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InnovExplo Inc.
WSP Canada 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
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
www.innovexplo.com
TABLE OF CONTENTS
SIGNATURE PAGE – INNOVEXPLO ........................................................................................................ 10
SIGNATURE PAGE – INNOVEXPLO ........................................................................................................ 11
SIGNATURE PAGE – WSP CANADA INC. ............................................................................................... 12
SIGNATURE PAGE – WSP CANADA INC. ............................................................................................... 13
SIGNATURE PAGE – WSP CANADA INC. ............................................................................................... 14
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
7.3.2.2
7.3.2.3
7.3.2.4
Porphyritic andesite ....................................................................................................................... 79
Conglomerate ................................................................................................................................ 80
Volcanic rocks ................................................................................................................................ 80
Graphitic schist .............................................................................................................................. 80
7.4.1.1
7.4.1.2
7.4.1.3
No. 1 Vein ...................................................................................................................................... 81
No. 4 Vein ...................................................................................................................................... 81
No. 9 Vein ...................................................................................................................................... 81
Pontiac Group ...................................................................................................................... 80
Mineralization ............................................................................................................................. 80
O’Brien mine ........................................................................................................................ 80
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
21.1.6.2
21.1.6.3
Site preparation and installation ................................................................................................... 158
Buildings ...................................................................................................................................... 159
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
43-101 Technical Report – O’Brien Project
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SIGNATURE PAGE – INNOVEXPLO
TECHNICAL REPORT FOR THE O’BRIEN PROJECT,
ABITIBI, QUÉBEC
(according to National Instrument 43-101 and Form 43-101F1)
Prepared for
Radisson Mining Resources Inc.
C.P. 307
Rouyn-Noranda, Québec
Canada J9X 5C3
Original signed and sealed (“Sylvie Poirier”)
Sylvie Poirier, Eng.
InnovExplo – Consulting Firm
Longueuil (Québec)
Signed at Longueuil on January 29, 2016
SIGNATURE PAGE – INNOVEXPLO
TECHNICAL REPORT FOR THE O’BRIEN PROJECT,
ABITIBI, QUÉBEC
(according to National Instrument 43-101 and Form 43-101F1)
Prepared for
Radisson Mining Resources Inc.
C.P. 307
Rouyn-Noranda, Québec
Canada J9X 5C3
Original signed and sealed (“Pierre-Luc Richard”)
Signed at Val-d’Or on January 29, 2016
Original signed and sealed (“Bruno Turcotte”)
Signed at Val-d’Or on January 29, 2016
Original signed and sealed (“Laurent Roy”)
Signed at Val-d’Or on January 29, 2016
Pierre-Luc Richard, M.Sc., P.Geo..
InnovExplo – Consulting Firm
Val-d’Or (Québec)
Bruno Turcotte, P.Geo.
InnovExplo – Consulting Firm
Val-d’Or (Québec)
Laurent Roy, Eng.
InnovExplo – Consulting Firm
Val-d’Or (Québec)
SIGNATURE PAGE – WSP CANADA INC.
TECHNICAL REPORT FOR THE O’BRIEN PROJECT,
ABITIBI, QUÉBEC
(according to National Instrument 43-101 and Form 43-101F1)
Prepared for
Radisson Mining Resources Inc.
C.P. 307
Rouyn-Noranda, Québec
Canada J9X 5C3
Original signed and sealed (“Annie Lavoie”)
Annie Lavoie. Eng.
WSP Canada inc.
Montréal (Québec)
Signed at Montreal on January 29, 2016
SIGNATURE PAGE – WSP CANADA INC.
TECHNICAL REPORT FOR THE O’BRIEN PROJECT,
ABITIBI, QUÉBEC
(according to National Instrument 43-101 and Form 43-101F1)
Prepared for
Radisson Mining Resources Inc.
C.P. 307
Rouyn-Noranda, Québec
Canada J9X 5C3
Original signed and sealed (“Éric Poirier”)
Éric Poirier, Eng.
WSP Canada inc.
Val-d’Or (Québec)
Signed at Val-d’Or on January 29, 2016
SIGNATURE PAGE – WSP CANADA INC.
TECHNICAL REPORT FOR THE O’BRIEN PROJECT,
ABITIBI, QUÉBEC
(according to National Instrument 43-101 and Form 43-101F1)
Prepared for
Radisson Mining Resources Inc.
C.P. 307
Rouyn-Noranda, Québec
Canada J9X 5C3
Original signed and sealed (“Marie-Claude Dion StPierre”)
Marie-Claude Dion St-Pierre, Eng
WSP Canada inc.
Quebec (Québec)
Signed at Québec on January 29, 2016
SIGNATURE PAGE – LAMONT INC.
TECHNICAL REPORT FOR THE O’BRIEN PROJECT,
ABITIBI, QUÉBEC
(according to National Instrument 43-101 and Form 43-101F1)
Prepared for
Radisson Mining Resources Inc.
C.P. 307
Rouyn-Noranda, Québec
Canada J9X 5C3
Original signed and sealed (“Ann Lamontagne”)
Ann Lamontagne, Eng., Ph.D.
Lamont inc.
Lac-Beauport (Québec)
Signed at Lac-Beauport on January 29, 2016
www.innovexplo.com
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.
www.innovexplo.com
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
www.innovexplo.com
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
www.innovexplo.com
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 43101) 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 43101F1) (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 43101F1, 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.
www.innovexplo.com
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.
4.
I am a member of the Ordre des Ingénieurs du Québec (OIQ #124421).
5.
6.
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.
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.
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).
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.
10. I have read Regulation 43-101 respecting standards of disclosure for mineral projects and Form 43101F1, 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)
Annie Lavoie, Eng.
WSP
www.innovexplo.com
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.
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 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.
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 43101F1, 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)
Éric Poirier, Eng.
WSP Canada inc.
www.innovexplo.com
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.
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 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 43101F1) (the “Technical Report”), effective as of November 29, 2015 and dated January 29, 2016,
prepared for Radisson Mining Resources Inc.
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 43101F1, 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)
Marie-Claude Dion St-Pierre, Eng. M.A.Sc.
WSP Canada inc.
www.innovexplo.com
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.
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 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.
www.innovexplo.com
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 Vald’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.
43-101 Technical Report – O’Brien Project
24
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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-talccarbonate 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 highgrade 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
43-101 Technical Report – O’Brien Project
25
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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 cutoff 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.
43-101 Technical Report – O’Brien Project
26
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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)
Indicated
Zone
All
Zones
















Cut-off
Tonnage
Inferred
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
All
Zones
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.
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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
48 km
Closed
NA
SigmaLamaque
Complex
Camflo Mill
Integra Gold
Gravity Concentration
& Leaching/CIP
1,000 to
2,200 tpd
1,200 to
2,400 tpd
67 km
Closed
No interest
Richmont
Mines
Iamgold
Leaching/MerrillCrowe
Gravity Concentration
& Leaching/CIP
Or
Gravity Concentration
& Flotation
Gravity Concentration
& Flotation & Flotation
Concentrate Leaching
800 to
1,200 tpd
2,400 tpd
35 km
Operating
No interest
19 km
Operating
Yes
75 km
Operating
Yes with
environmental
conditions
Westwood
Mill
Aurbel Mill
QMX
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800 tpd
500 to
800 tpd
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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
C$/t
289
280
Preparation and trucking
$/t
5.78
5.78
Custom milling
$/t
31
31
Smelting
$/t
45
NA
Total
$/t
81.78
36.78
Total3
Milling cost
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/2012147) 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, longhole mining with captive sublevels was selected.
The mineralized material will be transported to surface using a combination of 3.5cubic-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)
Production (t)
Grade (g/t)
Development (t)
Grade (g/t)
Total tonnage milled (t)
Grade (g/t)
Pre-production
Year 1
Year 2
Year 3
33,194 126,494
7.20
7.05
3,196
33,474 32,080
7.05
5.74
6.19
3,196
66,668 158,574
7.05
6.47
6.87
Production
Year 4
Year 5
129,593 134,524
7.39
5.66
40,298
52,409
5.95
5.11
169,891 186,933
7.04
5.50
Year 6
127,259
6.53
127,259
6.53
Total
551,064
6.68
161,457
5.70
712,521
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
Capitalized operating costs
Capitalized revenue
Pre-production
$21.33 M
-$19.30 M
Royalty payment
Development
Mobile equipment
Surface infrastructure
Mine service infrastructure
Closure costs
Salvage value
Sustaining
capital
$20.01 M
$0.21 M
$6.45 M
$7.29 M
EPCM
$0.77 M
Total
$36.76 M
Total cost
$21.33 M
-$19.30 M
$1.00 M
$1.00 M
$17.13 M
$0.18 M
$0.02 M
$0.78 M
$3.70 M
-$1.46 M
$37.14 M
$0.39 M
$6.48 M
$8.06 M
$3.70 M
-$1.46 M
$0.77 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
Definition drilling and sampling
Stope development
Contractor indirect costs
Mining costs
O'Brien staff and general
Energy costs
Milling and transportation
Environment
Total
Total cost
$2.47 M
$22.09 M
$18.06 M
$27.30 M
$12.38 M
$5.89 M
$23.64 M
Unit cost
($/t)
($/oz)
3.85
34.38
28.11
42.48
19.27
9.17
36.78
20.29
181.16
148.11
223.84
101.53
48.32
193.80
$1.97 M
3.06
16.14
$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 preproduction 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
Current mineral resources included (indicated and
inferred)
Results
712,521 tonnes @ 6.46 g/t Au
Mill recovery
91.5%
Life of mine ("LOM") (including 24 months of preproduction)
6 years
Daily mine production
440 tpd
Gold recovered over LOM
Gold price (USD)
Exchange rate (CAD/USD)
Gold price (CAD)
135,308 oz
$1,180
1.25
$1,475
Total gross revenue
$199.5M
Pre-production capital cost
$36.8M
Average operating cost per tonne
Average operating cost per ounce in US$
$178/tonne
US$752/ounce
PRE-TAX
LOM NPV at 5% discount rate (C$)
$0.2M
Internal Rate of Return (IRR)
5.18%
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Parameters
Payback period (years)
Results
5.6
AFTER-TAX
LOM NPV at 5% discount rate (C$)
IRR (%)
$(1.9)M
3.15%
Payback period (years)
5.8
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
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.
Potential Impact
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;
Possible Risk Mitigation
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.
Drilling from either surface or underground locations
could breach the confinement facility.
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).
The minimum mining width used for the resource
estimate might need to be adjusted if assumptions
differ from reality.
Surface and/or underground
geotechnical evaluations not
available
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
Regarding specifically at
Westwood mill opportunities:
A regrind mill could be
refurbished to reduce the grind
size before cyanidation.
Surface definition diamond
drilling
Explanation
Potential benefit
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.
Potential to upgrade some inferred
resources to the indicated category.
Adding indicated resources increases the
economic value of the mining project.
Potential to identify additional inferred
resources.
Adding inferred resources increases the
economic value of the mining project.
Potential to identify additional inferred
resources.
Adding inferred resources increases the
economic value of the mining project.
Potential to identify additional inferred
resources.
Adding inferred resources increases the
economic value of the mining project.
Potential to identify additional inferred
resources.
Adding inferred 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
Surface exploration diamond
drilling on Target 2
Extention at depth of the ore
shoot originating in the
Kewagama area
Surface exploration diamond
drilling on Target 3
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
1a
Property-scale compilation including 3D compilation of all
remaining historical openings and historical data
1b Surface exploration drilling (all inclusive)
1c Stakeholder mapping, communication plan
1d Environmental studies
1e 3D model and resource estimate update
1f PEA update
Contingencies (~ 15%)
Phase 1 subtotal
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Cost
$100,000
25,000 m
$2,500,000
$50,000
$300,000
$80,000
$250,000
$490,000
$3,770,000
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Budget
Phase 2 - Work Program
2a Surface exploration and/or conversion drilling (all inclusive)
2b 3D model and resource estimate update
2c Provision for an underground development program
2d
Provision for environmental and hydrogeological characterization
studies
2e Metallurgical testing
Description
Cost
25,000 m
$2,500,000
$80,000
$13,500,000
$600,000
$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 Vald’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 miningincompatible 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 Vald’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
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
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.
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Table 6.2 – Total gold production of the O’Brien mine from 1926 to 1957
Year
Metric Tonnes Metric Tonnes
Au g/t
Ounces of Gold Metric Tonnes
Mined (Hoisted)
Milled
Milled grade
Recovered
Development
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
Au g/t
Development
Metric Tonnes
Stopes
Au g/t
Stopes
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
118,635
10.07
936,427
16.17
1957
TOTAL
2,074.4
1,062,749
1,197,147
15.25
587,120.8
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
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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 D25) 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, D31 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 DS28, 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
Year
Metric Tonnes
Milled
Au g/t
Ounces of Gold
Milled grade
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
* 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.
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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 (213017 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 7.6m (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 (OB95-47 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) cutoff 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 RM11-18) 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
Year
1995
1996
1997
1998
2003
2004
2006
2007
2008
2011
2012
TOTAL
Number of
Holes
10
31
37
7
3
2
3
15
28
6
3
145
Total Length
(meter)
3,726.2
14,530.1
6,586.1
1,949.5
210.3
1,656.0
1.198.0
2,053.2
5,659.3
1,989.0
2,113.5
41,610.7
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.
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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. Thirtyone (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 (KW08151, 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
2003
Number of
Holes
1
Total Length
(meter)
176.0
2004
7
4,229.3
2005
5
3,030.0
2006
5
2,237.0
2008
11
4,946.8
2011
13
4,359.8
Total
42
18,978.9
Year
Table 6.6 ─ Best results obtained from Radisson’s drilling campaigns
From
to
Corelenght
Au g/t
Hole
(meter)
(meter)
(meter)
(uncut)
KW04-02W
1,228.8
1,229.8
1.0
17.46
KW04-03
515.4
523.6
8.2
5.45
KW05-11
564.0
565.5
1.5
5.42
KW06-13
183.1
184.1
1.0
10.4
KW06-15
202.6
203.8
1.2
10.60
KW06-16
443.0
443.8
0.8
20.7
KW08-151
554.0
557.0
3.0
4.92
KW08-155A
341.8
347.2
5.4
3.75
KW08-156
603.6
605.6
2.0
6.64
KW08-157
165.7
166.0
0.3
466.48
KW08-164
522.8
535.9
13.1
1,83
RM-11-11
132.8
137.0
4.2
2,53
RM-11-12
200.9
202.6
1.7
10.00
RM-11-13
131.3
132.4
1.0
18.90
RM-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.
2.
3.
4.
5.
6.
7.
Pre-2750 Ma volcanic episode;
Pacaud Assemblage (2750-2735 Ma);
Deloro Assemblage (2734-2724 Ma);
Stoughton-Roquemaure Assemblage (2723-2720 Ma);
Kidd-Munro Assemblage (2719-2711 Ma);
Tisdale Assemblage (2710-2704 Ma);
Blake River Assemblage (2704-2695 Ma).
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-talccarbonate 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 goldbearing 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 highgrade 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 eastwest (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 highgrade 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 goldbearing 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 quartzcarbonate 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 volcanoplutonic 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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Area 2
Area 1
Area 3
Area 4
Area 5
Area 6
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.
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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
No East Zone
1981
Mixed with 30% East
Zone
No gravity recovery
71.7
%
27.3
Mixed with 23%
East Zone
10.7
%
76.2
77.6
Ore origin
Gravity
recovery
Overall
recovery
1980
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
Flotation
Gravity + Flotation
Cyanidation of concentrate
Gravity + Amalgamation + Flotation +
Cyanidation
Gravity + Flotation + Fine regrind + CIP
23-75%
90-95%
 95%
53-78%
85%
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
KN-F-2
KN-F-3
KN-F-5
KN-F-6-R
KN-CN-F-4
KN-CN-2
9.97
10.39
10.57
14.59
11.00
11.23
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
µm
139
105
73
37
Gold recovery
%
90.2
94.0
95.8
94.9
Mass recovery
%
10.4
8.8
11.1
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
K 80
µm
Test
1
Test
2
Test
3
Gravity
Gold
Mass
recovery
recovery
%
%
Regrind
K80
%
Flotation
Gold
recovery
%
Mass
recovery
%
Total
Gold
recovery
%
102
55.3
0.03
NA
38.3
8.69
93.6
102
54.3
0.06
73
39.1
9.35
93.4
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
%
1.82
1.85
96.33
100
1st cleaner
Cleaner scavenger
Tailings
Total
As grade
%
20.2
5.51
0.05
0.52
As recovery
%
71.4
19.8
8.8
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
Cyclic test 1
Cyclic test 2
K 80
µm
Gravity
%
Flotation
%
Total
%
102
102
67.4
60.2
26.2
34.4
93.6
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
Gravity
Cyanidation
Test 1
Test 2
Grind size
µm
102
102
Au recovery
%
58
60.8
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
89.6
92.9
Total
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 43101 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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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
(Au g/t)
Capping
(Au g/t)
Cut
Cut
2.30
65.00
1.76
23
0.36%
-14.99%
853.40
2.65
30.00
1.56
36
1.13%
-36.10%
11,888
96.72
0.12
3.50
0.11
13
0.11%
-4.03%
5,694
19.46
0.17
4.00
0.16
9
0.16%
-4.31%
27,166
1,019.14
0.94
65
0.68
81
0.30%
-10.42%
Dataset
Block Code
Mineralized zones west of the fault
101 to 128
6,408
1,019.14
Mineralized zones east of the fault
201 to 230
3,176
Dilution envelope west of the fault
501
Dilution envelope east of the fault
502
Samples
T otal
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Histogram
Decile Analysis
300
100%
90%
250
80%
70%
200
150
Frequency
Contained Metal
Frequency
60%
50%
40%
100
30%
20%
50
10%
0
-5
-4.8
-4.6
-4.4
-4.2
-4
-3.8
-3.6
-3.4
-3.2
-3
-2.8
-2.6
-2.4
-2.2
-2
-1.8
-1.6
-1.4
-1.2
-1
-0.8
-0.6
-0.4
-0.2
0
0.2
0.4
0.6
0.8
1
1.2
1.4
1.6
1.8
2
2.2
2.4
2.6
2.8
3
3.2
3.4
3.6
3.8
4
4.2
4.4
4.6
4.8
5
0%
0-10 10-20 20-30 30-40 40-50 50-60 60-70 70-80 80-90 90-100
Classes
90-91 91-92 92-93 93-94 94-95 95-96 96-97 97-98 98-99 99-100
Decile
Upper Decile
80%
Probability Graph
1000
70%
60%
100
Grade
Cummulative metal
50%
10
40%
Capped
Raw
30%
20%
1
10%
0
0.01
0.10
0.50
0.90
Probability
0.99
0%
0%
1%
2%
3%
4%
5%
6%
7%
8%
9%
10%
Number of samples
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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Histogram
Decile Analysis
300
100%
90%
250
80%
70%
200
150
Frequency
Contained Metal
Frequency
60%
50%
40%
100
30%
20%
50
10%
0
-5
-4.8
-4.6
-4.4
-4.2
-4
-3.8
-3.6
-3.4
-3.2
-3
-2.8
-2.6
-2.4
-2.2
-2
-1.8
-1.6
-1.4
-1.2
-1
-0.8
-0.6
-0.4
-0.2
0
0.2
0.4
0.6
0.8
1
1.2
1.4
1.6
1.8
2
2.2
2.4
2.6
2.8
3
3.2
3.4
3.6
3.8
4
4.2
4.4
4.6
4.8
5
0%
0-10 10-20 20-30 30-40 40-50 50-60 60-70 70-80 80-90 90-100
90-91 91-92 92-93 93-94 94-95 95-96 96-97 97-98 98-99 99-100
Decile
Classes
Upper Decile
80%
Probability Graph
1000
70%
60%
100
Grade
Cummulative metal
50%
10
40%
Capped
Raw
30%
20%
1
10%
0
0.01
0.10
0.50
0.90
Probability
0.99
0%
0%
1%
2%
3%
4%
5%
6%
7%
8%
9%
10%
Number of samples
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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Decile Analysis
100%
450
90%
400
80%
350
70%
300
60%
250
Frequency
Contained Metal
Frequency
Histogram
500
50%
200
40%
150
30%
100
20%
50
10%
0
0%
-5 -4.7-4.4-4.1-3.8-3.5-3.2-2.9-2.6-2.3 -2 -1.7-1.4-1.1-0.8-0.5-0.2 0.1 0.4 0.7 1 1.3 1.6 1.9 2.2 2.5 2.8 3.1 3.4 3.7 4 4.3 4.6 4.9
0-10 10-20 20-30 30-40 40-50 50-60 60-70 70-80 80-90 90-100
Classes
90-91 91-92 92-93 93-94 94-95 95-96 96-97 97-98 98-99 99-100
Decile
Upper Decile
80%
Probability Graph
1000
70%
60%
100
Grade
Cummulative metal
50%
10
40%
Capped
Raw
30%
20%
1
10%
0
0.01
0.10
0.50
0.90
Probability
0.99
0%
0%
1%
2%
3%
4%
5%
6%
7%
8%
9%
10%
Number of samples
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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Decile Analysis
100%
450
90%
400
80%
350
70%
300
60%
250
Frequency
Contained Metal
Frequency
Histogram
500
50%
200
40%
150
30%
100
20%
50
10%
0
-5
-4.8
-4.6
-4.4
-4.2
-4
-3.8
-3.6
-3.4
-3.2
-3
-2.8
-2.6
-2.4
-2.2
-2
-1.8
-1.6
-1.4
-1.2
-1
-0.8
-0.6
-0.4
-0.2
0
0.2
0.4
0.6
0.8
1
1.2
1.4
1.6
1.8
2
2.2
2.4
2.6
2.8
3
3.2
3.4
3.6
3.8
4
4.2
4.4
4.6
4.8
5
0%
0-10 10-20 20-30 30-40 40-50 50-60 60-70 70-80 80-90 90-100
90-91 91-92 92-93 93-94 94-95 95-96 96-97 97-98 98-99 99-100
Decile
Classes
Upper Decile
80%
Probability Graph
1000
70%
60%
100
Grade
Cummulative metal
50%
10
40%
Capped
Raw
30%
20%
1
10%
0
0.01
0.10
0.50
0.90
Probability
0.99
0%
0%
1%
2%
3%
4%
5%
6%
7%
8%
9%
10%
Number of samples
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
Mineralized zones
west of the fault
Mineralized zones
east of the fault
Dilution envelope
west of the fault
Dilution envelope
east of the fault
43-101 Technical Report – O’Brien Project
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
501
501
41,431
3.34
0.03
0.12
3.32
502
502
27,184
4.38
0.04
0.13
3.27
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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
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%
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 resourcegrade 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
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
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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
Work-space
Description
GEMS Triangulation Name
Precedence
NAME1
NAME2
NAME3
Mineralized Zone 101
ZoneClip
101
F150227
3
Mineralized Zone 102
ZoneClip
102
F150227
3
Mineralized Zone 103
ZoneClip
103
F150227
3
Mineralized Zone 104
ZoneClip
104
F150227
3
Mineralized Zone 105
ZoneClip
105
F150227
3
Mineralized Zone 106
ZoneClip
106
F150227
3
Mineralized Zone 107
ZoneClip
107
F150227
3
Mineralized Zone 108
ZoneClip
108
F150227
3
Mineralized Zone 109
ZoneClip
109
F150227
3
Mineralized Zone 110
ZoneClip
110
F150227
3
Mineralized Zone 111
ZoneClip
111
F150227
3
Mineralized Zone 112
ZoneClip
112
F150227
3
Mineralized Zone 113
ZoneClip
113
F150227
3
Mineralized Zone 114
ZoneClip
114
F150227
3
Mineralized Zone 115
ZoneClip
115
F150227
3
Mineralized Zone 116
ZoneClip
116
F150227
3
Mineralized Zone 117
ZoneClip
117
F150227
3
Mineralized Zone 118
ZoneClip
118
F150227
3
Mineralized Zone 119
ZoneClip
119
F150227
3
Mineralized Zone 120
ZoneClip
120
F150227
3
Mineralized Zone 121
ZoneClip
121
F150227
3
Mineralized Zone 123
ZoneClip
123
F150227
3
Mineralized Zone 124
ZoneClip
124
F150227
3
Mineralized Zone 125
ZoneClip
125
F150227
3
Mineralized Zone 126
ZoneClip
126
F150227
3
Mineralized Zone 127
ZoneClip
127
F150227
3
Mineralized Zone 128
ZoneClip
128
F150227
3
Mineralized Zone 201
ZoneClip
201
F150227
4
Mineralized Zone 202
ZoneClip
202
F150227
4
Mineralized Zone 203
ZoneClip
203
F150227
4
Mineralized Zone 204
ZoneClip
204
F150227
4
Mineralized Zone 205
ZoneClip
205
F150227
4
Mineralized Zone 206
ZoneClip
206
F150227
4
Mineralized Zone 207
ZoneClip
207
F150227
4
Mineralized Zone 208
ZoneClip
208
F150227
4
Mineralized Zone 209
ZoneClip
209
F150227
4
Mineralized Zone 210
ZoneClip
210
F150227
4
Mineralized Zone 211
ZoneClip
211
F150227
4
Mineralized Zone 212
ZoneClip
212
F150227
4
Mineralized Zone 213
ZoneClip
213
F150227
4
Mineralized Zone 214
ZoneClip
214
F150227
4
Mineralized Zone 215
ZoneClip
215
F150227
4
Mineralized Zone 216
ZoneClip
216
F150227
4
Mineralized Zone 219
ZoneClip
219
F150227
4
Mineralized Zone 220
ZoneClip
220
F150227
4
Mineralized Zone 221
ZoneClip
221
F150227
4
Mineralized Zone 222
ZoneClip
222
F150227
4
Mineralized Zone 223
ZoneClip
223
F150227
4
Mineralized Zone 224
ZoneClip
224
F150227
4
Mineralized Zone 225
ZoneClip
225
F150227
4
Mineralized Zone 226
ZoneClip
226
F150227
4
Mineralized Zone 227
ZoneClip
227
F150227
4
Mineralized Zone 228
ZoneClip
228
F150227
4
Mineralized Zone 229
ZoneClip
229
F150227
4
Mineralized Zone 230
ZoneClip
230
F150227
4
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_100
Zones_200
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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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Indicated resource (default)
Indicated resource (upgraded)
Inferred category (default)
Inferred resource (downgraded)
Unca tegorized material
Figure 14.10 – Longitudinal view looking north showing all interpolated blocks of Zone
101 with respective categorization.
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Indicated resource (default)
Indicated resource (upgraded)
Inferred category (default)
Inferred resource (downgraded)
Unca tegorized material
Figure 14.11 – Longitudinal view looking north showing all interpolated blocks of Zone 222
with respective categorization.
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Au (g/t)
3.50 to 4.00
4.00 to 4.50
4.50 to 5.00
5.00 to 10.00
> 10.00
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.
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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
Input parameter
Value
Gol d pri ce ($US/oz)
Excha nge ra te
Gol d pri ce ($C/oz)
Gol d s el l i ng cos ts ($C/oz)
Net gold price ($C/oz)
1,200.00
1.00 USD : 1.20 CAD
1,440.00
5.00
1,435.00
Mi ni ng cos ts ($C/t)
94.98
Mi l l i ng cos ts ($C/t)
38.30
Total costs ($C/t)
133.28
Proces s i ng recovery (%)
92.50
Mi ni ng di l ution (%)
15.00
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.
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1.60
$1,600
1.55
$1,500
1.50
$1,400
1.45
$1,300
1.40
$1,200
1.35
$1,100
1.30
$1,000
1.25
$900
1.20
$800
1.15
$700
05/10/2014
Exchange Rate 1US = CAD
Gold Price
Gold Price and Exchange Rate
$1,700
1.10
04/11/2014
04/12/2014
Gold Price $USD
03/01/2015
Gold Price $CAD
02/02/2015
04/03/2015
03/04/2015
Exchange rate 1US=CAD
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 cutoff grade (O’Brien and Kewagama claim blocks), as well as the sensitivity at other cutoff 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.
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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
Indicated
Zone
All
Zones
















