0Sélénium (Se) 0.001 mg/L M-LIX-1.0Zinc (Zn) 0.001 mg/L M-LIX-1.0% Humidité 0.1 % M-HUM-1.0






Valeur

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F-02-13Version 3ième: 30/03/2011




fax.: (819) 762-6640

ParamètrespH N.D. M-LIX-1.0




ValeurLimite de détection rapportée


Page 4 de 4

F-02-13Version 3ième: 30/03/2011


Client : Genivar Inc Numéro de projet : MultipleResponsable : Mme Marie-Élise Viger Date de réception : 18 mai 2012

Adresse : 152, avenue Murdoch Nom du préleveur : Eugène GauthierRouyn-Noranda Québec J9X 1E1 Type d'échantillon : Minerai

tél.: (819) 797-3222 (298)fax.: (819) 762-6640

Paramètres Blanc Nom Obtenue Intervalle 1 2

Aluminium (Al) mg/L <0.006Aluminium (Al) mg/L <0.006 <0.006 <0.006Antimoine (Sb) mg/L <0.0001Antimoine (Sb) mg/L <0.0001 <0.0001 <0.0001Argent (Ag) mg/L <0.0005 <0.0005 <0.0005Argent (Ag) mg/L <0.0005Arsenic (As) mg/L <0.0005 0.1591 0.1246Arsenic (As) mg/L <0.0005Baryum (Ba) mg/L <0.0005 0.3346 0.3346Baryum (Ba) mg/L <0.0005Bore (lixiviation) mg/L <0.01 <0.01 <0.01Bore (lixiviation) mg/L <0.01Cadmium (Cd) mg/L <0.0001Cadmium (Cd) mg/L <0.0001 <0.0001 <0.0001Chrome (Cr) mg/L <0.0006Chrome (Cr) mg/L <0.0006 <0.0006 <0.0006Cobalt (Co) mg/L <0.001Cobalt (Co) mg/L <0.001 <0.001 <0.001Cuivre (Cu) mg/L <0.0005 0.0019 0.0013Cuivre (Cu) mg/L <0.0005Projet: 113485,113487,113489,113491,113493,113495,113497,113499


Duplicata




Standard

Page 1 de 2

F-02-15Version 3ième: 17/11/2011


Client : Genivar Inc Numéro de projet : MultipleResponsable : Mme Marie-Élise Viger Date de réception : 18 mai 2012


tél.: (819) 797-3222 (298)fax.: (819) 762-6640


Fer (Fe) mg/L <0.01 19.3 19.3Fer (Fe) mg/L <0.01Manganèse (Mn) mg/L <0.0005 18.85 18.85Manganèse (Mn) mg/L <0.0005Mercure (Hg) mg/L <0.00002Mercure (Hg) mg/L <0.00002 <0.00002 <0.00002Molybdene (Mo) mg/L <0.0005 <0.0005 <0.0005Molybdene (Mo) mg/L <0.0005Nickel (Ni) mg/L <0.0005 0.0285 0.0285Nickel (Ni) mg/L <0.0005pH 5.67 5.67pH Plomb (Pb) mg/L <0.0005Plomb (Pb) mg/L <0.0005 <0.0005 <0.0005Sélénium (Se) mg/L <0.001Sélénium (Se) mg/L <0.001 <0.001 <0.001Zinc (Zn) mg/L <0.001 <0.001 <0.001Zinc (Zn) mg/L <0.001

Projet: 113485,113487,113489,113491,113493,113495,113497,113499

Date d'émission : 05 juin 2012Les échantillons seront conservés pendant 30 jours à partir de la date du rapport à moins d'avis écrit du client.

Standard Duplicata



Page 2 de 2

F-02-15Version 3ième: 17/11/2011


Responsable : Mme Marie-Élise Viger Date de réception : 22 juin 2012Adresse : 152, avenue Murdoch Nom du préleveur : Eugène Gauthier