Cut-off
Tonnage
Inferred
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
All
Zones
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.
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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, longhole mining with captive sublevels was selected.
The mineralized material will be transported to surface using a combination of 3.5cubic-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
Inferred
570,800
918,300
6.53
6.38
119,819
188,466
38%
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 cutoff 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
Mint cost
Mill recovery
Mining
dilution
Cut-off grade
$173.28/t
$5.00/oz
91.50%
30%
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)
8 ft (2.4 m)
16.5 ft (5 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
Production (t)
Grade (g/t)
Development (t)
Grade (g/t)
Total tonnage milled (t)
Grade (g/t)
Pre-production
Year 1
Year 2
Year 3
33,194 126,494
7.20
7.05
3,196
33,474 32,080
7.05
5.74
6.19
3,196
66,668 158,574
7.05
6.47
6.87
Production
Year 4
Year 5
129,593 134,524
7.39
5.66
40,298
52,409
5.95
5.11
169,891 186,933
7.04
5.50
Year 6
127,259
6.53
127,259
6.53
Total
551,064
6.68
161,457
5.70
712,521
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
Development CAPEX
Ramp and main level
4.5 × 4.5 (m)
Alimak raise 2.4 × 2.4
(m)
Development OPEX
Sublevel 3.0 × 3.0 (m)
Conventional raise 2.4
× 2.4 (m)
Pre-production
Year 1
Year 2
Year 3
Production
Year 4
Year 5
1,053
3,145
2,357
766
-
10,289
107
191
-
351
2,482
3,306
2,984
-
11,601
114
18
-
132
2,968
53
255
2,574
Year 6
Total
Table 16.6 – O’Brien mine production rates
Tonnage summary
Mineralized material
(t)
Waste (t)
Total
Total mineralized
material per day (t/day)
Total mineralized
material and waste per
day (t/day)
Backfill (t)
Pre-production
Year 1
Year 2
Year 3
Production
Year 4
Year 5
Year 6
3,196
67,296
70,492
66,668
189,667
256,335
158,574
199,358
357,932
169,891
168,504
338,395
186,934
63,649
250,583
127,259
N/A
183
434
465
512
465
230
702
37,946
981
122,692
927
146,560
687
86,746
465
53,076
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127,259
Total
712,522
688,474
1,400,996
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
Administration
Manager
Secretary
Senior accountant
Intermediate accountant
Junior accountant
Senior purchaser
Clerk
Nurse
Surface services
Dryman
Gate keeper
Technical services
Geology
Chief geologist
Senior geologist
Junior geologist
Database technician
Senior geology technician
Engineering
Chief engineer
Senior engineer
Junior engineer
Mining technician
Senior surveyor
Surveyor
Pre-production
Year 1
Year 2
1
1
1
1
Year 3
Production
Year 4
Year 5
Year 6
1
1
1
1
1
1
2
1
1
1
1
1
1
1
2
1
1
1
1
0.75
0.75
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
2
1
1
4
2
4
2
4
2
4
1.5
4
1
4
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
0.5
0.5
1
1
1
1
1
2
1
1
1
1
2
1
1
1
1
1
2
1
1
1
1
1
2
1
1
1
0.5
0.5
2
0.5
1
1
1
1
1
2
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
Mill
Company
Process
Capacity
Distance
Mill status
(operating
or closed)
Interest for
custom
milling
Kiena Mill
Wesdome
Leaching/CIP
48 km
Closed
NA
SigmaLamaque
Complex
Camflo
Mill
Westwood
Mill
Integra
Gold
Gravity
Concentration &
Leaching/CIP
Leaching/MerrillCrowe
Gravity
Concentration &
Leaching/CIP
Or
Gravity
Concentration &
Flotation
Gravity
Concentration &
Flotation & Flotation
Concentrate
Leaching
1,000 to
2,200 tpd
1,200 to
2,400 tpd
67 km
Closed
No interest
800 to
1,200 tpd
2,400 tpd
35 km
Operating
No interest
19 km
Operating
Yes
75 km
Operating
Yes with
environmental
conditions
Aurbel Mill
Richmont
Mines
Iamgold
QMX
800 tpd
500 to
800 tpd
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
g/t
6.46
6.46
%
94.52
91.53
C$/t
289
280
Preparation and trucking
$/t
5.78
5.78
Custom milling
$/t
31
31
Smelting
$/t
45
NA
Total
$/t
81.78
36.78
Gold value
Ore grade1
Recovery
Total3
Milling cost
1 Based
on mining plan
test KN-F-3
3 Assumption section 17.1.2
BASED ON Gold PRICE at C$1475 /oz
2 URSTM
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
Gravity feed
size K80
µm
Gold Recovery in Laboratory Tests
Gravity
Regrind K80
Leach - 48 h
recovery
%
µm
%
Total
%
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 14120460-00-00-00-0001 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 (0¾ 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 offroad 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 1000CFM 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 68 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:
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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:
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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 nonpotentially 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 runoff 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/2012147) 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:
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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). Preproduction 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
Preproduction
Capitalized operating costs
Capitalized revenue
$21.33 M
-$19.30 M
Royalty payment
Development
Mobile equipment
Surface infrastructure
Mine service infrastructure
Closure costs
$20.01 M
$0.21 M
$6.45 M
$7.29 M
Salvage value
EPCM
$0.77 M
Total
$36.76 M
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Sustaining
capital
Total cost
$21.33 M
-$19.30 M
$1.00 M
$17.13 M
$0.18 M
$0.02 M
$0.78 M
$3.70 M
$1.00 M
$37.14 M
$0.39 M
$6.48 M
$8.06 M
$3.70 M
-$1.46 M
-$1.46 M
$0.77 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
Stope development
Contractor indirect costs
Mining costs
O'Brien staff and general
Energy costs
Milling and transportation
$0.27 M
$4.68 M
$2.33 M
$2.34 M
$6.36 M
$1.96 M
$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
Sustaining
Total cost
capital
$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
Description
Pre-production
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
Description
21.1.6.1
Pre-production
Site preparation and
installation
$4.45 M
Buildings
$0.56 M
Water management
$1.44 M
Total
$6.45 M
Sustaining
capital
Total cost
$4.45 M
$0.02 M
$0.58 M
$1.44 M
$0.02 M
$6.48 M
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 $4,453,238.
Table 21.5 presents the cost breakdown.
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Table 21.5 – Site preparation and installation
Description
Overburden pad
Waste pad
Powder magazine
Cap magazine
Fuel station (earthwork and concrete)
Garage (earthwork and concrete)
Electrical substation (concrete and mechanical)
Storage containers
Truck scale
79,834
1,520,573
36,361
32,631
353,492
Compressors (earthwork and concrete)
Loading station
157,392
279,444
5,344
258,899
6,367
5,189
Ventilation heating system (earthwork and concrete)
Propane vessel (earthwork and concrete)
Ramp portal
On-site roads
Office (earthwork)
Septic system
52,902
4,666
45,360
350,733
118,890
19,710
Site fencing
Employee parking
Potable water well
Water treatment pond
Final water treatment system (earthwork)
Water treatment polishing pond (earthwork)
286,474
87,825
25,944
330,000
70,784
147,724
Propane installation
Powder and cap magazine installation
Safety exit
Total
21.1.6.2
Total cost ($)
18,000
36,000
122,700
4,453,238
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
21.1.6.3
Total ($)
Warehouse
Processing plant
Garage
Storage containers
Office
7,020
32,760
167,946
25,208
296,412
Gate
Powder and cap magazine
Total
13,000
34,884
577,230
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 $1,444,902 and the
cost breakdown is presented in Table 21.7.
Table 21.7 – Water management and distribution
Description
Environmental study
Septic system (mechanical and piping)
Gate (piping)
Potable water well (mechanical and piping)
Water treatment pond (piping)
Water treatment system (mechanical and environment)
Water treatment polishing pond (piping)
Potable water treatment (mechanical)
Total
Total ($)
500,000
129,248
6,412
164,233
10,100
593,949
4,040
36,920
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
Mine dewatering
$1.05 M
Compressed air distribution
$0.77 M
Ventilation and air heating
Electrical distribution and
comm. systems
$1.41 M
Total
$7.29 M
Production
$0.52 M
Total
$1.56 M
$0.77 M
$0.26 M
$1.67 M
$4.06 M
$4.06 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 $177.10 per tonne.
Operating costs are summarized below for the production period (Table 21.9).
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Table 21.9 – Summary of operating costs
Description
Definition drilling and sampling
Stope development
Contractor indirect costs
Mining costs
O'Brien staff and general
Energy costs
Milling and transportation
Environment
Total
Total cost
Unit cost
($/t)
($/oz)
$2.47 M
$22.09 M
$18.06 M
$27.30 M
$12.38 M
$5.89 M
$23.64 M
3.85
34.38
28.11
42.48
19.27
9.17
36.78
20.29
181.16
148.11
223.84
101.53
48.32
193.80
$1.97 M
3.06
16.14
$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 $24.88/t. The mobile
equipment operating cost only includes the cost for equipment provided by Radisson,
namely pickup trucks and underground trucks. Manpower only includes the minimal
surface crew (dryman and gate keeper) and the material needed. All these costs
include 10% contingency.
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Table 21.10 – Mining costs
Description
Total cost
Stoping
Mobile equipment
Surface manpower and
material
Total
Unit cost
($/t)
($/oz)
$25.62 M
39.87
210.07
$0.43 M
0.67
3.53
$1.25 M
1.94
10.24
$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 $12.38 million.
The annual cost is detailed in Table 21.11.
Table 21.11 – O’Brien staff salaries
Manpower
Administration
Annual cost
($)
Manager
271,250
Secretary
Senior accountant
Intermediate accountant
Senior purchaser
Clerk
Nurse
82,800
157,300
117,300
128,700
138,000
103,500
Building rental
Material and others
121,080
877,257
Subtotal 1,997,187
Technical services
Chief geologist
Senior geologist
171,600
128,700
Junior geologist
Database technician
Senior geology technician
82,800
89,700
103,500
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Annual cost
($)
Manpower
Chief engineer
Technical services
171,600
Senior engineer
Junior engineer
Mining technician
Senior surveyor
128,700
89,700
103,500
207,000
Surveyor
Material and others
Subtotal
Total
69,000
240,720
1,586,520
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 $9.17 per tonne milled for Years 3-6.
Table 21.12 – Annual energy cost (average for Years 3-6)
Description
Electricity
Propane
Propane tank rental
Total
Annual cost ($)
1,124,069
635,905
3,900
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
Analyses and environmental monitoring
171,600
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):







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 preproduction 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
Parameters
Current mineral resources included (indicated and
inferred)
Results
712,521 tonnes @ 6.46 g/t Au
Mill recovery
91.5%
Life of mine ("LOM") (including 24 months of preproduction)
6 years
Daily mine production
440 tpd
Gold recovered over LOM
Gold price (USD)
Exchange rate (CAD/USD)
Gold price (CAD)
135,308 oz
$1,180
1.25
$1,475
Total gross revenue
$199.5M
Pre-production capital cost
$36.8M
Average operating cost per tonne
Average operating cost per ounce in US$
$178/tonne
US$752/ounce
PRE-TAX
LOM NPV at 5% discount rate (C$)
$0.2M
Internal Rate of Return (IRR)
5.18%
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Parameters
Payback period (years)
Results
5.6
AFTER-TAX
LOM NPV at 5% discount rate (C$)
IRR (%)
Payback period (years)
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Table 22.2 – Economic analysis for the O’Brien Project (figures in Canadian dollars)
Radisson - O'Brien Project
Cashflow summary
Pre production
Year 1
Production
Year 2
Year 3
Year 4
Year 5
Year 6
Total
PRODUCTION
ALL ZONES
Development (t)
3 196
33 474
32 080
40 298
52 409
-
Grade (g/t)
7,05
5,74
6,19
5,95
5,11
-
Long Hole (t)
-
33 194
126 494
129 593
134 524
Grade (g/t)
3 196
Grade (g/t)
7,20
7,05
7,39
5,66
6,53
6,68
158 574
169 891
186 934
127 259
712 521
7,05
Recovery (%)
91,5%
5,70
551 064
127 259
66 668
-
Total (tonne milled)
161 458
6,47
91,5%
6,87
91,5%
7,04
91,5%
5,50
91,5%
6,46
6,53
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
13
221
256
-
-
-
3 196
66 668
663
12 682
Gold Produced (oz) Royalty 2% Kewagama
Tonne milled assigned to capital
Gold Produced assigned to capital (oz)
490
69 864
13 345
Tonne milled assigned to production
158 574
Grade (g/t)
Gold Produced assigned to production(oz)
169 891
186 934
642 658
127 259
6,87
7,04
5,50
6,53
6,45
32 057
35 206
30 261
24 439
121 963
REVENUES
1 180 $
Gold Price ($US/oz)
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 $
977 850 $
18 706 105 $
47 284 419 $
51 928 456 $
44 635 278 $
36 047 344 $
199 579 452 $
1 989 $
38 046 $
96 172 $
105 617 $
90 784 $
73 317 $
19 556,99 $
325 260 $
377 849 $
956 304 $
18 342 799 $
Gross Revenue
Mint (cost 3,00$ per oz)
Kewagama royalty (NSR 2%)
Capitalized revenue
Net Revenue
405 924 $
722 666 $
19 299 103 $
46 810 398 $
51 822 839 $
44 544 495 $
35 974 027 $
179 151 759 $
OPERATING EXPENDITURES
Definition 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 $
339 466 $
1 992 621 $
3 244 190 $
4 871 956 $
5 787 907 $
4 160 058 $
20 396 199 $
299 255 $
2 038 487 $
6 240 115 $
6 734 672 $
7 510 252 $
6 814 711 $
29 637 492 $
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 $
571 600 $
571 600 $
571 600 $
253 700 $
Contractor indirect cost
Mining cost
O'Brien staff and General
Environment
Capitalized operating cost
345 860 $
471 600 $
4 904 510 $
16 425 451 $
0$
0$
Total operating costs
2 785 960 $
21 329 961 $
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 $
20 692 051 $
18 100 143 $
9 590 436 $
16 955 379 $
65 338 008 $
Operating Cash Flow
0$
0$
4 904 510 $
16 425 451 $
956 304 $
18 342 799 $
CAPITAL EXPENDITURES
Capitalized operating cost
Capitalized revenue
Royalty payment
Preproduction contractor
21 329 961 $
19 299 103 $
1 000 000 $
1 000 000 $
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 $
147 847 $
60 672 $
60 672 $
60 672 $
60 672 $
0$
390 536 $
4 453 238 $
0$
0$
0$
0$
0$
4 453 238 $
Mobile Equipment
Site preparation and installation
Buildings
Water management and distribution - Environment
Ventilation and Air heating
547 854 $
7 344 $
7 344 $
7 344 $
7 344 $
0$
577 230 $
1 444 902 $
0$
0$
0$
0$
0$
1 444 902 $
480 751 $
932 762 $
232 625 $
28 564 $
0$
0$
1 674 702 $
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 $
24 223 465 $
12 540 664 $
14 030 376 $
4 901 006 $
68 016 $
110 000 $
55 873 527 $
Electrical distribution & Communication system
Total capital expenditures
1 095,96 $
All-in sustaining cost/oz $CAN
876,77 $
All-in sustaining cost/oz $US
1 447,43 $
All-in cost/oz $CAN
1 157,94 $
All-in cost/oz $US
Salvage
1 460 908 $
Financial guarantee reimbursement
1 851 634 $
925 817 $
925 817 $
Closure Costs
1 460 908 $
3 703 268 $
0$
3 703 268 $
3 703 268 $
7 222 121 $
Net cashflow
24 223 465 $
12 540 664 $
4 810 041 $
12 273 320 $
8 596 603 $
18 306 286 $
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 $
Pre-tax IRR
After-tax NPV (5%)
(0,8)
5,18%
1 908 446 $
After-tax IRR
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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 $)
Revenue
Opex
Capex
-30%
-20%
-10%
(44.47)
18.40
9.28
(29.26)
11.60
5.76
(15.41)
5.05
2.00
Base Case
scenario
(1.91)
(1.91)
(1.91)
10%
20%
30%
8.17
(11.11)
(6.89)
17.84
(20.50)
(12.21)
27.36
(30.21)
(17.72)
Figure 22.1 – Sensitivity analysis of economic parameters on after-tax NPV at
5% (millions $)
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Table 22.4 – Sensitivity analysis of grade and Gold Price on after-tax NPV at
5% (millions $)
Grade (g/t)
Gold price
(US$/oz)
Resulting
NPV (M$)
-30%
-20%
-10%
8.39
7.75
7.10
Base
Case
scenario
6.46
826
944
1062
(44.47)
(29.26)
(15.41)
10%
20%
30%
5.81
5.16
4.52
1180
1298
1416
1534
(1.91)
8.17
17.84
27.36
Figure 22.2 – Sensitivity analysis of grade on after-tax NPV at 5% (millions $)
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Table 22.5 – Sensitivity analysis of economic parameters on after-tax IRR
Revenue
Opex
Capex
-30%
-20%
-10%
-47%
23%
18%
-26%
16%
12%
-11%
10%
7%
Base Case
scenario
3%
3%
3%
10%
20%
30%
13%
-6%
-1%
21%
-15%
-5%
29%
-26%
-9%
Figure 22.3 – Sensitivity analysis of economic parameters on after-tax IRR
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Table 22.6 – Sensitivity analysis of grade and Gold Price on after-tax IRR
Grade
(g/t)
Gold price
(US$/oz)
IRR
-30%
-20%
-10%
Base Case
scenario
10%
20%
30%
8.39
7.75
7.10
6.46
5.81
5.16
4.52
826
944
1062
1180
1298
1416
1534
-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 ThompsonCadillac 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 talcchlorite 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)
Mill
Company
Process
Capacity
Distance
Mill status
(operating
or closed)
Interest for
custom
milling
Kiena Mill
Wesdome
Leaching/CIP
48 km
Closed
NA
SigmaLamaque
Complex
Camflo
Mill
Westwood
Mill
Integra
Gold
Gravity
Concentration &
Leaching/CIP
Leaching/MerrillCrowe
Gravity
Concentration &
Leaching/CIP
Or
Gravity
Concentration &
Flotation
Gravity
Concentration &
Flotation & Flotation
Concentrate
Leaching
1,000 to
2,200 tpd
1,200 to
2,400 tpd
67 km
Closed
No interest
800 to
1,200 tpd
2,400 tpd
35 km
Operating
No interest
19 km
Operating
Yes
75 km
Operating
Yes with
environmental
conditions
Aurbel Mill
Richmont
Mines
Iamgold
QMX
800 tpd
500 to
800 tpd
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
g/t
6.46
6.46
%
94.52
91.53
C$/t
289
280
Preparation and trucking
$/t
5.78
5.78
Custom milling
$/t
31
31
Smelting
$/t
45
NA
Total
$/t
81.78
36.78
Gold value
Ore grade1
Recovery
Total3
Milling cost
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/2012147) 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 $1,46M.
Capital cost estimate (Table 21.1)
Description
Capitalized operating cost
Capitalized revenue
Royalty payment
Development
Pre-production
Sustaining
21.33 M$
-19.30 M$
20.01 M$
Mobile Equipment
Surface infrastructure
Mine service infrastructure
Closure cost
Salvage value
0.21 M$
6.45 M$
7.29 M$
EPCM
0.77 M$
Total
36.76 M$
Total cost
21.33 M$
-19.30 M$
1.00 M$
17.13 M$
1.00 M$
37.14 M$
0.18 M$
0.02 M$
0.78 M$
3.70 M$
-1.46 M$
0.39 M$
6.48 M$
8.06 M$
3.70 M$
-1.46 M$
0.77 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 $177.10 per tonne.
Operating costs are summarized below for the production period.
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Summary of total operating costs (Table 21.9)
Description
Definition drilling and sampling
Stope development
Contractor indirect cost
Mining cost
O'Brien staff and general
Energy cost
Milling and transportation
Environment
Total
Total cost
Unit cost
($/t)
($/oz)
2.47 M$
22.09 M$
18.06 M$
27.30 M$
12.38 M$
5.89 M$
23.64 M$
3,85
34,38
28,11
42,48
19,27
9,17
36,78
20,29
181,16
148,11
223,84
101,53
48,32
193,80
1.97 M$
3,06
16,14
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, longhole mining with captive sublevels was selected.
The mineralized material will be transported to surface using a combination of 3.5cubic-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)
Production (t)
Grade (g/t)
Development (t)
Grade (g/t)
Total tonnage milled (t)
Grade (g/t)
Pre-production
Year 1
Year 2
Year 3
33,194 126,494
7.20
7.05
3,196
33,474 32,080
7.05
5.74
6.19
3,196
66,668 158,574
7.05
6.47
6.87
Production
Year 4
Year 5
129,593 134,524
7.39
5.66
40,298
52,409
5.95
5.11
169,891 186,933
7.04
5.50
Year 6
127,259
6.53
127,259
6.53
Total
551,064
6.68
161,457
5.70
712,521
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.








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;
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 preproduction 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)
Parameters
Current mineral resources included (indicated and
inferred)
Results
712,521 tonnes @ 6.46 g/t Au
Mill recovery
91.5%
Life of mine ("LOM") (including 24 months of preproduction)
6 years
Daily mine production
440 tpd
Gold recovered over LOM
Gold price (USD)
Exchange rate (CAD/USD)
Gold price (CAD)
135,308 oz
$1,180
1.25
$1,475
Total gross revenue
$199.5M
Pre-production capital cost
$36.8M
Average operating cost per tonne
Average operating cost per ounce in US$
$178/tonne
US$752/ounce
PRE-TAX
LOM NPV at 5% discount rate (C$)
$0.2M
Internal Rate of Return (IRR)
5.18%
Payback period (years)
5.6
AFTER-TAX
LOM NPV at 5% discount rate (C$)
IRR (%)
Payback period (years)
$(1.9)M
3.15%
5.8
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
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.
Potential Impact
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;
Possible Risk Mitigation
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.
Drilling from either surface or underground locations
could breach the confinement facility.
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).
The minimum mining width used for the resource
estimate might need to be adjusted if assumptions
differ from reality.
Surface and/or underground
geotechnical evaluations not
available
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.
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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.
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.
Potential to upgrade some inferred
resources to the indicated category.
Adding indicated resources increases the
economic value of the mining project.
Potential to identify additional inferred
resources.
Adding inferred resources increases the
economic value of the mining project.
Potential to identify additional inferred
resources.
Adding inferred resources increases the
economic value of the mining project.
Potential to identify additional inferred
resources.
Adding inferred resources increases the
economic value of the mining project.
Potential to identify additional inferred
resources.
Adding inferred 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.
Surface definition diamond
drilling
Surface exploration diamond
drilling on Target 1
Extension of the mineralization
within the drilling gap between
the historical Kewagama mine
and the 36E area
Surface exploration diamond
drilling on Target 2
Extention at depth of the ore
shoot originating in the
Kewagama area
Surface exploration diamond
drilling on Target 3
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
o
o
o
o
o
Measure ball mill and abrasion work indexes to better estimate power
and grinding media consumption;
Conduct metallurgical tests in line with the Westwood mill flowsheet
(gravity concentration followed by cyanidation of gravity tails) to
optimize reagent consumption;
Conduct metallurgical tests with a longer retention time;
Conduct further diagnostic testing (via QEMSCAN or other) to
determine the nature of the unleached gold;
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;
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
Phase 1 - Work Program
Budget
Description
1a
Property-scale compilation including 3D compilation of all
remaining historical openings and historical data
1b Surface exploration drilling (all inclusive)
Cost
$100,000
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
1f
$80,000
PEA update
$250,000
Contingencies (~ 15%)
$490,000
Phase 1 subtotal
$3,770,000
Budget
Phase 2 - Work Program
2a Surface exploration and/or conversion drilling (all inclusive)
2b 3D model and resource estimate update
2c Provision for an underground development program
2d
Provision for environmental and hydrogeological characterization
studies
2e Metallurgical testing
Description
Cost
25,000 m
$2,500,000
$80,000
$13,500,000
$600,000
$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.
43-101 Technical Report – O’Brien Project
194
www.innovexplo.com
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.
43-101 Technical Report – O’Brien Project
195
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 197200.
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.
43-101 Technical Report – O’Brien Project
196
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.
43-101 Technical Report – O’Brien Project
197
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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 19341945. 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.
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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
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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 9101.
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.
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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
1 ft
1 acre
1 ounce (troy)
1 pound (avdp)
1 ton (short)
1 ounce (troy) / t (short)
25.4
0.3048
0.405
31.10348
0.454
0.907
34.286
mm
m
ha
g
kg
t
g/t
Abbreviations
°C
ha
g
kg
mm
cm
m
km
masl
’ or ft
cfm
m3/min
usgpm
Mbs
LOM
$M
$ or C$ or CAD
US$ or USD
degrees Celsius
hectares
grams
kilograms
millimetres
centimetres
metres
kilometres
metres above sea level
ft
cubic ft per minute
cubic metres per minute
US gallons per minute
megabytes per second
life-of-mine
millions of dollars
Canadian dollars
American dollars
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oz
avdp
st
oz/t
t
Mt
t.milled
t.moved
t.mined
tpd / tpy
g/t
ppb
ppm
hp
MW
kWh/t
kV/kVA
kPa/MPa
troy ounces
avoirdupois pound
short ton
ounces per short ton
metric ton (tonne)
millions of tonnes
tonnes milled
tonnes moved
tonnes mined
metric tons per day/year
grams per metric ton
parts per billion
parts per million
horsepower
megawatts
kilowatt-hours per tonne
kilovolts/kilovolt-amps
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%)
CDC
2429680
32D01
Active
57.37
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429681
32D01
Active
57.37
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429682
32D01
Active
57.37
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429683
32D01
Active
34.65
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429684
32D01
Active
29.92
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429685
32D01
Active
33.92
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429686
32D01
Active
4.57
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429687
32D01
Active
7.27
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429688
32D01
Active
14.76
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429689
32D01
Active
23.71
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429690
32D01
Active
29.69
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429691
32D01
Active
49.52
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429692
32D01
Active
19.99
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429693
32D01
Active
6.65
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429694
32D01
Active
24.02
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429695
32D01
Active
24.12
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429696
32D01
Active
24.75
July 30, 2015
March 1, 2017
Radisson Mining Resources Inc. (100%)
CDC
2429697
32D01
Active
32.13
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
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
O'Brien
Elmac
Elmac
CDC
CDC
CDC
CDC
CDC
CDC
CDC
CDC
CDC
CDC
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
CL
2169717
2169718
2169719
2169720
2169721
2169722
2169723
2169724
2169725
2169726
3295141
3350491
3350492
3350493
3350494
3350495
3350501
3350502
3350504
3350505
3350511
3350512
3350513
3350514
3350515
3350521
3350522
3350523
3350524
4261152
5274288
5274289
5274290
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
32D01
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Cadillac
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
Active
12.33
35.61
26.71
17.45
3.91
6.80
19.90
26.42
36.85
8.62
81.47
17.47
14.55
16.26
14.64
15.92
16.42
8.92
9.86
6.87
6.78
9.76
11.47
9.76
6.93
6.55
6.96
13.98
4.80
9.84
16.00
16.36
8.36
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Radisson Mining Resources Inc. (100%)
Kewagama
CL
C005451
32D01
Cadillac
Active
21.17
Radisson Mining Resources Inc. (100%)
Kewagama
CL
C006763
32D01
Cadillac
Active
7.88
Radisson Mining Resources Inc. (100%)
Kewagama
CLD
P007770
32D01
Cadillac
Active
83.02
Radisson Mining Resources Inc. (100%)
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
No Royalty
2% NSR to KWG
Resources Inc.
2% NSR to KWG
Resources Inc.
2% NSR to KWG
Resources Inc.
43-101 Technical Report – O’Brien Project
211
www.innovexplo.com
APPENDIX V – SURFACE PLANS
43-101 Technical Report – O’Brien Project
212
www.innovexplo.com
APPENDIX VI – ENVIRONMENTAL CHARACTERIZATION
43-101 Technical Report – O’Brien Project
215
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
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 i
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 du gisement O’Brien - PU-2013-12-860
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 :
Tableau 2 :
Tableau 3 :
Tableau 4 :
Tableau 5 :
Tableau 6 :
Tableau 7 :
Tableau 8 :
Tableau 9 :
Tableau 10 :
Tableau 11 :
Tableau 12 :
Tableau 13 :
Liste d'échantillons utilisés pour réaliser le lot composite utilisé pour
les essais métallurgiques .......................................................................................... 2
Synthèse des essais d’évaluation de la proportion d’or libre récupérable .............. 5
Résumé des résultats obtenus des essais F-1 à F-4 ................................................. 7
Bilan (Au) métallurgique de l’essai KN-F-1 ............................................................... 8
Bilan métallurgique de l’essai KN-F-2..................................................................... 11
Bilan métallurgique (Au) de l’essai KN-F-3 ............................................................. 13
Bilan métallurgique (As) de l’essai KN-F-3 ............................................................. 13
Bilan métallurgique de l’essai cyclique KN-F-5....................................................... 17
Bilan métallurgique de l’essai KN-F-6-R ................................................................. 19
Tableau comparatif des résultats des essais KN-F-3, KN-F-5 et KN-F-6-R.............. 20
Bilan métallurgique de l’essai KN-CN-F-4 avec K80 = 102µ à la cyanuration ........ 22
Récupération et consommation en réactifs de la cyanuration uniquement
(CN-F-4)................................................................................................................... 22
Bilan métallurgique de l’essai KN-CN-2 avec K80 = 37µ à la cyanuration.............. 24
Page ii
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 du gisement O’Brien - PU-2013-12-860
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 :
Figure 2 :
Figure 3 :
Figure 4 :
Figure 5 :
Figure 6 :
Figure 7 :
Figure 8 :
Figure 9 :
Figure 10 :
Figure 11 :
Figure 12 :
Figure 13 :
Figure 14 :
Localisation du site O’Brien de Ressources minières Radisson ............................... 1
Diviseur rotatif utilisé pour la division en lots homogènes...................................... 3
Schéma de concassage des échantillons .................................................................. 3
Schéma expérimental utilisé pour l’évaluation de la proportion d’or libre............. 4
Montage expérimental pour l’évaluation de la proportion d’or libre ..................... 4
Graphique de la récupération gravimétrique de l’or vs la granulométrie
de broyage ................................................................................................................ 5
Photographie de l’or libre sur la table de Mozley (Essai KN-4) ................................ 6
Protocole de flottation utilisé pour les essais F-1 à F-4 ........................................... 6
Graphique de la récupération en or et en argent, en fonction de la
granulométrie de broyage ........................................................................................ 7
Protocole expérimental de l’essai KN-F-1 ................................................................ 9
Protocole expérimental de l’essai KN-F-2 .............................................................. 10
Protocole expérimental de l’essai KN-F-3 .............................................................. 12
Protocole expérimental de l’essai KN-F-4 .............................................................. 15
Protocole de l’essai cyclique KN-F-5 ...................................................................... 16
Page iii
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 du gisement O’Brien - PU-2013-12-860
Figure 15 :
Figure 16 :
Figure 17 :
Figure 18 :
Figure 19 :
Protocole expérimental de l’essai cyclique KN-F-6-R............................................. 18
Protocole de l’essai KN-CN-F-4 ............................................................................... 21
Protocole expérimental de l’essai KN-CN-2 ........................................................... 23
Photographies au microscope optique montrant les trois statuts de l’or ............. 41
Représentation graphique des teneurs en or dans l’arsénopyrite dans
l’échantillon «concentré de flottation» ................................................................ 42
Page iv
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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 caractérisations 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 1
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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 RouynNoranda.
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 2
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
Figure 2 : Diviseur rotatif utilisé pour la division en lots homogènes
Séchage 40o C
Concasseur
à mâchoires
Concasseur
à rouleaux
1re division avec
l’échantillonneur
rotatif
2e division avec
l’échantillonneur
rotatif
Division en lots de
0,5 kg
Figure 3 : Schéma de concassage des échantillons
Page 3
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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.
Evaluation de la proportion d'or libre récupérable
Broyage
Pompe Masterflex
Knelson
concentré
Table de Mozzley
Rejet table de Mozzley
Rejet Knelson
Concentré or libre
Figure 4 : Schéma expérimental utilisé pour l’évaluation de la proportion d’or libre
Pompe
Masterflex
Figure 5 : Montage expérimental pour l’évaluation de la proportion d’or libre
Page 4
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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
Au
Description
Ag
Teneur
Teneur
Alim.
Alim.
du conc. Rejet
du conc. Rejet
calc. Réc. Ag
calc.
or libre g Au/mt
or libre g Ag/tm
g Au/tm
g Ag/tm
(g Au/tm)
(g Au/tm)
Essai
K80
(µm)
% < 200
mailles
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
Evaluation
proportion
d'or libre
récupérable
Réc.
Au
Moyenne:
11,96
0,90

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.
% or libre récupéré Vs K80
Zone 36-Est - Ressources Radisson
100%
90%
80%
% or libre récupéré
70%
60%
50%
40%
30%
20%
10%
0%
30
40
50
60
70
80
90
100
K80 (µm)
110
120
130
140
150
Figure 6 : Graphique de la récupération gravimétrique de l’or vs la granulométrie de broyage
Page 5
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860

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:
Conditionnement
5 min
pH nat = 8,2
250 g/tm CuS O 4
15 g/tm A-407
20 g/tm KAX
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
13 g/tm MIBC
1,0min
1,15mi
3,0min
Conc. 1
Conc. 2
Conc. 3
5 g/tm A-407
10 g/tm KAX
13 g/t MIBC
2,0min
Rejet
Conc. 4
Figure 8 : Protocole de flottation utilisé pour les essais F-1 à F-4
Page 6
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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
Description
Essais de
flottation
%
% < 200 massique
mailles
au
concentré
Au
Ag
Teneur
Teneur
Alim.
Alim.
moy. du Rejet
moy. du
Rejet
calc. Réc. Ag
calc.
conc. g Au/mt
conc.
g Ag/tm
g Au/tm
g Ag/tm
(g Au/tm)
(g Ag/tm)
Essai
K80
(µm)
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
Réc.
Au
Moyenne:
10,22
1,17

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.
Essais de flottation F-1 à F-4
% récupération Vs K80
Zone 36-Est - Ressources Radisson
100%
90%
Au
80%
% récupération
70%
60%
50%
Ag
40%
30%
20%
10%
0%
30
40
50
60
70
80
90
100
K80 (µm)
110
120
130
140
150
Figure 9 : Graphique de la récupération en or et en argent, en fonction de la granulométrie de broyage
Page 7
Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860