fax.: (819) 762-6640

No Multilab Direct 115747 115748 115749 115750 115751 115752 115753 115754Échantillon 136107 136108 136109 136110 136111 136112 136113 136114Date prélèvement 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012% Humidité % <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 Aluminium (Al) mg/Kg 21085 16831 16955 15252 19168 18404 14402 16620 Antimoine (Sb) mg/Kg 0.2 <0.1 2.0 1.8 1.0 1.4 0.5 2.0 Arsenic (As) mg/Kg <0.05 (< A) 15.2 (A-B) 29.5 (A-B) 18144 (> C) 17737 (> C) 13615 (> C) 13833 (> C) 11659 (> C)Baryum (Ba) mg/Kg 35.4 (< A) <0.01 (< A) 122 (< A) 69.4 (< A) 188 (< A) 128 (< A) 100 (< A) 162 (< A)Bore (B) mg/Kg <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 Cadmium (Cd) mg/Kg 8.07 (B-C) 0.268 (< A) 0.158 (< A) 0.182 (< A) 9.90 (B-C) 0.122 (< A) 0.145 (< A) 0.127 (< A)Chrome (Cr) mg/Kg 125 (A-B) 81.2 (< A) 170 (A-B) 82.3 (< A) 116 (A-B) 81.2 (< A) 74.6 (< A) 136 (A-B)Cobalt (Co) mg/Kg 9.3 (< A) 33.0 (A-B) 5.0 (< A) 22.2 (A-B) 25.1 (A-B) 6.6 (< A) 4.5 (< A) 14.2 (< A)Cuivre (Cu) mg/Kg 23 (< A) 146 (B-C) 19 (< A) 97 (A-B) 52 (A-B) 52 (A-B) 60 (A-B) 42 (< A)Fer (Fe) mg/Kg 47093 60180 37423 68876 69851 47200 47263 55753 Manganèse (Mn) mg/Kg 473 (< A) 1373 (B-C) 385 (< A) 1384 (B-C) 1376 (B-C) 802 (< A) 987 (< A) 1038 (B-C)Molybdene (Mo) mg/Kg 2.6 (< A) 1.0 (< A) 1.5 (< A) 0.92 (< A) 1.0 (< A) 1.9 (< A) 1.0 (< A) 1.0 (< A)Nickel (Ni) mg/Kg 88.5 (A-B) 53.4 (A-B) 84.0 (A-B) 46.2 (< A) 42.9 (< A) 37.2 (< A) 34.0 (< A) 44.2 (< A)Plomb (Pb) mg/Kg <0.05 (< A) <0.05 (< A) <0.05 (< A) <0.05 (< A) <0.05 (< A) <0.05 (< A) <0.05 (< A) <0.05 (< A)Potentiel générateur acide Potentiel neutralisaton net (PNN) kg CaCO3/t 28.7 285 18.0 87.3 141 33.7 74.9 120 Potentiel neutralisation brut (PN) kg CaCO3/t 34 314 24 157 207 83 126 177 Potentiel d'acidité maximal (PA) kg CaCO3/t 5.3 19 6 69.7 66.3 49.3 51.1 57.5 Soufre % 0.17 0.61 0.19 2.2 2.1 1.6 1.6 1.8

Date d'émission : 17 juillet 2012




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F-02-13Version 3ième: 30/03/2011




fax.: (819) 762-6640

No Multilab Direct 115747 115748 115749 115750 115751 115752 115753 115754Échantillon 136107 136108 136109 136110 136111 136112 136113 136114Date prélèvement 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012Préparation d'échantillon Sélénium (Se) mg/Kg <0.05 (< A) <0.05 (< A) <0.05 (< A) <0.05 (< A) <0.05 (< A) 5.6 (B-C) <0.05 (< A) <0.05 (< A)Zinc (Zn) mg/Kg 57.8 (< A) 10.4 (< A) 71.2 (< A) 64.8 (< A) <0.05 (< A) 38.2 (< A) 33.0 (< A) 8.1 (< A)Argent (Ag) mg/Kg <2 <2 <2 <2 <2 <2 <2 2 Mercure (Hg) mg/Kg <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1





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F-02-13Version 3ième: 30/03/2011




fax.: (819) 762-6640

Paramètres% Humidité 0.1 % M-HUM-1.0Aluminium (Al) 0.6 mg/Kg M-MET-3.0Antimoine (Sb) 0.1 mg/Kg M-MET-3.0Arsenic (As) 0.05 mg/Kg M-MET-3.0Baryum (Ba) 0.01 mg/Kg M-MET-3.0Bore (B) 0.01 mg/Kg M-MET-3.0Cadmium (Cd) 0.005 mg/Kg M-MET-3.0Chrome (Cr) 0.05 mg/Kg M-MET-3.0Cobalt (Co) 0.05 mg/Kg M-MET-3.0Cuivre (Cu) 5 mg/Kg M-MET-3.0Fer (Fe) 0.5 mg/Kg M-MET-3.0Manganèse (Mn) 0.05 mg/Kg M-MET-3.0Molybdene (Mo) 0.05 mg/Kg M-MET-3.0Nickel (Ni) 0.05 mg/Kg M-MET-3.0Plomb (Pb) 0.05 mg/Kg M-MET-3.0Sélénium (Se) 0.05 mg/Kg M-MET-3.0Zinc (Zn) 0.05 mg/Kg M-MET-3.0Argent (Ag) 2 mg/Kg M-MET-4.0Mercure (Hg) 0.1 mg/Kg M-HG-2.0