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
Concentré d'or libre:
0,28
0,03%
19700 g Au/tm
5,52 mg Au
55,3%
Concentré flottation
86,20
8,69%
44,27 g Au/tm
3,82 mg Au
38,3%
Rejet flottation
905,7
91,28%
0,71 g Au/tm
0,64 mg Au
6,4%
Alimentation calculée
992,18
100%
10,05 g Au/tm
9,97
100%
Récupération or dans concentré or libre:
Récupération or dans concentré flottation:
Récupération or combinée:
55,3%
38,3%
93,6%

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.
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Essai:
KN-F-1
Description:
Essai Knelson + flottation (sans rebroyage avant la flottation)
Broyage: :
992,2 g
20 kg de tiges
6' 45"
50% solide
1,0 kg de
minerai/broyage
K80 = 102 µ
(66,5% < 200 mailles)
Pompe Masterflex
Knelson
13,1 kPa
concentré
Rejet Knelson
Table de Mozzley
Rejet table de Mozzley
0,28 g
19700 g Au/tm
Concentré or libre
% or récupéré = 55,3%
Conditionnement
5 min
pH nat = 8,2
250 g/tm CuS O 4
15 g/tm A-407
20 g/tm KAX
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
13 g/tm MIBC
1,0min
1,15mi
2,0min
5 g/tm A-407
10 g/tm KAX
13 g/t MIBC
2,0min
S-10 et S-16
S -11 et S -18
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.
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Il faut également noter que cet essai ne comporte pas d’étape de nettoyage.
Essai:
KN-F-2
Description:
Essai Knelson + flottation (avec rebroyage avant la flottation)
Broyage: :
997,2 g
20 kg de tiges
6' 45"
50% solide
1,0 kg de
minerai/broyage
K80 = 102 µ
(66,5% < 200 mailles)
Pompe Masterflex
Knelson
13,1 kPa
concentré
Rejet Knelson
Table de Mozzley
Rejet table de Mozzley
0,55 g
10263 g Au/tm
Concentré or libre
% or récupéré = 54,3%
Broyage: 2' 00"
K80 = 73µ
Conditionnement
5 min
pH nat = 8,2
250 g/tm CuS O 4
15 g/tm A-407
20 g/tm KAX
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
13 g/tm MIBC
1,0min
1,15mi
2,0min
5 g/tm A-407
10 g/tm KAX
13 g/t MIBC
2,0min
S-13 et S-17
S -14 et S -19
Figure 11 : Protocole expérimental de l’essai KN-F-2
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Tableau 5 : Bilan métallurgique de l’essai KN-F-2
Masse (g)
% poids
Teneur
mg Au
% distribution
Concentré d'or libre:
0,55
0,06%
10263 g Au/tm
5,64 mg Au
54,3%
Concentré flottation
93,20
9,35%
43,65 g Au/tm
4,07 mg Au
39,1%
Rejet flottation
903,4
90,60%
0,75 g Au/tm
0,68 mg Au
6,5%
Alimentation calculée
997,15
100%
10,42 g Au/tm
10,39
100%
Récupération or dans concentré or libre:
Récupération or dans concentré flottation:
Récupération or combinée:
54,3%
39,1%
93,5%

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
3.3.1
Essais combinant la concentration gravimétrique et la flottation
avec étapes de nettoyage
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.
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Essai: KN-F-3
Date: 17-mai-14
Description: Essai Knelson + flottation (avec étapes de nettoyage)
Broyage: :
994,0 g
20 kg de tiges
7' 00"
50% solide
1,0 kg de
minerai/broyage
K80 = 102 µ
(65,8% < 200 mailles)
Pompe Masterflex
Knelson
13,1 kPa
concentré
Rejet Knelson
Table de Mozzley
Rejet table de Mozzley
0,29 g
22947 g Au/tm
Concentré or libre
% or récupéré = 62,9%
Conditionnement
5 min
pH nat = 8,2
250 g/tm CuS O 4
15 g/tm A-407
20 g/tm KAX
13 g/tm MIBC
1,5min
1,0min
S-31
20,0% As
156,7 g Au/tm
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
1,5min
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
2,0min
5 g/tm A-407
10 g/tm KAX
13 g/t MIBC
2,0min
2,0min
S-30 et S-34
0,03% As
0,57 g Au/tm
S-33
0,30% As
1,5 g Au/tm
S-35
5,51% As
27,0 g Au/tm
Figure 12 : Protocole expérimental de l’essai KN-F-3
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Tableau 6 : Bilan métallurgique (Au) de l’essai KN-F-3
Masse (g)
% poids
Teneur
mg Au
% distribution
Concentré d'or libre:
0,29
0,03%
22947 g Au/tm
6,65 mg Au
62,9%
Concentré 1er nettoyage
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%
Rejet nettoyeur épuiseur
45,30
4,56%
1,5 g Au/tm
0,07 mg Au
0,7%
Rejet épuiseur
911,9
91,74%
0,57 g Au/tm
0,52 mg Au
4,9%
Rejets combinés
957,2
96,30%
0,61 g Au/tm
0,58 mg Au
5,5%
Alimentation calculée
993,99
100%
10,64 g Au/tm
10,57 mg Au
100%
Concentré nettoyeur
épuiseur
Concentrés flottation
combinés
Récupération or dans concentré or libre:
Récupération combinée des conc. 1er nettoyeur et conc. nettoyeur épuiseur:
Récupération combinée or libre + concentrés flottation :
62,9%
31,5%
94,5%

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%
Rejet nettoyeur épuiseur
45,30
4,56%
0,30%
0,14 g As
2,7%
Rejet épuiseur
911,9
91,77%
0,03%
0,32 g As
6,2%
Rejets combinés
957,2
96,33%
0,05%
0,45 g As
8,8%
Alimentation calculée
993,70
100%
0,52%
5,12 g As
100%
Concentré 1er nettoyage
Concentré nettoyeur
épuiseur
Concentrés flottation
combinés

On observe tout d’abord que la teneur calculée de l’alimentation se situe à 0,52 % As.
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
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.
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Essai: KN-F-4
Date: 24-juin-14
Description: Essai Knelson + flottation (avec étapes de nettoyage)
Broyage: :
992,7 g
20 kg de tiges
7' 00"
50% solide
1,0 kg de
minerai/broyage
K80 = 102 µ
(65,8% < 200 mailles)
Pompe Masterflex
Knelson
13,1 kPa
concentré
Rejet Knelson
Table de Mozzley
Rejet table de Mozzley
0,22 g
27095 g Au/tm
Concentré or libre
% or récupéré = 63,1%
Conditionnement
5 min
pH nat = 8,2
250 g/tm CuS O 4
15 g/tm A-407
20 g/tm KAX
13 g/tm MIBC
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
1,5min
1,5min
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
2,0min
47,8 g Au/tm
1er nettoyage
S-31-B
5,1 g Au/tm
1,0min
S-30-B
119,7 g Au/tm
5 g/tm A-407
10 g/tm KAX
13 g/t MIBC
2,0min
4,2 g Au/tm
Nettoyeur-épuiseur
2,0min
S-34-B
0,55 g Au/tm
Rejets combinés
0,57 g Au/tm
S-33-B
1,30 g Au/tm
S-32-B
13,7 g Au/tm
Figure 13 : Protocole expérimental de l’essai KN-F-4
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Essai: KN-F-5
Date: 25-juin-14
Description: Essai Knelson + flottations cycliques (Cycle 4)
Broyage: :
20 kg de tiges
7' 00"
50% solide
1,0 kg de
minerai/broyage
4 kg
K80 = 102 µ
(65,8% < 200 mailles)
Pompe Masterflex
Knelson
13,1 kPa
concentré
Rejet Knelson
Table de Mozzley
Rejet table de Mozzley
0,63 g
62143 g Au/tm
Concentré or libre
% or récupéré = 67,4%
Conditionnement
5 min
pH nat = 8,2
250 g/tm CuS O 4
15 g/tm A-407
20 g/tm KAX
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
13 g/tm MIBC
1,5min
1,5min
2,0min
5 g/tm A-407
10 g/tm KAX
13 g/t MIBC
2,0min
Rejet 1
0,02% As
0,82 g Au/mt
Nettoyeur-épuiseur
1er nettoyage
S-48
0,51% As
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
2,0min
1,5min
3,58 g Au/mt
Concentré final
7,26% As
94,5 g Au/tm
Rejet 2
0,17% As
6,79 g Au/mt
S-47
1,75% As
114,3 g Au/tm
Figure 14 : Protocole de l’essai cyclique KN-F-5
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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
962,86 g
964,06 g
994,76 g
14,96 g Au/mt
15,68 g Au/mt
14,59 g Au/mt
0,11% As
0,35% As
0,32% As
% distribution Au
concentré
flottation
26,4%
29,6%
26,2%
% distribution or
libre récupéré
67,9%
64,8%
67,4%
% récupération
globale Au
94,3%
94,4%
93,6%
Concentré or libre
Concentré final
Rejet 1 + Rejet 2
0,02% As
Alim. calc.
0,24% As

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.
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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-R
Date: 06-juil-14
Description: Reprise de l'essai KN-F-6 :Essai Knelson + flottations cycliques (Cycle 4)
Broyage: :
20 kg de tiges
7' 00"
50% solide
1,0 kg de
minerai/broyage
4 kg
K80 = 102 µ
(65,8% < 200 mailles)
Pompe Masterflex
Knelson
13,1 kPa
concentré
Rejet Knelson
Table de Mozzley
Rejet table de Mozzley
0,69 g
38114 g Au/tm
Concentré or libre
% or récupéré = 60,2%
Conditionnement
5 min
pH nat = 8,2
250 g/tm CuS O 4
15 g/tm A-407
20 g/tm KAX
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
13 g/tm MIBC
1,5min
1,5min
10 g/tm A-407
15 g/tm KAX
13 g/t MIBC
2,0min
2,0min
1er nettoyage
1,5min
Concentré final
12,09% As
96,1 g Au/tm
5 g/tm A-407
10 g/tm KAX
13 g/t MIBC
Rejet 1
0,02% As
0,58 g Au/mt
Nettoyeur-épuiseur
2,0min
Rejet final
0,03% As
0,62 g Au/mt
Rejet 2
0,13% As
1,06 g Au/mt
1,61% As
15,9 g Au/mt
Figure 15 : Protocole expérimental de l’essai cyclique KN-F-6-R
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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
116,8 g Au/mt
138,5 g Au/mt
88,2 g Au/mt
96,1 g Au/mt
13,96% As
13,4% As
12,5% As
12,1% As
891,80 g
951,20 g
947,70 g
951,00 g
0,53 g Au/mt
0,92 g Au/mt
0,93 g Au/mt
0,62 g Au/mt
0,02% As
0,04% As
0,03% As
0,03% As
923,40 g
984,50 g
988,10 g
990,10 g
10,7 g Au/mt
12,26 g Au/mt
11,15 g Au/mt
11,03 g Au/mt
0,50% As
0,49% As
0,54% As
0,51% As
% distribution Au
concentré
flottation
34,4%
38,2%
32,4%
34,4%
% distribution or
libre récupéré
61,2%
54,5%
59,7%
60,2%
% récupération
globale Au
95,6%
92,7%
92,0%
94,6%
Concentré or libre
Concentré final
Rejet 1 + Rejet 2
Alim. calc.

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.
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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
91,3 g Au/mt
94,5 g Au/mt
96,1 g Au/mt
12,79% As
7,3% As
12,1% As
957,20 g
929,90 g
0,00 g
0,61 g Au/mt
0,97 g Au/mt
0,6 g Au/mt
0,05% As
0,02% As
0,03% As
993,99 g
994,76 g
990,10 g
10,6 g Au/mt
14,59 g Au/mt
11,0 g Au/mt
0,52% As
0,32% As
0,51% As
% distribution Au
concentré
flottation
31,5%
26,2%
34,4%
% distribution or
libre récupéré
62,9%
67,4%
60,2%
% récupération
globale Au
94,5%
93,6%
94,6%
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 KNF-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.
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
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:
Description:
KN-CN-F-4
Récupération gravimétrique de l'or suivie de la cyanuration du rejet
gravimétrique
Broyage:
20 kg de tiges
7' 00"
1,0 kg de minerais
50% solide
K80 = 102 µ
Pompe Masterflex
Knelson
concentré
13 kPa
Rejet Knelson
Nettoyage sur table Mozley
Cyanuration
48h
Rejet de table de Mozley
concentré
d'or libre
% or libre récupéré
58,0%
Solution
Rejet solide
30508 g Au/tm Récupération globale en or = 87,9%
Figure 16 : Protocole de l’essai KN-CN-F-4
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Tableau 11 : Bilan métallurgique de l’essai KN-CN-F-4 avec K80 = 102µ à la cyanuration
Concentré or libre
Masse (g)
% poids
Teneur
mg Au
% distribution
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%
Solution après 25 h de
cyanuration
Solution après 34 h de
cyanuration
Solution après 48 h de
cyanuration
Rejet solide de cyanuration
983,1
99,98%
1,17 g Au/tm
1,15 mg Au
10,4%
Alimentation calculée
983,3
100%
11,23 g Au/tm
11,04 mg Au
100%
Récupération globale en or après 25 h
Récupération globale en or après 34 h
Récupération globale en or après 48 h
% or libre récupéré
=
=
=
=
87,9%
88,6%
89,6%
58,0%
Consommation en cyanure de sodium = 0,33 kg NaCN / tm
Consommation en chaux hydratée = 2,08 kg Ca(OH)2 / tm

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)
Durée
Récupération
Au
NaCN kg/tm
de minerai
Ca(OH)2 kg/tm de
minerai
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

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.
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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:
Description:
KN-CN-F-2
Récupération gravimétrique de l'or suivie de la cyanuration du rejet
gravimétrique avec broyage du rejet Knelson
Broyage:
20 kg de tiges
7' 00"
1,0 kg de minerais
50% solide
K80 = 102 µ
Pompe Masterflex
Knelson
concentré
Nettoyage sur table Mozley
13 kPa
Rejet Knelson
K80 = 37 µ
Cyanuration
48 h
Rejet de table de Mozley
concentré
d'or libre
% or libre récupéré
60,8%
Solution
Rejet solide
25598 g Au/tm Récupération globale en or = 92,9%
Figure 17 : Protocole expérimental de l’essai KN-CN-2
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Tableau 13 : Bilan métallurgique de l’essai KN-CN-2 avec K80 = 37µ à la cyanuration
Masse (g)
% poids
Teneur
mg Au
% distribution
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%
Concentré or libre
Solution après 25,6 h de
cyanuration
Solution après 38,3 h de
cyanuration
Solution après 48 h de
cyanuration
Rejet solide de cyanuration
994,3
99,97%
0,84 g Au/tm
0,84 mg Au
7,1%
Alimentation calculée
994,6
100%
11,86 g Au/tm
11,79 mg Au
100%
Récupération globale en or après 25,6 h
Récupération globale en or après 38,3 h
Récupération globale en or après 48 h
% or libre récupéré
=
=
=
=
91,2%
92,3%
92,9%
60,8%
Consommation en cyanure de sodium = 0,49 kg NaCN / tm
Consommation en chaux hydratée = 3,19 kg Ca(OH)2 / tm
Tableau 14 : Récupération et consommation en réactifs de la cyanuration uniquement (KN-CN-F-2)
Durée
Récupération
Au
NaCN kg/tm
de minerai
Ca(OH)2 kg/tm de
minerai
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

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 pourraient 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
Composite Éponte 6
Composite Éponte 4
Composite Épontes 2 + 5
Composite Épontes 1 + 3
Nom échantillon brut
#URSTM
Volc Centre = #3 (45236)
32145
Volc Centre = #3 (45237)
32146
Volc Sud Éponte #1A (49918)
32147
Volc Sud #1B (49934)
32148
Volc Sud #1B (49935)
32149
Porph Nord (45221)
32150
Porph Nord (45222)
32151
Porph Nord (45223)
32152
Porph Sud (45248)
32153
Porph Sud (45249)
32154
WCgl (45228)
32155
WCgl (45229)
32156
WCgl (45230)
32157
WCgl (45231)
32158
W-Volc-Sud N (45204)
32159
W-Volc-Sud N (45205)
32160
W-Volc-Sud N (45206)
32161
W-Volc-Sud N (45207)
32162
W-Volc-Sud N (45208)
32163
W-Volc-Sud N (45209)
32164
Remarques
Concassé passant ~ 1 cm
Masse approx: 2.85 kg
Concassé passant ~ 1 cm
Masse approx: 2.5 kg
Concassé passant ~ 2 mm
Masse approx: 2.4 kg
Concassé passant ~ 2 mm
Masse approx: 2 kg
Concassé passant ~ 2 mm
Masse approx: 2,07 kg
Concassé passant ~ 1 cm
Masse approx: 2,06 kg
Concassé passant ~ 1 cm
Masse approx: 1,54 kg
Concassé passant ~ 1 cm
Masse approx: 2,51 kg
Concassé passant ~ 1 cm
Masse approx: 2,86 kg
Concassé passant ~ 1 cm
Masse approx: 2,65 kg
Concassé passant ~ 1 cm
Masse approx: 2,66 kg
Concassé passant ~ 1 cm
Masse approx: 2,8 kg
Concassé passant ~ 1 cm
Masse approx: 1,85 kg
Concassé passant ~ 1 cm
Masse approx: 2.75 kg
Concassé passant ~ 1 cm
Masse approx: 1 kg
Concassé passant ~ 1 cm
Masse approx: 1,58 kg
Concassé passant ~ 1 cm
Masse approx: 2,32 kg
Concassé passant ~ 1 cm
Masse approx: 2,2 kg
Concassé passant ~ 1 cm
Masse approx: 2,87 kg
Concassé passant ~ 1 cm
Masse approx: 1,86 kg
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Tableau 16 : Différentes moutures des échantillons pour les essais environnementaux
#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 µm
Analyses chimiques, Essais
statiques
Composite Épontes 2 + 5 pulv
32170
< 200 µm
Analyses chimiques, Essais
statiques
Composite Éponte 4 pulv
32171
< 200 µm
Analyses chimiques, Essais
statiques
Composite Éponte 6 pulv
32172
< 200 µm
Analyses chimiques, Essais
statiques
Rejet de flottation
34290
< 200 µm
TCLP, Analyses chimiques,
Essais statiques
Nom
2. Méthodes
2.1
Caractérisations chimiques des solides
Pour évaluer les éléments majeurs, les échantillons ont été envoyés dans un laboratoire soustraitant 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 ICPAES;
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 cidessous ».
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, S total, 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 environnemental, 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 échantillons.
Selon les analyses chimiques, les échantillons semblent principalement composés d’aluminosilicates 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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Roche totale
Tableau 17 : Analyses chimiques réalisées sur les échantillons de l’étude environnementale
Paramètre
Unités
LDM
Composite
Éponte 1+3
pulv.
U32169
Composite
Éponte 2+5
pulv.
U32170
Composite
Éponte 4 pulv.
Composite
Éponte 6 pulv.
Rejet de
flottation
U32171
U32172
U34290
Fe 2O3
% 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
MgO
K2O
CaO
P2O5
MnO
TiO2
% p/p
0,01
% p/p
0,01
% p/p
0,01
% p/p
0,01
% p/p
0,01
% p/p
0,01
% p/p
0,01
2,37
6,27
1,77
8,29
0,14
0,20
0,85
4,56
3,29
1,74
3,86
0,16
0,08
0,35
2,82
2,88
2,60
3,35
0,16
0,16
0,61
1,69
5,54
1,53
8,29
0,13
0,17
0,74
3,05
4,04
2,33
5,16
0,14
0,12
0,49
Cr2O3
% p/p
0,01
0,02
0,04
0,03
0,02
0,05
V2O5
% p/p
LOI
% p/p
Mass Balance % p/p
0,01
0,04
6,78
100,35
0,01
4,31
99,84
0,02
3,00
98,70
0,03
8,56
98,62
0,02
5,59
97,74
Ag 1
Al
As
mg/kg
2
mg/kg
60
mg/kg
30
<2
66130
88,9
<2
47300
1690
<2
62380
382
<2
62760
257
<2
68940
251
mg/kg
0,01
mg/kg
5
mg/kg
1
mg/kg
30
10,0
266
<1
<30
60960
<5
74,3
187
104
78660
10,5
753
<1
<30
25360
<5
28,7
497
32,9
27080
7,2
542
<1
<30
23830
<5
52,3
401
77,2
54590
3,0
218
<1
<30
59550
<5
61,7
217
93,8
60780
<0,01
588
<1
<30
41430
<5
27
301
72
39220
<0,01
32820
1274
58,7
82,9
<5
<4
<3
<5
4389
<2
4677
<55
0,10
8991
496
38,4
204
<5
30,8
<3
<5
3815
<2
1227
<55
<0,01
11960
908
50,2
166
<5
10
<3
<5
5563
<2
3174
<55
<0,01
27810
1026
53,3
86,1
<5
<4
<3
<5
4977
<2
3026
<55
0,02
20420
775
48
120
<5
<4
<3
<5
3558
<2
9359
<55
Analyses chimiques
1
1
B
Ba
Be
Bi
Ca
Cd
Co
Cr
Cu
Fe
1
Hg
Mg
Mn
Mo
Ni
Pb
Sb
Se
Sn
Stotal (ICP)
Te
Ti
Zn
0,01
0,01
mg/kg
60
mg/kg
5
mg/kg
5
mg/kg
5
mg/kg
10
mg/kg
10
mg/kg
0,01
mg/kg
15
mg/kg
5
mg/kg
5
mg/kg
5
mg/kg
5
mg/kg
4
mg/kg
3
mg/kg
5
mg/kg
200
mg/kg
2
mg/kg
25
mg/kg
55
: l a méthode de di ges ti on pour ces él éments doi t être cons i dérée pa rti el l e, ma i s conforme à MA. 200 – Mét. 1.2
n/a : non a na l ys é
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Tableau 18 : Comparaison des résultats des analyses chimiques avec les critères de la PPSRTC
Éléments
Argent (Ag)
Arsenic (As)
Baryum (Ba)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Étain (Sn)
Manganèse (Mn)
Mercure (Hg)
Molybdène (Mo)
Nickel (Ni)
Plomb (Pb)
Selenium (Se)
Zinc (Zn)
Unités
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
Échantillon
Éponte 1+3 Éponte 2+5 Éponte 4 Éponte 6
Critère A* Critère B* Critère C*
Rejet
fottation
U32169
U32170
U32171
U32172
U34290
<2
88,9
266
<5
187
74,3
104
<5
1274
<0,01
58,7
82,9
<5
<3
<55
<2
1690
753
<5
497
28,7
32,9
<5
496
0,1
38,4
204
<5
<3
<55
<2
382
542
<5
401
52,3
77,2
<5
908
<0,01
50,2
166
<5
<3
<55
<2
257
218
<5
217
61,7
93,8
<5
1026
<0,01
53,3
86,1
<5
<3
<55
<2
251
588
<5
301
27
72
<5
775
0,02
48
120
<5
<3
<55
0,5
5
200
0,9
85
20
50
5
1000
0,3
6
50
40
3
120
20
30
500
5
250
50
100
50
1000
2
10
100
500
3
500
40
50
2000
20
800
300
500
300
2200
10
40
500
1000
10
1500
*Cri tères de l a Pol i tique de protection des s ol s et de réha bi l i tation des terra i ns contami nés pour l a provi nce du s upéri eur
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.
U32169
Composite
Éponte 2+5
pulv.
U32170
Composite
Composite
Éponte 4 pulv. Éponte 6 pulv.
U32171
U32172
Rejet de
flottation
U34290
Rejet
cyanuration
(S-21 S-24 )
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
PN 2
PNN
PN/PA
Acidogène 3
1
kg Ca CO3/t n/d
10,4
13,7
17,4
13,8
8,6
30,5
kg Ca CO3/t n/d
74,9
61
5,5
44,9
28
2,6
150
136
10,9
94,3
85,7
10,9
112
81,5
3,7
Non
Oui
Non
Non
Non
-
-
115
105
11,1
-
-
Non
kg Ca CO3/t n/d
: l a ma tri ce des s ul fures a été touchée pa r l 'extra ction des s ul fa tes , l e PAM es t donc ca l cul é à pa rtir du S total
2
: La l i mi te de l 'es s a i de La wrence et Wa ng 1997 es t de 150 kgCa CO 3/t, i l s e pourra i t que l e PN véri tabl e s oi t
pl us él evé que l a va l eur ra pportée da ns ce ca s .
3
: s el on l es cri tères de l a Di rective 019, ma rs 2012
n/d : non détermi né
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 ».
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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
Solution #
pH initial
pH final
Aluminium
Antimoine 1
Argent
Arsenic
Baryum
Béryllium
Bismuth
Bore
Cadmium
Calcium
Chrome
Cobalt
Cuivre
Fer
Magnésium
Manganèse
Mercure
Molybdène
Nickel
Plomb
Potassium
Sélénium
Sodium2
Titane
Zinc
Sulfates
LDR
-
-
-
pH
-
-
pH
-
-
(Al)
mg/l
0,01
(Sb)
mg/l
0,09
(Ag)
mg/l
0,0001
(As)
mg/l
0,06
(Ba)
mg/l
0,001
(Be)
mg/l
0,001
(Bi)
mg/l
0,02
(B)
mg/l
0,01
(Cd)
mg/l
0,003
(Ca)
mg/l
0,03
(Cr)
mg/l
0,003
(Co)
mg/l
0,004
(Cu)
mg/l
0,003
(Fe)
mg/l
0,006
0,001
(Mg)
mg/l
(Mn)
mg/l
0,002
(Hg)
mg/l
0,00001
(Mo)
mg/l
0,009
(Ni)
mg/l
0,004
(Pb)
mg/l
0,02
(K)
mg/l
0,05
(Se)
mg/l
0,1
(Na)
mg/l
0,05
(Ti)
mg/l
0,002
(Zn)
mg/l
0,005
(SO 4 2-)
mg/l
0,6
Blanc
solution 1
Blanc
solution 2
Composite
Éponte 1+3
grossier
Composite
Éponte 2+5
grossier
Composite
Éponte 4
grossier
Composite
Éponte 6
grossier
Rejets
Flottation
Rejets
Flottation
(double)
-
-
U32165
U32166
U32167
U32168
U34290
U34290 (d)
1
4,88
4,9
0,099
0,752
<0,0001
<0,06
0,072
<0,001
<0,02
<0,01
<0,003
0,783
<0,003
<0,004
0,007
0,081
0,017
0,003
0,00007
0,012
<0,004
<0,02
0,51
<0,1
1280
<0,002
0,057
22,8
2
2,90
2,89
0,051
0,647
<0,0001
<0,06
0,074
<0,001
<0,02
<0,01
<0,003
0,126
<0,003
<0,004
0,009
0,060
0,013
<0,002
<0,00001
0,011
<0,004
<0,02
<0,05
<0,1
0,430
<0,002
0,062
14,8
2
2,90
5,64
<0,01
0,167
<0,0001
<0,06
0,017
<0,001
<0,02
<0,01
<0,003
1850
<0,003
<0,004
<0,003
17,8
12,9
21,3
0,00057
<0,009
<0,004
<0,02
54,4
<0,1
1,70
<0,002
0,119
14,5
1
4,88
6,15
0,123
0,399
0,0017
<0,06
1,09
<0,001
<0,02
0,200
<0,003
729
<0,003
<0,004
0,008
5,17
7,54
11,6
0,00033
<0,009
0,012
<0,02
21,7
<0,1
1502
<0,002
0,132
10,4
1
4,88
5,55
0,338
0,550
<0,0001
0,083
0,773
<0,001
<0,02
<0,01
<0,003
624
0,013
0,005
0,004
8,52
3,66
8,41
0,00009
<0,009
0,017
<0,02
28,7
<0,1
1414
<0,002
0,167
<0,6
1
4,88
6,34
0,102
0,555
0,0020
<0,06
0,284
<0,001
<0,02
0,040
<0,003
956
<0,003
0,006
0,003
0,217
4,20
8,23
0,00008
<0,009
0,028
<0,02
7,70
<0,1
1516
<0,002
0,148
7,90
2
2,90
5,18
0,580
0,697
<0,0001
<0,06
0,575
<0,001
<0,02
<0,01
<0,003
1540
0,005
0,009
0,531
13,4
33,2
20,5
<0,00001
<0,009
0,033
<0,02
25,8
<0,1
1,40
<0,002
0,271
14,0
2
2,90
5,18
0,588
<0,09
<0,0001
<0,06
0,628
<0,001
<0,02
<0,01
<0,003
1550
0,006
0,006
0,536
12,6
33,5
21,0
<0,00001
<0,009
0,033
<0,02
26,1
<0,1
1,80
<0,002
0,290
15,2
Critères d'eau
Critères de
souterraine
résurgences
aux fins de
dans les eaux
consommation
de surface
s/o
s/o
s/o
s/o
0,006
0,1
0,025
1
s/o
s/o
s/o
0,005
s/o
0,05
s/o
1*
s/o
s/o
0,05*
0,001
0,07
0,020
0,01
s/o
0,01
200
s/o
5*
s/o
s/o
s/o
s/o
0,75
s/o
0,00062
0,34
5,3
s/o
s/o
s/o
0,00021
s/o
0,2
0,5
0,0073
s/o
s/o
s/o
0,00013
2
0,26
0,034
s/o
0,02
s/o
s/o
0,0167
s/o
Résidus
miniers à
risques élevés
(Directive 019)
s/o
s/o
s/o
s/o
s/o
s/o
5,0
100
s/o
s/o
500
0,5
s/o
5,0
s/o
s/o
s/o
s/o
s/o
0,1
s/o
s/o
5,0
s/o
1,0
s/o
s/o
s/o
s/o
s /o : s a ns objet da ns l a PPSRTC et/ou l a Di recti ve 019
*: Des objecti fs d’ordre es théti ques s ont di s poni bl es pour certa i ns pa ra mètres .
1
: Une conta mi na ti on en Anti moi ne da ns l es s ol uti ons d'extra cti on nous empêche d'uti l i s er l es rés ul ta ts des a na l ys es .
2
: Le s odi um entre da ns l a compos i ti on de l a s ol uti on #1 pour l es extra cti ons TCLP.
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Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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.
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Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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.
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Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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 caractérisation 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 ICPAES 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
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Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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
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de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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.
A
B
C
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.
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de la zone 36 Est du gisement O’Brien - PU-2013-12-860
500
450
400
Teneur en Au (ppm)
350
300
250
200
150
Limite de
détection
de la MSE
100
50
0
0
20
40
60
80
100
Nombre de mesure
120
140
160
180
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 %
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Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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.
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Caractérisation minéralogique, métallurgique et environnementale d’échantillons
de la zone 36 Est du gisement O’Brien - PU-2013-12-860
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)
Au pondéré sur l'aire total des arsénopyrites (ppm)
396 757
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
Page 44
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)
Caractérisation physicochimique du
minerai et des stériles à la propriété
O’Brien
Ressources Minières Radisson Inc.
Rouyn-Noranda (Cadillac), Québec
152, avenue Murdoch ~ Rouyn-Noranda (Québec) CANADA J9X 1E2
Tél. : 819 797-3222 ~ Fax : 819 762-6640 ~ www.genivar.com
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.
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
Ressources Minières Radisson Inc.
121-13415-00
i
GENIVAR
Juillet 2012
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
Ressources Minières Radisson Inc.
121-13415-00
iii
GENIVAR
Juillet 2012
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
[email protected]
Courriel :
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
[email protected]
Courriel :
M. Éric Gingras, M.Sc., EESA®
Chargé de projet
Courriel :
[email protected]
Ressources Minières Radisson Inc.
121-13415-00
1
GENIVAR
Juillet 2012
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 Vald’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.
GENIVAR
Juillet 2012
2
Ressources Minières Radisson inc.
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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.
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GENIVAR
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3
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.
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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.
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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.
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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.
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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.
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GENIVAR
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Annexe A
Tableaux de résultats
Tableau 1 (1 de 2)
Caractérisation géochimique des échantillons de minerai et de stériles
Rouyn-Noranda (Cadillac), Qc
N/réf: 121-13415-00
Échantillon
Éléments
Aluminium
Al
Antimoine
Sb
Argent
Ag
Arsenic
As
Baryum
Ba
Bore
B
Cadmium
Cd
Chrome
Cr
Cobalt
Co
Cuivre
Cu
Fer
Fe
Manganèse
Mn
Mercure
Hg
Molybdène
Mo
Nickel
Ni
Plomb
Pb
Sélénium
Se
Soufre
S
Zinc
Zn
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
%
mg/kg
Critères
LDR(1)
A
B
C
------0,6
------0,1
0,5
20
40
2
5
30
50
0,05
200 500 2000 0,01
------0,01
0,9
5
20
0,005
85
250 800
0,05
20
50
300
0,05
50
100 500
5
------0,5
1000 1000 2200 0,05
0,3
2
10
0,1
6
10
40
0,05
50
100 500
0,05
40
500 1000 0,05
3
3
10
0,05
0,04 0,1
0,2
0,1
110 500 1500 0,05
Minerai #1
136099
14424
7,9
<2
3226
119
<0,01
17,160
142
18,3
54
32990
604
<0,01
6,9
54,9
16,6
0,06
1,30
41,9
Minerai #2
136100
13916
7,5
<2
4382
99
<0,01
21,300
157
16,5
35
29893
508
<0,01
6,5
50,2
17,4
0,13
1,30
47,7
Minerai #3
136101
22307
7,7
<2
2050
139
<0,01
11,610
138
32,5
68
62208
1248
<0,01
2,9
39,8
5,8
<0,05
2,30
8,5
Minerai #4
136110
15252
1,8
<2
18144
69,4
<0,01
0,182
82
22,2
97
68876
1384
<0,01
0,92
46,2
<0,05
<0,05
2,20
65
NOTE:
(1)
(2)
: Limite de détection rapportée par le laboratoire.
: Pour une valeur inférieure à la limite de détection rapportée (LDR), la concentration utilisée correspond à |LDR|.
LÉGENDE:
100
100
100
100
100
: Concentration < A
: Concentration = A
: Concentration > A et < B
: Concentration > B et < C
: Concentration > C
Minerai #5
136111
19168
1
<2
17737
188
<0,01
9,900
116
25,1
52
69851
1376
<0,01
1,0
42,9
<0,05
<0,05
2,10
<0,05
Minerai #6
136112
18404
1,4
<2
13615
128
<0,01
0,122
81
6,6
52
47200
802
<0,01
1,9
37,2
<0,05
5,6
1,60
38
Minerai #7
136113
14402
0,5
<2
13833
100
<0,01
0,145
75
4,5
60
47263
987
<0,01
1,0
34
<0,05
<0,05
1,60
33,0
Minerai #8
136114
16620
2
2
11659
162
<0,01
0,127
136
14,2
42
55753
1038
<0,01
1,0
44,2
<0,05
<0,05
1,80
8
Moyenne
minerai(2)
16812
3,7
<2
10581
126
<0,01
7,568
116
17,5
58
51754
993
<0,01
2,8
43,7
5,0
0,8
1,78
34,6
Tableau 1 (2 de 2)
Caractérisation géochimique des échantillons de minerai et de stériles
Rouyn-Noranda (Cadillac), Qc
N/réf: 121-13415-00
Échantillon
Éléments
Aluminium
Al
Antimoine
Sb
Argent
Ag
Arsenic
As
Baryum
Ba
Bore
B
Cadmium
Cd
Chrome
Cr
Cobalt
Co
Cuivre
Cu
Fer
Fe
Manganèse
Mn
Mercure
Hg
Molybdène
Mo
Nickel
Ni
Plomb
Pb
Sélénium
Se
Soufre
S
Zinc
Zn
Critères
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
mg/kg
%
mg/kg
A
B
C
LDR(1)
------0,6
------0,1
0,5
20
40
2
5
30
50
0,05
200 500 2000 0,01
------0,01
0,9
5
20
0,005
85
250 800
0,05
20
50
300
0,05
50
100 500
5
------0,5
1000 1000 2200 0,05
0,3
2
10
0,1
6
10
40
0,05
50
100 500
0,05
40
500 1000 0,05
3
3
10
0,05
0,04 0,1
0,2
0,1
110 500 1500 0,05
Stérile #1
136102
24659
10,6
<2
28,40
177
<0,01
<0,005
247
24,9
48
37116
360
<0,1
6,6
86
12,2
<0,05
0,27
63
Stérile #2
136103
31985
23,6
<2
667,00
4
<0,01
2,170
1954
67,0
64
58712
1259
0,2
4,3
685
2,4
<0,05
0,11
28
Stérile #3
136104
25495
0,4
<2
458,0
191
<0,01
1,970
220
25,0
61
47190
663
<0,1
6,2
63
8,1
<0,05
0,61
59
Stérile #4
136105
22364
9,7
<2
199,00
301
<0,01
0,341
387
22,4
33
37175
643
<0,1
2,4
112
6,3
<0,05
0,30
59
NOTE:
(1)
(2)
: Limite de détection rapportée par le laboratoire.
: Pour une valeur inférieure à la limite de détection rapportée (LDR), la concentration utilisée correspond à |LDR|.
LÉGENDE:
100
100
100
100
100
: Concentration < A
: Concentration = A
: Concentration > A et < B
: Concentration > B et < C
: Concentration > C
Stérile #5
136106
40183
10,1
<2
16,20
9
<0,01
<0,005
173
45,7
91
74008
1124
<0,1
7,6
56
7,6
<0,05
0,33
106
Stérile #6
136107
21085
0,2
<2
<0,05
35
<0,01
8,07
125
9,3
23
47093
473
<0,1
2,6
89
<0,05
<0,05
0,17
58
Stérile #7
136108
16831
<0,1
<2
15,2
<0,01
<0,01
0,268
81,2
33
146
60180
1373
<0,1
1
53
<0,05
<0,05
0,61
10
Stérile #8
136109
16955
2
<2
29,5
122
<0,01
0,158
170
5
19
37423
385
<0,1
1,5
84
<0,05
<0,05
0,19
71
Moyenne
stérile(2)
28937
7,1
<2
176,67
105
<0,01
2,163
510,4
29,0
59
50840
810
0,1
4,0
153
4,6
<0,05
0,32
57
Tableau 2 (1 de 2)
Résultats du potentiel générateur d'acide pour les échantillons de minerai et de stériles
Rouyn-Noranda (Cadillac), Qc
N/réf: 121-13415-00
Échantillon
PGA
PN
kg CaCO3/ t
PA
kg CaCO3/ t
PNN
kg CaCO3/ t
PN/PA
ratio
S
%
Potentiel
---
Minerai #1
136099
63
40,0
23,0
1,6
1,3
NPGA
Minerai #2
136100
89
39,4
49,6
2,3
1,3
NPGA
Minerai #3
136101
236
71,9
164,0
3,3
2,3
NPGA
Minerai #4
136110
157
69,7
87,3
2,3
2,2
NPGA
Minerai #5
136111
207
66,3
141,0
3,1
2,1
NPGA
Minerai #6
136112
83
49,3
33,7