Page 3 de 3

F-02-13Version 3ième: 30/03/2011


Client : Genivar Inc Numéro de projet : MultipleResponsable : Mme Marie-Élise Viger Date de réception : 22 juin 2012


tél.: (819) 797-3222 (298)fax.: (819) 762-6640


Aluminium (Al) mg/Kg <0.6 D-076-540 7551 7140 - 9660Antimoine (Sb) mg/Kg <0.1Argent (Ag) mg/Kg <2 D-076-540 38.0 29.2 - 39.6Arsenic (As) mg/Kg <0.05 D-076-540 94.6 73.2 - 115.8Baryum (Ba) mg/Kg <0.01 D-076-540 187 142 - 192Bore (B) mg/Kg <0.01 D-076-540 127 82 - 130Cadmium (Cd) mg/Kg <0.005 D-076-540 66.1 46.9 - 74.1Chrome (Cr) mg/Kg <0.05 D-076-540 60.3 59.8 - 81.0Cobalt (Co) mg/Kg <0.05 D-076-540 108 87 - 117Cuivre (Cu) mg/Kg <5 D-076-540 73.0 67.7 - 91.5Fer (Fe) mg/Kg <0.5 D-076-540 11966 9688 - 15313Manganèse (Mn) mg/Kg <0.05 D-076-540 302 219 - 347Mercure (Hg) mg/Kg <0.1Molybdene (Mo) mg/Kg <0.05 D-076-540 36.9 32.3 - 51.1Nickel (Ni) mg/Kg <0.05 D-076-540 47.3 44.6 - 70.6Plomb (Pb) mg/Kg <0.05 D-076-540 79.0 71.1 - 112.5Sélénium (Se) mg/Kg <0.05 D-076-540 90.5 67.0 - 105.8Zinc (Zn) mg/Kg <0.05 D-076-540 143 119 - 161

Projet: 115747:115754


Duplicata




Standard

Page 1 de 1

F-02-15Version 3ième: 17/11/2011




fax.: (819) 762-6640

No Multilab Direct 115755 115756 115757 115758 115759 115760 115761 115762Échantillon 136107 136108 136109 136110 136111 136112 136113 136114Date prélèvement 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012 22-06-2012Aluminium (Al) mg/L 0.703 0.030 1.95 0.219 0.038 <0.006 1.66 0.189 Antimoine (Sb) mg/L 0.0025 0.0003 0.0014 0.0072 0.0074 0.0040 0.0063 0.0081 Argent (Ag) mg/L <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 <0.0005 Arsenic (As) mg/L <0.0005 <0.0005 <0.0005 1.504 0.9565 0.3325 3.956 1.086 Baryum (Ba) mg/L 0.0683 0.0275 0.1479 0.1896 0.3855 0.2319 0.2770 0.3477 Bore (lixiviation) mg/L <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 Cadmium (Cd) mg/L <0.0001 0.0012 <0.0001 0.0007 0.0008 0.0005 0.0013 0.0005 Chrome (Cr) mg/L 0.0360 0.0293 0.0421 0.0242 0.0296 0.0289 0.0646 0.0372 Cobalt (Co) mg/L 0.006 0.026 0.004 0.025 0.022 0.009 0.027 0.024 Cuivre (Cu) mg/L 0.0239 0.0014 0.0254 0.0019 0.0012 0.0226 0.0017 0.0012 Fer (Fe) mg/L 7.3 11.1 7.5 13.3 12.7 1.1 23.1 13.5 Lixiviation (TCLP) Manganèse (Mn) mg/L 0.9857 12.30 0.7625 17.70 19.98 8.891 21.35 17.14 Mercure (Hg) mg/L <0.00002 <0.00002 <0.00002 <0.00002 <0.00002 <0.00002 <0.00002 <0.00002 Molybdene (Mo) mg/L 0.0021 0.0007 0.0022 <0.0005 <0.0005 0.0021 0.0009 <0.0005 Nickel (Ni) mg/L 0.1142 0.4630 0.0573 0.4904 0.4821 0.1917 0.4249 0.4978 Plomb (Pb) mg/L 0.0118 0.0013 0.0013 0.0152 0.0073 0.0035 0.0281 0.0110 Sélénium (Se) mg/L <0.001 <0.001 <0.001 <0.001 <0.001 <0.001 <0.001 <0.001 Zinc (Zn) mg/L <0.001 0.004 0.02 0.034 0.015 0.032 0.037 0.009

Certificat corrigé, remplace le certificat multiple émis le 04 juillet 2012.