1,7
1,6
NPGA
Minerai #7
136113
126,0
51,1
74,9
2,5
1,6
NPGA
Minerai #8
136114
177
57,5
120
3,1
1,8
NPGA
Moyenne
minerai(1)
142,25
55,65
86,69
2,46
1,78
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'acide
PA: Potentiel d'acidité des sulfures
PNN: Potentiel de neutralisation net, calculé comme PN-PA
0,1
15
5
NPGA
ZG
PGA
: PN/PA < 3.0
: PNN < 20 kg CaCO3/ton
: sulfure > 0,3 %
: non potentiel générateur d'acide
: Zone grise
: 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.
Tableau 2 (2 de 2)
Résultats du potentiel générateur d'acide pour les échantillons de minerai et de stériles
Rouyn-Noranda (Cadillac), Qc
N/réf: 121-13415-00
Échantillon
PGA
PN
kg CaCO3/ t
PA
kg CaCO3/ t
PNN
kg CaCO3/ t
PN/PA
ratio
S
%
Potentiel
---
Stérile #1
136102
46
8,40
37,6
5,5
0,27
NPGA
Stérile #2
136103
259
3,40
256,0
76,2
0,11
NPGA
Stérile #3
136104
29
19,10
9,9
1,5
0,61
PGA
Stérile #4
136105
158
9,40
149,0
16,8
0,30
NPGA
Stérile #5
136106
149
10,30
139,0
14,5
0,33
NPGA
Stérile #6
136107
34
5,30
28,7
6,4
0,17
NPGA
Stérile #7
136108
314
19,0
285,0
16,5
0,61
NPGA
Stérile #8
136109
24
6,0
18,0
4,0
0,19
NPGA
Moyenne
stérile(2)
126,63
10,11
115,40
17,67
0,32
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'acide
PA: Potentiel d'acidité des sulfures
PNN: Potentiel de neutralisation net, calculé comme PN-PA
0,1
15
5
NPGA
ZG
PGA
: PN/PA < 3.0
: PNN < 20 kg CaCO3/ton
: sulfure > 0,3 %
: non potentiel générateur d'acide
: Zone grise
: 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.
Tableau 3 (1 de 2)
Résultats des analyses chimiques (essai de lixiviation TCLP 1311) des échantillons de minerai et de stériles
Rouyn-Noranda (Cadillac), Qc
N/réf: 121-13415-00
Échantillons
Paramètres
pH
Aluminium
Antimoine
Argent
Arsenic
Baryum
Bore
Cadmium
Chrome
Cobalt
Cuivre
Fer
Manganèse
Mercure
Molybdène
Nickel
Plomb
Sélénium
Zinc
Al
Sb
Ag
As
Ba
B
Cd
Cr
Co
Cu
Fe
Mn
Hg
Mo
Ni
Pb
Se
Zn
EC(1)
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
6,5> <8,5
--0,006
0,1
0,025
1
--0,005
0,05
--1
--0,05
0,001
0,07
0,02
0,01
0,01
5
RÉSIE(2)
--0,75
-0,00062
0,34
5,3
--0,0021
0,0016
0,5
0,0073
----0,00013
2
0,26
0,034
0,02
0,067
Directive
019(3)
--------5
100
--0,5
5
--------0,1
----5
1
---
LDR
(4)
--0,006
0,0001
0,0005
0,0005
0,0005
0,01
0,0001
0,0006
0,001
0,0005
0,01
0,0005
0,00002
0,0005
0,0005
0,0005
0,001
0,001
Minerai #1
136099
5,53
0,43
0,0030
<0,0005
0,5905
0,130
<0,01
0,0003
0,0225
0,004
0,0023
6,4
8,3870
<0,00002
0,0029
0,0638
0,0353
0,001
0,039
Minerai #2
136100
6,35
0,03
0,0076
<0,0005
0,1070
0,129
<0,01
0,0001
0,0066
0,003
0,0060
2,0
7,7020
<0,00002
0,0025
0,0776
<0,0005
<0,001
0,030
Minerai #3
136101
5,67
<0,006
<0,0001
<0,0005
0,1591
0,335
<0,01
<0,0001
<0,0006
<0,001
0,0196
19,3
18,8500
<0,00002
<0,0005
0,0285
0,0005
<0,001
<0,001
Minerai #4
136110
0,219
0,0072
<0,0005
1,504
0,190
<0,01
0,0007
0,0242
0,025
0,0019
13,3
17,7000
<0,00002
<0,0005
0,4904
0,0152
<0,001
0,034
Minerai #5
136111
0,038
0,0074
<0,0005
0,9565
0,386
<0,01
0,0008
0,0296
0,022
0,0012
12,7
19,98
<0,00002
<0,0005
0,4821
0,0073
<0,001
0,015
Minerai #6
136112
<0,006
0,0040
<0,0005
0,3325
0,232
<0,01
0,0005
0,0289
0,009
0,0226
1,1
8,8910
<0,00002
0,0021
0,1917
0,0035
<0,001
0,032
NOTES:
: 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|.
(1)
LÉGENDE:
100
100
: Non défini ou non analysé
: Concentration supérieure au critère RÉSIE
: Concentration supérieure au critère de la directive 019
Minerai #7
136113
1,66
0,0063
<0,0005
3,9560
0,277
<0,01
0,0013
0,0646
0,027
0,0017
23,1
21,3500
<0,00002
0,0009
0,4249
0,0281
<0,001
0,037
Minerai #8
136114
0,189
0,0081
<0,0005
1,0860
0,348
<0,01
0,0005
0,0372
0,024
0,0012
13,5
17,1400
<0,00002
<0,0005
0,4978
0,0110
<0,001
0,009
(5)
Moyenne
5,850
0,322
0,005
<0,0005
1,086
0,253
<0,01
0,0006
0,0268
0,014
0,0071
11,425
15,0000
<0,00002
0,0021
0,2821
0,0144
<0,001
0,026
Tableau 3 (2 de 2)
Résultats des analyses chimiques (essai de lixiviation TCLP 1311) des échantillons de minerai et de stériles
Rouyn-Noranda (Cadillac), Qc
N/réf: 121-13415-00
Échantillons
Paramètres
pH
Aluminium
Antimoine
Argent
Arsenic
Baryum
Bore
Cadmium
Chrome
Cobalt
Cuivre
Fer
Manganèse
Mercure
Molybdène
Nickel
Plomb
Sélénium
Zinc
Al
Sb
Ag
As
Ba
B
Cd
Cr
Co
Cu
Fe
Mn
Hg
Mo
Ni
Pb
Se
Zn
EC(1)
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
6,5> <8,5
--0,006
0,1
0,025
1
--0,005
0,05
--1
--0,05
0,001
0,07
0,02
0,01
0,01
5
RÉSIE(2)
--0,75
-0,00062
0,34
5,3
--0,0021
0,0016
0,5
0,0073
----0,00013
2
0,26
0,034
0,02
0,067
Directive
019(3)
--------5
100
--0,5
5
--------0,1
----5
1
---
LDR
(4)
--0,006
0,0001
0,0005
0,0005
0,0005
0,01
0,0001
0,0006
0,001
0,0005
0,01
0,0005
0,00002
0,0005
0,0005
0,001
0,001
0,001
Stérile #1
136102
5,47
<0,006
<0,0001
<0,0005
<0,0005
0,207
<0,01
<0,0001
<0,0006
<0,001
0,0196
2,4
1,9580
<0,00002
<0,0005
0,0148
<0,0005
<0,001
<0,001
Stérile #2
136103
5,02
<0,006
<0,0001
<0,0005
0,0501
0,037
<0,01
<0,0001
<0,0006
0,009
0,0070
40,9
12,7200
<0,00002
<0,0005
0,1271
<0,0005
<0,001
<0,001
Stérile #3
136104
5,26
<0,006
<0,0001
<0,0005
0,2080
0,277
<0,01
<0,0001
<0,0006
<0,001
0,0197
4,3
3,1400
<0,00002
<0,0005
0,0163
<0,0005
<0,001
<0,001
Stérile #4
136105
6,2
<0,006
<0,0001
<0,0005
<0,0005
0,424
<0,01
<0,0001
<0,0006
<0,001
0,0202
4,4
4,8650
<0,00002
<0,0005
0,0029
<0,0005
<0,001
<0,001
Stérile #5
136106
5,73
0,34
0,0011
<0,0005
<0,0005
0,019
<0,01
0,0010
0,0238
0,008
0,0037
13,8
16,0100
<0,00002
<0,0005
0,0660
<0,0005
<0,001
0,017
Stérile #6
136107
0,703
0,0025
<0,0005
<0,0005
0,068
<0,01
<0,0001
0,036
0,006
0,0239
7,3
0,9857
<0,00002
0,0021
0,1142
0,0118
<0,001
<0,001
NOTES:
: 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|.
(1)
LÉGENDE:
100
100
: Non défini ou non analysé
: Concentration supérieure au critère RÉSIE
: Concentration supérieure au critère de la directive 019
Stérile #7
136108
0,03
0,0003
<0,0005
<0,0005
0,0275
<0,01
0,0012
0,0293
0,026
0,0014
11,1
12,30
<0,00002
0,0007
0,4630
0,0013
<0,001
0,040
Stérile #8
136109
1,95
0,0014
<0,0005
<0,0005
0,148
<0,01
<0,0001
0,0421
0,004
0,0254
7,5
0,7625
<0,00002
0,0022
0,0573
0,0013
<0,001
0,020
Moyenne
(2)
stérile
5,54
0,381
0,0057
<0,0005
0,033
0,151
<0,01
0,0004
0,0167
0,007
0,015
11,463
6,5927
<0,00002
0,0017
0,1077
0,0048
<0,001
0,010
Annexe B
Certificats d’analyse du laboratoire
Certificat d'analyse
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.
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 1 de 7
Certificat d'analyse
Numéro de projet : C-113484
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136099
Lieu de prélèvement : Radisson
Paramètres
Résultats
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Préparation d'échantillon
Sélénium (Se)
Zinc (Zn)
Potentiel générateur acide
Soufre
14424
7.9
3226(> C)
119(< A)
<0.01
17.16(B-C)
142(A-B)
18.3(< A)
54(A-B)
32990
604(< A)
6.9(A-B)
54.9(A-B)
16.6(< A)
Potentiel d'acidité maximal (PA)
Potentiel neutralisation brut (PN)
Potentiel neutralisaton net (PNN)
% Humidité
Mercure (Hg)
Argent (Ag)
Méthode d'analyse
Date d'analyse
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
30 mai 2012
30 mai 2012
05 juin 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
Sous-traitance\Laboratoire Expert Inc.
0.06(< A) mg/Kg
41.9(< A) mg/Kg
M-MET-3.0
M-MET-3.0
05 juin 2012
30 mai 2012
Sous-traitance\Maxxam Analytics Inc
1.3
40.0
63
23.0
<0.1
<0.1
<2
%
kg CaCO3/t
kg CaCO3/t
kg CaCO3/t
%
mg/Kg
mg/Kg
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
M-HUM-1.0
M-HG-2.0
M-MET-4.0
12 juin 2012
08 juin 2012
08 juin 2012
08 juin 2012
29 mai 2012
05 juin 2012
30 mai 2012
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 2 de 7
Limite de détection rapportée
Numéro de projet : C-113484
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136099
Lieu de prélèvement : Radisson
Paramètre
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
% Humidité
Mercure (Hg)
Argent (Ag)
Valeur
0.6
0.1
0.05
0.01
0.01
0.005
0.05
0.05
5
0.5
0.05
0.05
0.05
0.05
0.05
0.05
0.1
0.1
2
Unité
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
%
mg/Kg
mg/Kg
Méthode
Accréditation
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HUM-1.0
M-HG-2.0
M-MET-4.0
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 3 de 7
Certificat contrôle qualité
Numéro de projet : C-113484
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136099
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Argent (Ag) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Cobalt (Co) mg/Kg
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
<0.6
DMR.0231-2012-1
36907
97.1%
32300 - 43700
<0.1
<2
D-076-540
33.0
95.9%
29.2 - 39.6
<0.05
D-076-540
75.6
80%
73.2 - 115.8
<0.01
D-076-540
189
86.8%
142 - 192
<0.01
D-076-540
98
92.5%
82 - 130
<0.005
D-076-540
63.1
95.7%
46.9 - 74.1
<0.05
D-076-540
63.8
90.6%
59.8 - 81.0
<0.05
D-076-540
111
91.2%
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 4 de 7
Certificat contrôle qualité
Numéro de projet : C-113484
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136099
Lieu de prélèvement : Radisson
Paramètres
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Mercure (Hg) mg/Kg
Mercure (Hg) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Plomb (Pb) mg/Kg
Sélénium (Se) mg/Kg
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
87 - 117
<5
D-076-540
85.0
93.2%
67.7 - 91.5
<0.5
<0.05
D-076-540
328
84.1%
219 - 347
<0.1
DOLT-4
2.20
85.3%
2.12 - 3.04
DORM-3
0.40
97.6%
0.34 - 0.48
<0.05
D-076-540
50.0
80.1%
32.3 - 51.1
<0.05
D-076-540
68.7
80.7%
44.6 - 70.6
<0.05
D-076-540
100.0
91.1%
71.1 - 112.5
<0.05
D-076-540
77.4
89.6%
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 5 de 7
Certificat contrôle qualité
Numéro de projet : C-113484
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136099
Lieu de prélèvement : Radisson
Paramètres
Zinc (Zn) mg/Kg
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
67.0 - 105.8
<0.05
D-076-540
152
91.4%
119 - 161
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 6 de 7
Informations supplémentaires
Numéro de projet : C-113484
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136099
Lieu de prélèvement : Radisson
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
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 7 de 7
Certificat d'analyse
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-113486
Lieu de prélèvement : Radisson
Date de prélèvement : 17 avril 2012
Échantillon : 136100
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.
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 1 de 7
Certificat d'analyse
Numéro de projet : C-113486
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136100
Lieu de prélèvement : Radisson
Paramètres
Résultats
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Mercure (Hg)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Préparation d'échantillon
Sélénium (Se)
Zinc (Zn)
Potentiel générateur acide
Soufre
13916
7.5
4382(> C)
99.4(< A)
<0.01
21.30(> C)
157(A-B)
16.5(< A)
35(< A)
29893
508(< A)
<0.1
6.5(A-B)
50.2(A-B)
17.4(< A)
Potentiel d'acidité maximal (PA)
Potentiel neutralisation brut (PN)
Potentiel neutralisaton net (PNN)
% Humidité
Argent (Ag)
Méthode d'analyse
Date d'analyse
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HG-2.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
30 mai 2012
30 mai 2012
05 juin 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
05 juin 2012
30 mai 2012
30 mai 2012
30 mai 2012
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
Sous-traitance\Laboratoire Expert Inc.
0.13(< A) mg/Kg
47.7(< A) mg/Kg
M-MET-3.0
M-MET-3.0
05 juin 2012
30 mai 2012
Sous-traitance\Maxxam Analytics Inc
1.3
39.4
89
49.6
<0.1
<2
%
kg CaCO3/t
kg CaCO3/t
kg CaCO3/t
%
mg/Kg
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
M-HUM-1.0
M-MET-4.0
12 juin 2012
08 juin 2012
08 juin 2012
08 juin 2012
29 mai 2012
30 mai 2012
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 2 de 7
Limite de détection rapportée
Numéro de projet : C-113486
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136100
Lieu de prélèvement : Radisson
Paramètre
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Mercure (Hg)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
% Humidité
Argent (Ag)
Valeur
0.6
0.1
0.05
0.01
0.01
0.005
0.05
0.05
5
0.5
0.05
0.1
0.05
0.05
0.05
0.05
0.05
0.1
2
Unité
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
%
mg/Kg
Méthode
Accréditation
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HG-2.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HUM-1.0
M-MET-4.0
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 3 de 7
Certificat contrôle qualité
Numéro de projet : C-113486
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136100
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Argent (Ag) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Cobalt (Co) mg/Kg
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
<0.6
DMR.0231-2012-1
36907
97.1%
32300 - 43700
<0.1
<2
D-076-540
33.0
95.9%
29.2 - 39.6
<0.05
D-076-540
75.6
80%
73.2 - 115.8
<0.01
D-076-540
189
86.8%
142 - 192
<0.01
D-076-540
98
92.5%
82 - 130
<0.005
D-076-540
63.1
95.7%
46.9 - 74.1
<0.05
D-076-540
63.8
90.6%
59.8 - 81.0
<0.05
D-076-540
111
91.2%
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 4 de 7
Certificat contrôle qualité
Numéro de projet : C-113486
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136100
Lieu de prélèvement : Radisson
Paramètres
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Mercure (Hg) mg/Kg
Mercure (Hg) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Plomb (Pb) mg/Kg
Sélénium (Se) mg/Kg
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
87 - 117
<5
D-076-540
85.0
93.2%
67.7 - 91.5
<0.5
<0.05
D-076-540
328
84.1%
219 - 347
<0.1
DOLT-4
2.20
85.3%
2.12 - 3.04
DORM-3
0.40
97.6%
0.34 - 0.48
<0.05
D-076-540
50.0
80.1%
32.3 - 51.1
<0.05
D-076-540
68.7
80.7%
44.6 - 70.6
<0.05
D-076-540
100.0
91.1%
71.1 - 112.5
<0.05
D-076-540
77.4
89.6%
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 5 de 7
Certificat contrôle qualité
Numéro de projet : C-113486
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136100
Lieu de prélèvement : Radisson
Paramètres
Zinc (Zn) mg/Kg
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
67.0 - 105.8
<0.05
D-076-540
152
91.4%
119 - 161
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 6 de 7
Informations supplémentaires
Numéro de projet : C-113486
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136100
Lieu de prélèvement : Radisson
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
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 7 de 7
Certificat d'analyse
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-113488
Lieu de prélèvement : Radisson
Date de prélèvement : 17 avril 2012
Échantillon : 136101
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.
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 1 de 7
Certificat d'analyse
Numéro de projet : C-113488
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136101
Lieu de prélèvement : Radisson
Paramètres
Résultats
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Préparation d'échantillon
Sélénium (Se)
Zinc (Zn)
Potentiel générateur acide
Soufre
22307
7.7
2050(> C)
139(< A)
<0.01
11.61(B-C)
138(A-B)
32.5(A-B)
68(A-B)
62208
1248(B-C)
2.9(< A)
39.8(< A)
5.8(< A)
Potentiel d'acidité maximal (PA)
Potentiel neutralisation brut (PN)
Potentiel neutralisaton net (PNN)
% Humidité
Argent (Ag)
Mercure (Hg)
Méthode d'analyse
Date d'analyse
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
30 mai 2012
30 mai 2012
05 juin 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
Sous-traitance\Laboratoire Expert Inc.
<0.05(< A) mg/Kg
8.5(< A) mg/Kg
M-MET-3.0
M-MET-3.0
05 juin 2012
30 mai 2012
Sous-traitance\Maxxam Analytics Inc
2.3
71.9
236
164
<0.1
<2
<0.1
%
kg CaCO3/t
kg CaCO3/t
kg CaCO3/t
%
mg/Kg
mg/Kg
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
M-HUM-1.0
M-MET-4.0
M-HG-2.0
12 juin 2012
08 juin 2012
08 juin 2012
08 juin 2012
29 mai 2012
30 mai 2012
05 juin 2012
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 2 de 7
Limite de détection rapportée
Numéro de projet : C-113488
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136101
Lieu de prélèvement : Radisson
Paramètre
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
% Humidité
Argent (Ag)
Mercure (Hg)
Valeur
0.6
0.1
0.05
0.01
0.01
0.005
0.05
0.05
5
0.5
0.05
0.05
0.05
0.05
0.05
0.05
0.1
2
0.1
Unité
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
%
mg/Kg
mg/Kg
Méthode
Accréditation
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HUM-1.0
M-MET-4.0
M-HG-2.0
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 3 de 7
Certificat contrôle qualité
Numéro de projet : C-113488
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136101
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Argent (Ag) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
<0.6
DMR.0231-2012-1
36907
97.1%
32300 - 43700
22307-24289
<0.1
7.7-12.9
<2
D-076-540
33.0
95.9%
29.2 - 39.6
<2-<2
<0.05
D-076-540
75.6
80%
73.2 - 115.8
2050-2101
<0.01
D-076-540
189
86.8%
142 - 192
139-149
<0.01
D-076-540
98
92.5%
82 - 130
<0.01-<0.01
<0.005
D-076-540
63.1
95.7%
46.9 - 74.1
11.61-11.02
<0.05
D-076-540
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 4 de 7
Certificat contrôle qualité
Numéro de projet : C-113488
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136101
Lieu de prélèvement : Radisson
Paramètres
Cobalt (Co) mg/Kg
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Mercure (Hg) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
63.8
90.6%
59.8 - 81.0
138-147
<0.05
D-076-540
111
91.2%
87 - 117
32.5-33.2
<5
D-076-540
85.0
93.2%
67.7 - 91.5
68-74
<0.5
62208-67454
<0.05
D-076-540
328
84.1%
219 - 347
1248-1309
<0.1
DORM-3
0.40
97.6%
0.34 - 0.48
<0.05
D-076-540
50.0
80.1%
32.3 - 51.1
2.9-2.4
<0.05
D-076-540
68.7
80.7%
44.6 - 70.6
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 5 de 7
Certificat contrôle qualité
Numéro de projet : C-113488
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136101
Lieu de prélèvement : Radisson
Paramètres
Plomb (Pb) mg/Kg
Sélénium (Se) mg/Kg
Zinc (Zn) mg/Kg
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Duplicata
39.8-40.7
<0.05
D-076-540
100.0
91.1%
71.1 - 112.5
5.8-4.9
<0.05
D-076-540
77.4
89.6%
67.0 - 105.8
<0.05-<0.05
<0.05
D-076-540
152
91.4%
119 - 161
8.5-9.2
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 6 de 7
Informations supplémentaires
Numéro de projet : C-113488
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136101
Lieu de prélèvement : Radisson
Méthode laboratoire
Méthode de référence
M-MET-3.0
MA.200-Mét. 1.2
M-MET-4.0
EPA-3050b
M-HG-2.0
MA.207-Hg 2.0
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 7 de 7
Certificat d'analyse
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-113490
Lieu de prélèvement : Radisson
Date de prélèvement : 17 avril 2012
Échantillon : 136102
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.
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 1 de 7
Certificat d'analyse
Numéro de projet : C-113490
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136102
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Préparation d'échantillon
Sélénium (Se)
Zinc (Zn)
Potentiel générateur acide
Soufre
Potentiel d'acidité maximal (PA)
Potentiel neutralisation brut (PN)
Potentiel neutralisaton net (PNN)
% Humidité
Argent (Ag)
Mercure (Hg)
Résultats
24659
10.6
28.4(A-B)
177(< A)
<0.01
<0.005(< A)
247(A-B)
24.9(A-B)
48(< A)
37116
360(< A)
6.6(A-B)
86.4(A-B)
12.2(< A)
Méthode d'analyse
Date d'analyse
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
30 mai 2012
30 mai 2012
05 juin 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
Sous-traitance\Laboratoire Expert Inc.
<0.05(< A) mg/Kg
62.5(< A) mg/Kg
M-MET-3.0
M-MET-3.0
05 juin 2012
30 mai 2012
Sous-traitance\Maxxam Analytics Inc
0.27
8.4
46
37.6
<0.1
<2
<0.1
%
kg CaCO3/t
kg CaCO3/t
kg CaCO3/t
%
mg/Kg
mg/Kg
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
M-HUM-1.0
M-MET-4.0
M-HG-2.0
12 juin 2012
08 juin 2012
08 juin 2012
08 juin 2012
29 mai 2012
30 mai 2012
05 juin 2012
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 2 de 7
Limite de détection rapportée
Numéro de projet : C-113490
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136102
Lieu de prélèvement : Radisson
Paramètre
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
% Humidité
Argent (Ag)
Mercure (Hg)
Valeur
0.6
0.1
0.05
0.01
0.01
0.005
0.05
0.05
5
0.5
0.05
0.05
0.05
0.05
0.05
0.05
0.1
2
0.1
Unité
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
%
mg/Kg
mg/Kg
Méthode
Accréditation
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HUM-1.0
M-MET-4.0
M-HG-2.0
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 3 de 7
Certificat contrôle qualité
Numéro de projet : C-113490
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136102
Lieu de prélèvement : Radisson
Paramètres
% Humidité %
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Argent (Ag) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Cobalt (Co) mg/Kg
Duplicata
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
<0.1-<0.1
<0.6
DMR.0231-2012-1
36907
97.1%
32300 - 43700
<0.1
<2
D-076-540
33.0
95.9%
29.2 - 39.6
<0.05
D-076-540
75.6
80%
73.2 - 115.8
<0.01
D-076-540
189
86.8%
142 - 192
<0.01
D-076-540
98
92.5%
82 - 130
<0.005
D-076-540
63.1
95.7%
46.9 - 74.1
<0.05
D-076-540
63.8
90.6%
59.8 - 81.0
<0.05
D-076-540
111
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 4 de 7
Certificat contrôle qualité
Numéro de projet : C-113490
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136102
Lieu de prélèvement : Radisson
Paramètres
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Mercure (Hg) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Plomb (Pb) mg/Kg
Sélénium (Se) mg/Kg
Zinc (Zn) mg/Kg
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
91.2%
87 - 117
<5
D-076-540
85.0
93.2%
67.7 - 91.5
<0.5
<0.05
D-076-540
328
84.1%
219 - 347
<0.1
DORM-3
0.40
97.6%
0.34 - 0.48
<0.05
D-076-540
50.0
80.1%
32.3 - 51.1
<0.05
D-076-540
68.7
80.7%
44.6 - 70.6
<0.05
D-076-540
100.0
91.1%
71.1 - 112.5
<0.05
D-076-540
77.4
89.6%
67.0 - 105.8
<0.05
D-076-540
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 5 de 7
Certificat contrôle qualité
Numéro de projet : C-113490
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136102
Lieu de prélèvement : Radisson
Paramètres
Valeur obtenue 152
Justesse 91.4%
Intervalle 119 - 161
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 6 de 7
Informations supplémentaires
Numéro de projet : C-113490
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136102
Lieu de prélèvement : Radisson
Méthode laboratoire
Méthode de référence
M-MET-3.0
MA.200-Mét. 1.2
M-MET-4.0
EPA-3050b
M-HG-2.0
MA.207-Hg 2.0
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 7 de 7
Certificat d'analyse
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-113492
Lieu de prélèvement : Radisson
Date de prélèvement : 17 avril 2012
Échantillon : 136103
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.
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 1 de 7
Certificat d'analyse
Numéro de projet : C-113492
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136103
Lieu de prélèvement : Radisson
Paramètres
Résultats
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Préparation d'échantillon
Sélénium (Se)
Zinc (Zn)
Potentiel générateur acide
Soufre
31985
23.6
667(> C)
4.2(< A)
<0.01
2.17(A-B)
1954(> C)
67.0(B-C)
64(A-B)
58712
1259(B-C)
4.3(< A)
685(> C)
2.4(< A)
Potentiel d'acidité maximal (PA)
Potentiel neutralisation brut (PN)
Potentiel neutralisaton net (PNN)
% Humidité
Argent (Ag)
Mercure (Hg)
Méthode d'analyse
Date d'analyse
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
30 mai 2012
30 mai 2012
05 juin 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
Sous-traitance\Laboratoire Expert Inc.
<0.05(< A) mg/Kg
28.4(< A) mg/Kg
M-MET-3.0
M-MET-3.0
05 juin 2012
30 mai 2012
Sous-traitance\Maxxam Analytics Inc
0.11
3.4
259
256
<0.1
<2
0.2
%
kg CaCO3/t
kg CaCO3/t
kg CaCO3/t
%
mg/Kg
mg/Kg
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
M-HUM-1.0
M-MET-4.0
M-HG-2.0
12 juin 2012
08 juin 2012
08 juin 2012
08 juin 2012
29 mai 2012
30 mai 2012
05 juin 2012
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 2 de 7
Limite de détection rapportée
Numéro de projet : C-113492
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136103
Lieu de prélèvement : Radisson
Paramètre
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
% Humidité
Argent (Ag)
Mercure (Hg)
Valeur
0.6
0.1
0.05
0.01
0.01
0.005
0.05
0.05
5
0.5
0.05
0.05
0.05
0.05
0.05
0.05
0.1
2
0.1
Unité
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
%
mg/Kg
mg/Kg
Méthode
Accréditation
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HUM-1.0
M-MET-4.0
M-HG-2.0
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 3 de 7
Certificat contrôle qualité
Numéro de projet : C-113492
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136103
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Argent (Ag) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Cobalt (Co) mg/Kg
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
<0.6
DMR.0231-2012-1
36907
97.1%
32300 - 43700
<0.1
<2
D-076-540
33.0
95.9%
29.2 - 39.6
<0.05
D-076-540
75.6
80%
73.2 - 115.8
<0.01
D-076-540
189
86.8%
142 - 192
<0.01
D-076-540
98
92.5%
82 - 130
<0.005
D-076-540
63.1
95.7%
46.9 - 74.1
<0.05
D-076-540
63.8
90.6%
59.8 - 81.0
<0.05
D-076-540
111
91.2%
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 4 de 7
Certificat contrôle qualité
Numéro de projet : C-113492
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136103
Lieu de prélèvement : Radisson
Paramètres
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Mercure (Hg) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Plomb (Pb) mg/Kg
Sélénium (Se) mg/Kg
Zinc (Zn) mg/Kg
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
87 - 117
<5
D-076-540
85.0
93.2%
67.7 - 91.5
<0.5
<0.05
D-076-540
328
84.1%
219 - 347
<0.1
DORM-3
0.40
97.6%
0.34 - 0.48
<0.05
D-076-540
50.0
80.1%
32.3 - 51.1
<0.05
D-076-540
68.7
80.7%
44.6 - 70.6
<0.05
D-076-540
100.0
91.1%
71.1 - 112.5
<0.05
D-076-540
77.4
89.6%
67.0 - 105.8
<0.05
D-076-540
152
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 5 de 7
Certificat contrôle qualité
Numéro de projet : C-113492
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136103
Lieu de prélèvement : Radisson
Paramètres
Justesse 91.4%
Intervalle 119 - 161
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 6 de 7
Informations supplémentaires
Numéro de projet : C-113492
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136103
Lieu de prélèvement : Radisson
Méthode laboratoire
Méthode de référence
M-MET-3.0
MA.200-Mét. 1.2
M-MET-4.0
EPA-3050b
M-HG-2.0
MA.207-Hg 2.0
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 7 de 7
Certificat d'analyse
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-113494
Lieu de prélèvement : Radisson
Date de prélèvement : 17 avril 2012
Échantillon : 136104
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.
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 1 de 7
Certificat d'analyse
Numéro de projet : C-113494
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136104
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Préparation d'échantillon
Sélénium (Se)
Zinc (Zn)
Potentiel générateur acide
Soufre
Potentiel d'acidité maximal (PA)
Potentiel neutralisation brut (PN)
Potentiel neutralisaton net (PNN)
% Humidité
Argent (Ag)
Mercure (Hg)
Résultats
25495
0.4
458(> C)
191(< A)
<0.01
1.97(A-B)
220(A-B)
25.0(A-B)
61(A-B)
47190
663(< A)
6.2(A-B)
62.5(A-B)
8.1(< A)
Méthode d'analyse
Date d'analyse
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
30 mai 2012
30 mai 2012
05 juin 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
Sous-traitance\Laboratoire Expert Inc.
<0.05(< A) mg/Kg
58.6(< A) mg/Kg
M-MET-3.0
M-MET-3.0
05 juin 2012
30 mai 2012
Sous-traitance\Maxxam Analytics Inc
0.61
19.1
29
9.90
<0.1
<2
<0.1
%
kg CaCO3/t
kg CaCO3/t
kg CaCO3/t
%
mg/Kg
mg/Kg
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
M-HUM-1.0
M-MET-4.0
M-HG-2.0
12 juin 2012
08 juin 2012
08 juin 2012
08 juin 2012
29 mai 2012
30 mai 2012
05 juin 2012
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 2 de 7
Limite de détection rapportée
Numéro de projet : C-113494
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136104
Lieu de prélèvement : Radisson
Paramètre
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
% Humidité
Argent (Ag)
Mercure (Hg)
Valeur
0.6
0.1
0.05
0.01
0.01
0.005
0.05
0.05
5
0.5
0.05
0.05
0.05
0.05
0.05
0.05
0.1
2
0.1
Unité
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
%
mg/Kg
mg/Kg
Méthode
Accréditation
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HUM-1.0
M-MET-4.0
M-HG-2.0
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 3 de 7
Certificat contrôle qualité
Numéro de projet : C-113494
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136104
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Argent (Ag) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Cobalt (Co) mg/Kg
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
<0.6
DMR.0231-2012-1
36907
97.1%
32300 - 43700
<0.1
<2
D-076-540
33.0
95.9%
29.2 - 39.6
<0.05
D-076-540
75.6
80%
73.2 - 115.8
<0.01
D-076-540
189
86.8%
142 - 192
<0.01
D-076-540
98
92.5%
82 - 130
<0.005
D-076-540
63.1
95.7%
46.9 - 74.1
<0.05
D-076-540
63.8
90.6%
59.8 - 81.0
<0.05
D-076-540
111
91.2%
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 4 de 7
Certificat contrôle qualité
Numéro de projet : C-113494
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136104
Lieu de prélèvement : Radisson
Paramètres
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Mercure (Hg) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Plomb (Pb) mg/Kg
Sélénium (Se) mg/Kg
Zinc (Zn) mg/Kg
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
87 - 117
<5
D-076-540
85.0
93.2%
67.7 - 91.5
<0.5
<0.05
D-076-540
328
84.1%
219 - 347
<0.1
DORM-3
0.40
97.6%
0.34 - 0.48
<0.05
D-076-540
50.0
80.1%
32.3 - 51.1
<0.05
D-076-540
68.7
80.7%
44.6 - 70.6
<0.05
D-076-540
100.0
91.1%
71.1 - 112.5
<0.05
D-076-540
77.4
89.6%
67.0 - 105.8
<0.05
D-076-540
152
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 5 de 7
Certificat contrôle qualité
Numéro de projet : C-113494
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136104
Lieu de prélèvement : Radisson
Paramètres
Justesse 91.4%
Intervalle 119 - 161
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 6 de 7
Informations supplémentaires
Numéro de projet : C-113494
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136104
Lieu de prélèvement : Radisson
Méthode laboratoire
Méthode de référence
M-MET-3.0
MA.200-Mét. 1.2
M-MET-4.0
EPA-3050b
M-HG-2.0
MA.207-Hg 2.0
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 7 de 7
Certificat d'analyse
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-113496
Lieu de prélèvement : Radisson
Date de prélèvement : 17 avril 2012
Échantillon : 136105
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.
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 1 de 7
Certificat d'analyse
Numéro de projet : C-113496
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136105
Lieu de prélèvement : Radisson
Paramètres
Résultats
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Préparation d'échantillon
Sélénium (Se)
Zinc (Zn)
Potentiel générateur acide
Soufre
22364
9.7
199(> C)
301(A-B)
<0.01
0.341(< A)
387(B-C)
22.4(A-B)
33(< A)
37175
643(< A)
2.4(< A)
112(B-C)
6.3(< A)
Potentiel d'acidité maximal (PA)
Potentiel neutralisation brut (PN)
Potentiel neutralisaton net (PNN)
% Humidité
Argent (Ag)
Mercure (Hg)
Méthode d'analyse
Date d'analyse
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
30 mai 2012
30 mai 2012
05 juin 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
Sous-traitance\Laboratoire Expert Inc.
<0.05(< A) mg/Kg
58.9(< A) mg/Kg
M-MET-3.0
M-MET-3.0
05 juin 2012
30 mai 2012
Sous-traitance\Maxxam Analytics Inc
0.30
9.4
158
149
<0.1
<2
<0.1
%
kg CaCO3/t
kg CaCO3/t
kg CaCO3/t
%
mg/Kg
mg/Kg
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
M-HUM-1.0
M-MET-4.0
M-HG-2.0
12 juin 2012
08 juin 2012
08 juin 2012
08 juin 2012
29 mai 2012
30 mai 2012
05 juin 2012
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 2 de 7
Limite de détection rapportée
Numéro de projet : C-113496
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136105
Lieu de prélèvement : Radisson
Paramètre
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
% Humidité
Argent (Ag)
Mercure (Hg)
Valeur
0.6
0.1
0.05
0.01
0.01
0.005
0.05
0.05
5
0.5
0.05
0.05
0.05
0.05
0.05
0.05
0.1
2
0.1
Unité
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
%
mg/Kg
mg/Kg
Méthode
Accréditation
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HUM-1.0
M-MET-4.0
M-HG-2.0
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 3 de 7
Certificat contrôle qualité
Numéro de projet : C-113496
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136105
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Argent (Ag) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Cobalt (Co) mg/Kg
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
<0.6
DMR.0231-2012-1
36907
97.1%
32300 - 43700
<0.1
<2
D-076-540
33.0
95.9%
29.2 - 39.6
<0.05
D-076-540
75.6
80%
73.2 - 115.8
<0.01
D-076-540
189
86.8%
142 - 192
<0.01
D-076-540
98
92.5%
82 - 130
<0.005
D-076-540
63.1
95.7%
46.9 - 74.1
<0.05
D-076-540
63.8
90.6%
59.8 - 81.0
<0.05
D-076-540
111
91.2%
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 4 de 7
Certificat contrôle qualité
Numéro de projet : C-113496
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136105
Lieu de prélèvement : Radisson
Paramètres
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Mercure (Hg) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Plomb (Pb) mg/Kg
Sélénium (Se) mg/Kg
Zinc (Zn) mg/Kg
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
87 - 117
<5
D-076-540
85.0
93.2%
67.7 - 91.5
<0.5
<0.05
D-076-540
328
84.1%
219 - 347
<0.1
DORM-3
0.40
97.6%
0.34 - 0.48
<0.05
D-076-540
50.0
80.1%
32.3 - 51.1
<0.05
D-076-540
68.7
80.7%
44.6 - 70.6
<0.05
D-076-540
100.0
91.1%
71.1 - 112.5
<0.05
D-076-540
77.4
89.6%
67.0 - 105.8
<0.05
D-076-540
152
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 5 de 7
Certificat contrôle qualité
Numéro de projet : C-113496
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136105
Lieu de prélèvement : Radisson
Paramètres
Justesse 91.4%
Intervalle 119 - 161
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 6 de 7
Informations supplémentaires
Numéro de projet : C-113496
Date de prélèvement : 17 avril 2012
Heure de prélèvement : N/D
Échantillon : 136105
Lieu de prélèvement : Radisson
Méthode laboratoire
Méthode de référence
M-MET-3.0
MA.200-Mét. 1.2
M-MET-4.0
EPA-3050b
M-HG-2.0
MA.207-Hg 2.0
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 7 de 7
Certificat d'analyse
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-113498
Lieu de prélèvement : Radisson
Date de prélèvement : 17 mai 2012
Échantillon : 136106
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.
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 1 de 7
Certificat d'analyse
Numéro de projet : C-113498
Date de prélèvement : 17 mai 2012
Heure de prélèvement : N/D
Échantillon : 136106
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Préparation d'échantillon
Sélénium (Se)
Zinc (Zn)
Potentiel générateur acide
Soufre
Potentiel d'acidité maximal (PA)
Potentiel neutralisation brut (PN)
Potentiel neutralisaton net (PNN)
% Humidité
Argent (Ag)
Mercure (Hg)
Résultats
40183
10.1
16.2(A-B)
8.6(< A)
<0.01
<0.005(< A)
173(A-B)
45.7(A-B)
91(A-B)
74008
1124(B-C)
7.6(A-B)
55.9(A-B)
7.6(< A)
Méthode d'analyse
Date d'analyse
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
30 mai 2012
30 mai 2012
05 juin 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
30 mai 2012
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
Sous-traitance\Laboratoire Expert Inc.
<0.05(< A) mg/Kg
106(< A) mg/Kg
M-MET-3.0
M-MET-3.0
05 juin 2012
30 mai 2012
Sous-traitance\Maxxam Analytics Inc
0.33
10.3
149
139
<0.1
<2
<0.1
%
kg CaCO3/t
kg CaCO3/t
kg CaCO3/t
%
mg/Kg
mg/Kg
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
Sous-traitance\Maxxam Analytics Inc
M-HUM-1.0
M-MET-4.0
M-HG-2.0
12 juin 2012
08 juin 2012
08 juin 2012
08 juin 2012
29 mai 2012
30 mai 2012
05 juin 2012
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 2 de 7
Limite de détection rapportée
Numéro de projet : C-113498
Date de prélèvement : 17 mai 2012
Heure de prélèvement : N/D
Échantillon : 136106
Lieu de prélèvement : Radisson
Paramètre
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
% Humidité
Argent (Ag)
Mercure (Hg)
Valeur
0.6
0.1
0.05
0.01
0.01
0.005
0.05
0.05
5
0.5
0.05
0.05
0.05
0.05
0.05
0.05
0.1
2
0.1
Unité
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
%
mg/Kg
mg/Kg
Méthode
Accréditation
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-HUM-1.0
M-MET-4.0
M-HG-2.0
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Oui