Page 1 de 2

F-02-13Version 3ième: 30/03/2011




fax.: (819) 762-6640

ParamètresAluminium (Al) 0.006 mg/L M-LIX-1.0Antimoine (Sb) 0.0001 mg/L M-LIX-1.0Argent (Ag) 0.0005 mg/L M-LIX-1.0Arsenic (As) 0.0005 mg/L M-LIX-1.0Baryum (Ba) 0.0005 mg/L M-LIX-1.0Bore (lixiviation) 0.01 mg/L M-LIX-1.0Cadmium (Cd) 0.0001 mg/L M-LIX-1.0Chrome (Cr) 0.0006 mg/L M-LIX-1.0Cobalt (Co) 0.001 mg/L M-LIX-1.0Cuivre (Cu) 0.0005 mg/L M-LIX-1.0Fer (Fe) 0.01 mg/L M-LIX-1.0Lixiviation (TCLP) N.D. M-LIX-1.0Manganèse (Mn) 0.0005 mg/L M-LIX-1.0Mercure (Hg) 0.00002 mg/L M-LIX-1.0Molybdene (Mo) 0.0005 mg/L M-LIX-1.0Nickel (Ni) 0.0005 mg/L M-LIX-1.0Plomb (Pb) 0.0005 mg/L M-LIX-1.0Sélénium (Se) 0.001 mg/L M-LIX-1.0Zinc (Zn) 0.001 mg/L M-LIX-1.0

Certificat corrigé, remplace le certificat multiple émis le 04 juillet 2012.






Page 2 de 2

F-02-13Version 3ième: 30/03/2011


Client : Genivar Inc Numéro de projet : MultipleResponsable : Mme Marie-Élise Viger Date de réception : 22 juin 2012


tél.: (819) 797-3222 (298)fax.: (819) 762-6640


Aluminium (Al) mg/L <0.006Antimoine (Sb) mg/L <0.0001Argent (Ag) mg/L <0.0005Arsenic (As) mg/L <0.0005Baryum (Ba) mg/L <0.0005Bore (lixiviation) mg/L <0.01Cadmium (Cd) mg/L <0.0001Chrome (Cr) mg/L <0.0006Cobalt (Co) mg/L <0.001Cuivre (Cu) mg/L <0.0005Fer (Fe) mg/L <0.01Manganèse (Mn) mg/L <0.0005Mercure (Hg) mg/L <0.00002Molybdene (Mo) mg/L <0.0005Nickel (Ni) mg/L <0.0005Plomb (Pb) mg/L <0.0005Sélénium (Se) mg/L <0.001Zinc (Zn) mg/L <0.001

Projet: 115755:115762Certificat corrigé, remplace le certificat multiple émis le 04 juillet 2012.


Duplicata




Standard

Page 1 de 1

F-02-15Version 3ième: 17/11/2011

Annexe C Exigences de rejet – Directive 019

Paramètre

Colonne I Concentration moyenne

acceptable (moyenne arithmétique

mensuelle)

Colonne II Concentration maximale

acceptable dans un échantillon instantané

Arsenic extractible

Cuivre extractible

Fer extractible

Nickel extractible

Plomb extractible

Zinc extractible

Cyanures totaux

Hydrocarbures (C10C50)

Matières en suspension

0,200 mg/l

0,300 mg/l

3,000 mg/l

0,500 mg/l

0,200 mg/l

0,500 mg/l

1,000 mg/l

---------

15,000 mg/l

0,400 mg/l

0,600 mg/l

6,000 mg/l

1,000 mg/l

0,400 mg/l

1,000 mg/l

2,000 mg/l

2,000 mg/l

30,000 mg/l

* Selon la nature du minerai, du procédé, des résidus miniers ou du calcul des objectifs environnementaux de rejet, d’autres exigences au point de déversement de l’effluent final pourraient s’ajouter en vertu de l’article 20 de la Loi de la délivrance du certificat d’autorisation. Source : Directive 019 sur l’industrie minière

43-101 technical report for pea of the o'brien project

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