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 3 de 7
Certificat contrôle qualité
Numéro de projet : C-113498
Date de prélèvement : 17 mai 2012
Heure de prélèvement : N/D
Échantillon : 136106
Lieu de prélèvement : Radisson
Paramètres
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Argent (Ag) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Cobalt (Co) mg/Kg
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
<0.6
DMR.0231-2012-1
36907
97.1%
32300 - 43700
<0.1
<2
D-076-540
33.0
95.9%
29.2 - 39.6
<0.05
D-076-540
75.6
80%
73.2 - 115.8
<0.01
D-076-540
189
86.8%
142 - 192
<0.01
D-076-540
98
92.5%
82 - 130
<0.005
D-076-540
63.1
95.7%
46.9 - 74.1
<0.05
D-076-540
63.8
90.6%
59.8 - 81.0
<0.05
D-076-540
111
91.2%
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 4 de 7
Certificat contrôle qualité
Numéro de projet : C-113498
Date de prélèvement : 17 mai 2012
Heure de prélèvement : N/D
Échantillon : 136106
Lieu de prélèvement : Radisson
Paramètres
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Mercure (Hg) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Plomb (Pb) mg/Kg
Sélénium (Se) mg/Kg
Zinc (Zn) mg/Kg
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
Justesse
Intervalle
Blanc
Nom Standard
Valeur obtenue
87 - 117
<5
D-076-540
85.0
93.2%
67.7 - 91.5
<0.5
<0.05
D-076-540
328
84.1%
219 - 347
<0.1
DORM-3
0.40
97.6%
0.34 - 0.48
<0.05
D-076-540
50.0
80.1%
32.3 - 51.1
<0.05
D-076-540
68.7
80.7%
44.6 - 70.6
<0.05
D-076-540
100.0
91.1%
71.1 - 112.5
<0.05
D-076-540
77.4
89.6%
67.0 - 105.8
<0.05
D-076-540
152
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 5 de 7
Certificat contrôle qualité
Numéro de projet : C-113498
Date de prélèvement : 17 mai 2012
Heure de prélèvement : N/D
Échantillon : 136106
Lieu de prélèvement : Radisson
Paramètres
Justesse 91.4%
Intervalle 119 - 161
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 6 de 7
Informations supplémentaires
Numéro de projet : C-113498
Date de prélèvement : 17 mai 2012
Heure de prélèvement : N/D
Échantillon : 136106
Lieu de prélèvement : Radisson
Méthode laboratoire
Méthode de référence
M-MET-3.0
MA.200-Mét. 1.2
M-MET-4.0
EPA-3050b
M-HG-2.0
MA.207-Hg 2.0
Sauf indication contraire, tous les échantillons ont été reçus en bon état.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
F-02-06
Version 3ième: 26/10/2005
Page 7 de 7
Sommaire des résultats
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
No Multilab Direct
Échantillon
Date prélèvement
Aluminium (Al) mg/L
Antimoine (Sb) mg/L
Argent (Ag) mg/L
Arsenic (As) mg/L
Baryum (Ba) mg/L
Bore (lixiviation) mg/L
Cadmium (Cd) mg/L
Chrome (Cr) mg/L
Cobalt (Co) mg/L
Cuivre (Cu) mg/L
Fer (Fe) mg/L
Lixiviation (TCLP)
Manganèse (Mn) mg/L
Mercure (Hg) mg/L
Molybdene (Mo) mg/L
Nickel (Ni) mg/L
Plomb (Pb) mg/L
Sélénium (Se) mg/L
Zinc (Zn) mg/L
% Humidité %
Date de réception : 18 mai 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
113485
136099
17-04-2012
0.425
0.0030
<0.0005
0.5905
0.1302
<0.01
0.0003
0.0225
0.004
0.0023
6.4
113487
136100
17-04-2012
0.032
0.0076
<0.0005
0.1070
0.1285
<0.01
0.0001
0.0066
0.003
0.0060
2.0
113489
136101
17-04-2012
<0.006
<0.0001
<0.0005
0.1591
0.3346
<0.01
<0.0001
<0.0006
<0.001
0.0019
19.3
113491
136102
17-04-2012
<0.006
<0.0001
<0.0005
<0.0005
0.2065
<0.01
<0.0001
<0.0006
<0.001
0.0196
2.4
113493
136103
17-04-2012
<0.006
<0.0001
<0.0005
0.0501
0.0365
<0.01
<0.0001
<0.0006
0.009
0.0070
40.9
113495
136104
17-04-2012
<0.006
<0.0001
<0.0005
0.2080
0.2770
<0.01
<0.0001
<0.0006
<0.001
0.0197
4.3
113497
136105
17-04-2012
<0.006
<0.0001
<0.0005
<0.0005
0.4235
<0.01
<0.0001
<0.0006
<0.001
0.0202
4.4
113499
136106
17-05-2012
0.344
0.0011
<0.0005
<0.0005
0.0190
<0.01
0.0010
0.0238
0.008
0.0037
13.8
8.387
<0.00002
0.0029
0.0638
0.0353
0.001
0.039
<0.1
7.702
<0.00002
0.0025
0.0776
<0.0005
<0.001
0.030
<0.1
18.85
<0.00002
<0.0005
0.0285
<0.0005
<0.001
<0.001
<0.1
1.958
<0.00002
<0.0005
0.0148
<0.0005
<0.001
<0.001
<0.1
12.72
<0.00002
<0.0005
0.1271
<0.0005
<0.001
<0.001
<0.1
3.140
<0.00002
<0.0005
0.0163
<0.0005
<0.001
<0.001
<0.1
4.865
<0.00002
<0.0005
0.0029
<0.0005
<0.001
<0.001
<0.1
16.01
<0.00002
<0.0005
0.0660
<0.0005
<0.001
0.017
<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.
F-02-13
Version 3ième: 30/03/2011
Date d'émission : 05 juin 2012
Page 1 de 4
Sommaire des résultats
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
No Multilab Direct
Échantillon
Date prélèvement
pH
113485
136099
17-04-2012
5.53
113487
136100
17-04-2012
6.35
Date de réception : 18 mai 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
113489
136101
17-04-2012
5.67
113491
136102
17-04-2012
5.47
113493
136103
17-04-2012
5.02
113495
136104
17-04-2012
5.26
113497
136105
17-04-2012
6.2
113499
136106
17-05-2012
5.73
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.
F-02-13
Version 3ième: 30/03/2011
Date d'émission : 05 juin 2012
Page 2 de 4
Sommaire des résultats
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
Date de réception : 18 mai 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
Limite de détection rapportée
Valeur
Paramètres
Aluminium (Al)
Antimoine (Sb)
Argent (Ag)
Arsenic (As)
Baryum (Ba)
Bore (lixiviation)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Lixiviation (TCLP)
Manganèse (Mn)
Mercure (Hg)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
% Humidité
0.006
0.0001
0.0005
0.0005
0.0005
0.01
0.0001
0.0006
0.001
0.0005
0.01
N.D.
0.0005
0.00002
0.0005
0.0005
0.0005
0.001
0.001
0.1
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
%
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-HUM-1.0
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.
F-02-13
Version 3ième: 30/03/2011
Date d'émission : 05 juin 2012
Page 3 de 4
Sommaire des résultats
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
Date de réception : 18 mai 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
Limite de détection rapportée
Paramètres
pH
Valeur
N.D.
M-LIX-1.0
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.
F-02-13
Version 3ième: 30/03/2011
Date d'émission : 05 juin 2012
Page 4 de 4
Certificat contrôle qualité
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
Paramètres
Blanc
Numéro de projet : Multiple
Date de réception : 18 mai 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
Nom
Standard
Obtenue Intervalle
Aluminium (Al) mg/L
<0.006
Aluminium (Al) mg/L
<0.006
Antimoine (Sb) mg/L
<0.0001
Antimoine (Sb) mg/L
<0.0001
Argent (Ag) mg/L
<0.0005
Argent (Ag) mg/L
<0.0005
Arsenic (As) mg/L
<0.0005
Arsenic (As) mg/L
<0.0005
Baryum (Ba) mg/L
<0.0005
Baryum (Ba) mg/L
<0.0005
Bore (lixiviation) mg/L
<0.01
Bore (lixiviation) mg/L
<0.01
Cadmium (Cd) mg/L
<0.0001
Cadmium (Cd) mg/L
<0.0001
Chrome (Cr) mg/L
<0.0006
Chrome (Cr) mg/L
<0.0006
Cobalt (Co) mg/L
<0.001
Cobalt (Co) mg/L
<0.001
Cuivre (Cu) mg/L
<0.0005
Cuivre (Cu) mg/L
<0.0005
Projet: 113485,113487,113489,113491,113493,113495,113497,113499
Duplicata
1
2
<0.006
<0.006
<0.0001
<0.0005
<0.0001
<0.0005
0.1591
0.1246
0.3346
0.3346
<0.01
<0.01
<0.0001
<0.0001
<0.0006
<0.0006
<0.001
0.0019
<0.001
0.0013
Les résultats ne se rapportent qu'aux échantillons soumis pour analyse.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
Les échantillons seront conservés pendant 30 jours à partir de la date du rapport à moins d'avis écrit du client.
Date d'émission : 05 juin 2012
F-02-15
Version 3ième: 17/11/2011
Page 1 de 2
Certificat contrôle qualité
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
Paramètres
Blanc
Fer (Fe) mg/L
Fer (Fe) mg/L
Manganèse (Mn) mg/L
Manganèse (Mn) mg/L
Mercure (Hg) mg/L
Mercure (Hg) mg/L
Molybdene (Mo) mg/L
Molybdene (Mo) mg/L
Nickel (Ni) mg/L
Nickel (Ni) mg/L
pH
pH
Plomb (Pb) mg/L
Plomb (Pb) mg/L
Sélénium (Se) mg/L
Sélénium (Se) mg/L
Zinc (Zn) mg/L
Zinc (Zn) mg/L
Numéro de projet : Multiple
Date de réception : 18 mai 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
Nom
Standard
Obtenue Intervalle
<0.01
<0.01
<0.0005
<0.0005
<0.00002
<0.00002
<0.0005
<0.0005
<0.0005
<0.0005
<0.0005
<0.0005
<0.001
<0.001
<0.001
<0.001
Duplicata
1
2
19.3
19.3
18.85
18.85
<0.00002
<0.0005
<0.00002
<0.0005
0.0285
0.0285
5.67
5.67
<0.0005
<0.0005
<0.001
<0.001
<0.001
<0.001
Projet: 113485,113487,113489,113491,113493,113495,113497,113499
Les résultats ne se rapportent qu'aux échantillons soumis pour analyse.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
Les échantillons seront conservés pendant 30 jours à partir de la date du rapport à moins d'avis écrit du client.
Date d'émission : 05 juin 2012
F-02-15
Version 3ième: 17/11/2011
Page 2 de 2
Sommaire des résultats
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
No Multilab Direct
Échantillon
Date prélèvement
% Humidité %
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Cobalt (Co) mg/Kg
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Plomb (Pb) mg/Kg
Potentiel générateur acide
Potentiel neutralisaton net (PNN) kg CaCO3/t
Potentiel neutralisation brut (PN) kg CaCO3/t
Potentiel d'acidité maximal (PA) kg CaCO3/t
Soufre %
Date de réception : 22 juin 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
115747
136107
22-06-2012
<0.1
21085
0.2
<0.05 (< A)
35.4 (< A)
<0.01
8.07 (B-C)
125 (A-B)
9.3 (< A)
23 (< A)
47093
473 (< A)
2.6 (< A)
88.5 (A-B)
<0.05 (< A)
115748
136108
22-06-2012
<0.1
16831
<0.1
15.2 (A-B)
<0.01 (< A)
<0.01
0.268 (< A)
81.2 (< A)
33.0 (A-B)
146 (B-C)
60180
1373 (B-C)
1.0 (< A)
53.4 (A-B)
<0.05 (< A)
115749
136109
22-06-2012
<0.1
16955
2.0
29.5 (A-B)
122 (< A)
<0.01
0.158 (< A)
170 (A-B)
5.0 (< A)
19 (< A)
37423
385 (< A)
1.5 (< A)
84.0 (A-B)
<0.05 (< A)
115750
136110
22-06-2012
<0.1
15252
1.8
18144 (> C)
69.4 (< A)
<0.01
0.182 (< A)
82.3 (< A)
22.2 (A-B)
97 (A-B)
68876
1384 (B-C)
0.92 (< A)
46.2 (< A)
<0.05 (< A)
115751
136111
22-06-2012
<0.1
19168
1.0
17737 (> C)
188 (< A)
<0.01
9.90 (B-C)
116 (A-B)
25.1 (A-B)
52 (A-B)
69851
1376 (B-C)
1.0 (< A)
42.9 (< A)
<0.05 (< A)
115752
136112
22-06-2012
<0.1
18404
1.4
13615 (> C)
128 (< A)
<0.01
0.122 (< A)
81.2 (< A)
6.6 (< A)
52 (A-B)
47200
802 (< A)
1.9 (< A)
37.2 (< A)
<0.05 (< A)
115753
136113
22-06-2012
<0.1
14402
0.5
13833 (> C)
100 (< A)
<0.01
0.145 (< A)
74.6 (< A)
4.5 (< A)
60 (A-B)
47263
987 (< A)
1.0 (< A)
34.0 (< A)
<0.05 (< A)
115754
136114
22-06-2012
<0.1
16620
2.0
11659 (> C)
162 (< A)
<0.01
0.127 (< A)
136 (A-B)
14.2 (< A)
42 (< A)
55753
1038 (B-C)
1.0 (< A)
44.2 (< A)
<0.05 (< A)
28.7
34
5.3
0.17
285
314
19
0.61
18.0
24
6
0.19
87.3
157
69.7
2.2
141
207
66.3
2.1
33.7
83
49.3
1.6
74.9
126
51.1
1.6
120
177
57.5
1.8
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.
F-02-13
Version 3ième: 30/03/2011
Date d'émission : 17 juillet 2012
Page 1 de 3
Sommaire des résultats
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
No Multilab Direct
Échantillon
Date prélèvement
Préparation d'échantillon
Sélénium (Se) mg/Kg
Zinc (Zn) mg/Kg
Argent (Ag) mg/Kg
Mercure (Hg) mg/Kg
Date de réception : 22 juin 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
115747
136107
22-06-2012
115748
136108
22-06-2012
115749
136109
22-06-2012
115750
136110
22-06-2012
115751
136111
22-06-2012
115752
136112
22-06-2012
115753
136113
22-06-2012
115754
136114
22-06-2012
<0.05 (< A)
57.8 (< A)
<2
<0.1
<0.05 (< A)
10.4 (< A)
<2
<0.1
<0.05 (< A)
71.2 (< A)
<2
<0.1
<0.05 (< A)
64.8 (< A)
<2
<0.1
<0.05 (< A)
<0.05 (< A)
<2
<0.1
5.6 (B-C)
38.2 (< A)
<2
<0.1
<0.05 (< A)
33.0 (< A)
<2
<0.1
<0.05 (< A)
8.1 (< A)
2
<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.
F-02-13
Version 3ième: 30/03/2011
Date d'émission : 17 juillet 2012
Page 2 de 3
Sommaire des résultats
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
Date de réception : 22 juin 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
Limite de détection rapportée
Paramètres
% Humidité
Aluminium (Al)
Antimoine (Sb)
Arsenic (As)
Baryum (Ba)
Bore (B)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Manganèse (Mn)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
Argent (Ag)
Mercure (Hg)
Valeur
0.1
0.6
0.1
0.05
0.01
0.01
0.005
0.05
0.05
5
0.5
0.05
0.05
0.05
0.05
0.05
0.05
2
0.1
%
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
mg/Kg
M-HUM-1.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-3.0
M-MET-4.0
M-HG-2.0
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.
F-02-13
Version 3ième: 30/03/2011
Date d'émission : 17 juillet 2012
Page 3 de 3
Certificat contrôle qualité
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 : Multiple
Date de réception : 22 juin 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
Standard
Obtenue Intervalle
Paramètres
Blanc
Nom
Aluminium (Al) mg/Kg
Antimoine (Sb) mg/Kg
Argent (Ag) mg/Kg
Arsenic (As) mg/Kg
Baryum (Ba) mg/Kg
Bore (B) mg/Kg
Cadmium (Cd) mg/Kg
Chrome (Cr) mg/Kg
Cobalt (Co) mg/Kg
Cuivre (Cu) mg/Kg
Fer (Fe) mg/Kg
Manganèse (Mn) mg/Kg
Mercure (Hg) mg/Kg
Molybdene (Mo) mg/Kg
Nickel (Ni) mg/Kg
Plomb (Pb) mg/Kg
Sélénium (Se) mg/Kg
Zinc (Zn) mg/Kg
<0.6
<0.1
<2
<0.05
<0.01
<0.01
<0.005
<0.05
<0.05
<5
<0.5
<0.05
<0.1
<0.05
<0.05
<0.05
<0.05
<0.05
D-076-540
7551
7140 - 9660
D-076-540
D-076-540
D-076-540
D-076-540
D-076-540
D-076-540
D-076-540
D-076-540
D-076-540
D-076-540
38.0
94.6
187
127
66.1
60.3
108
73.0
11966
302
29.2 - 39.6
73.2 - 115.8
142 - 192
82 - 130
46.9 - 74.1
59.8 - 81.0
87 - 117
67.7 - 91.5
9688 - 15313
219 - 347
D-076-540
D-076-540
D-076-540
D-076-540
D-076-540
36.9
47.3
79.0
90.5
143
32.3 - 51.1
44.6 - 70.6
71.1 - 112.5
67.0 - 105.8
119 - 161
Duplicata
1
2
Projet: 115747:115754
Les résultats ne se rapportent qu'aux échantillons soumis pour analyse.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
Les échantillons seront conservés pendant 30 jours à partir de la date du rapport à moins d'avis écrit du client.
Date d'émission : 17 juillet 2012
F-02-15
Version 3ième: 17/11/2011
Page 1 de 1
Sommaire des résultats
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
No Multilab Direct
Échantillon
Date prélèvement
Aluminium (Al) mg/L
Antimoine (Sb) mg/L
Argent (Ag) mg/L
Arsenic (As) mg/L
Baryum (Ba) mg/L
Bore (lixiviation) mg/L
Cadmium (Cd) mg/L
Chrome (Cr) mg/L
Cobalt (Co) mg/L
Cuivre (Cu) mg/L
Fer (Fe) mg/L
Lixiviation (TCLP)
Manganèse (Mn) mg/L
Mercure (Hg) mg/L
Molybdene (Mo) mg/L
Nickel (Ni) mg/L
Plomb (Pb) mg/L
Sélénium (Se) mg/L
Zinc (Zn) mg/L
Date de réception : 22 juin 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
115755
136107
22-06-2012
0.703
0.0025
<0.0005
<0.0005
0.0683
<0.01
<0.0001
0.0360
0.006
0.0239
7.3
115756
136108
22-06-2012
0.030
0.0003
<0.0005
<0.0005
0.0275
<0.01
0.0012
0.0293
0.026
0.0014
11.1
115757
136109
22-06-2012
1.95
0.0014
<0.0005
<0.0005
0.1479
<0.01
<0.0001
0.0421
0.004
0.0254
7.5
115758
136110
22-06-2012
0.219
0.0072
<0.0005
1.504
0.1896
<0.01
0.0007
0.0242
0.025
0.0019
13.3
115759
136111
22-06-2012
0.038
0.0074
<0.0005
0.9565
0.3855
<0.01
0.0008
0.0296
0.022
0.0012
12.7
115760
136112
22-06-2012
<0.006
0.0040
<0.0005
0.3325
0.2319
<0.01
0.0005
0.0289
0.009
0.0226
1.1
115761
136113
22-06-2012
1.66
0.0063
<0.0005
3.956
0.2770
<0.01
0.0013
0.0646
0.027
0.0017
23.1
115762
136114
22-06-2012
0.189
0.0081
<0.0005
1.086
0.3477
<0.01
0.0005
0.0372
0.024
0.0012
13.5
0.9857
<0.00002
0.0021
0.1142
0.0118
<0.001
<0.001
12.30
<0.00002
0.0007
0.4630
0.0013
<0.001
0.004
0.7625
<0.00002
0.0022
0.0573
0.0013
<0.001
0.02
17.70
<0.00002
<0.0005
0.4904
0.0152
<0.001
0.034
19.98
<0.00002
<0.0005
0.4821
0.0073
<0.001
0.015
8.891
<0.00002
0.0021
0.1917
0.0035
<0.001
0.032
21.35
<0.00002
0.0009
0.4249
0.0281
<0.001
0.037
17.14
<0.00002
<0.0005
0.4978
0.0110
<0.001
0.009
Certificat corrigé, remplace le certificat multiple émis le 04 juillet 2012.
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.
F-02-13
Version 3ième: 30/03/2011
Date d'émission : 26 juillet 2012
Page 1 de 2
Sommaire des résultats
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
Date de réception : 22 juin 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
Limite de détection rapportée
Paramètres
Aluminium (Al)
Antimoine (Sb)
Argent (Ag)
Arsenic (As)
Baryum (Ba)
Bore (lixiviation)
Cadmium (Cd)
Chrome (Cr)
Cobalt (Co)
Cuivre (Cu)
Fer (Fe)
Lixiviation (TCLP)
Manganèse (Mn)
Mercure (Hg)
Molybdene (Mo)
Nickel (Ni)
Plomb (Pb)
Sélénium (Se)
Zinc (Zn)
Valeur
0.006
0.0001
0.0005
0.0005
0.0005
0.01
0.0001
0.0006
0.001
0.0005
0.01
N.D.
0.0005
0.00002
0.0005
0.0005
0.0005
0.001
0.001
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
mg/L
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
M-LIX-1.0
Certificat corrigé, remplace le certificat multiple émis le 04 juillet 2012.
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.
F-02-13
Version 3ième: 30/03/2011
Date d'émission : 26 juillet 2012
Page 2 de 2
Certificat contrôle qualité
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
Paramètres
Blanc
Aluminium (Al) mg/L
Antimoine (Sb) mg/L
Argent (Ag) mg/L
Arsenic (As) mg/L
Baryum (Ba) mg/L
Bore (lixiviation) mg/L
Cadmium (Cd) mg/L
Chrome (Cr) mg/L
Cobalt (Co) mg/L
Cuivre (Cu) mg/L
Fer (Fe) mg/L
Manganèse (Mn) mg/L
Mercure (Hg) mg/L
Molybdene (Mo) mg/L
Nickel (Ni) mg/L
Plomb (Pb) mg/L
Sélénium (Se) mg/L
Zinc (Zn) mg/L
Numéro de projet : Multiple
Date de réception : 22 juin 2012
Nom du préleveur : Eugène Gauthier
Type d'échantillon : Minerai
Nom
Standard
Obtenue Intervalle
Duplicata
1
2
<0.006
<0.0001
<0.0005
<0.0005
<0.0005
<0.01
<0.0001
<0.0006
<0.001
<0.0005
<0.01
<0.0005
<0.00002
<0.0005
<0.0005
<0.0005
<0.001
<0.001
Projet: 115755:115762
Certificat corrigé, remplace le certificat multiple émis le 04 juillet 2012.
Les résultats ne se rapportent qu'aux échantillons soumis pour analyse.
Toute reproduction, sinon en entier, est interdite sans l'autorisation écrite du laboratoire.
Les échantillons seront conservés pendant 30 jours à partir de la date du rapport à moins d'avis écrit du client.
Date d'émission : 26 juillet 2012
F-02-15
Version 3ième: 17/11/2011
Page 1 de 1
Annexe C
Exigences de rejet – Directive 019
Colonne I
Concentration moyenne
acceptable
(moyenne arithmétique
mensuelle)
Colonne II
Concentration maximale
acceptable dans un
échantillon instantané
Arsenic extractible
0,200 mg/l
0,400 mg/l
Cuivre extractible
0,300 mg/l
0,600 mg/l
Fer extractible
3,000 mg/l
6,000 mg/l
Nickel extractible
0,500 mg/l
1,000 mg/l
Plomb extractible
0,200 mg/l
0,400 mg/l
Zinc extractible
0,500 mg/l
1,000 mg/l
Cyanures totaux
1,000 mg/l
2,000 mg/l
---------
2,000 mg/l
15,000 mg/l
30,000 mg/l
Paramètre
Hydrocarbures
(C10C50)
Matières en
suspension
* 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
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