TECHNICAL REPORT AND PRELIMINARY ECONOMIC … · Richmont Mines at Francoeur and Beaufor mines,...

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

TECHNICAL REPORT AND PRELIMINARY ECONOMIC ASSESSMENT FOR THE LAMAQUE PROJECT

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

Project Location

Bourlamaque Township, Province of Québec, Canada

(NTS: 32C/04) (UTM: 293960E, 5329260N)

(NAD 83, Zone 18)

Prepared for

Integra Gold Corp. 1101 3e Avenue Est,

Val-d’Or, Québec, Canada, J9P 4P5

Prepared by:

Sylvie Poirier, Eng. Laurent Roy, Eng.

Christian D’Amours, P.Geo. Michel Caron, Eng.

InnovExplo Inc. Val-d’Or (Québec)

GéoPointCom Inc, Val-d’Or (Québec)

WSP Canada Inc. Val-d’Or (Québec)

Stephan Bergeron, P. Geo., M.Eng. AMEC Environment and Infrastructure Dorval (Québec)

Daniel Gaudreault, Eng. Geologica Groupe-Conseil Inc., Val-d’Or (Québec)

Effective Date: 28 February 2014 Signature Date: 25 April 2014

www.innovexplo.com

2014NI43‐101TechnicalReportontheLamaquepropertyApril25,2014IntegraGoldCorp. 2

CERTIFICATE OF AUTHOR – DANIEL GAUDREAULT I, Daniel Gaudreault, Eng, do hereby certify that: 1. I am currently employed as a geological engineer by Geologica Groupe-Conseil Inc., 450, 3rd

Avenue, Suite 202, P.O. Box 1891, Val-d’Or (Québec), J9P 6C5.

2. I graduated with a degree in Geological Engineering (“Eng.”) from the University of Québec in Chicoutimi in 1983.

3. I am a member of the “Ordre des ingénieurs du Québec (OIQ)”, #39834, of the Québec Mining Exploration Association (AEMQ) and the Prospectors and Developers Association of Canada (PDAC).

4. I have worked as a geologist for a total of 31 years since my graduation from university. As an engineer specializing in geology and mining, I have been involved with all aspects of planning, organization and supervision of mineral exploration projects, especially in remote areas of Abitibi, Québec. I have been in charge of teams of professionals and technicians on geological projects in the most severe conditions. I have also completed several geoscientific compilations and technical reports on areas of interest in Québec, Ontario, USA (California & Nevada) and South America (mainly Peru).

5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, 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 technical parts of Items 4, 5, 6, 7, 8, 9, 10, 11, 12 and 23, and co-author of Items 1, 25 to 27 of the report titled “Technical Report and Preliminary Economic Assessment for the Lamaque Project (according to National Instrument 43-101 and Form 43-101F1)”, effective date of February 28, 2014 and signature date of April 25, 2014, prepared for Integra Gold Corp. I have recently visited the subject property and I have planned, organized, supervised and revised the drilling programs between 2009 and 2014.

7. 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.

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

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

Signed on this 25th day of April, 2014 (Original signed and sealed)

Daniel Gaudreault, Eng. (OIQ #39834) Géologica Groupe-Conseil 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 eighteen (18) years since graduating from university. My mining expertise was acquired while working for Talpa Mining Contractor, Richmont Mines at Francoeur and Beaufor mines, Doyon and CasaBerardi mines. I have been a consulting engineer for InnovExplo Inc. since September 2012.

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

6. I am responsible for the preparation of Sections 2, 3, 15, 16, 18, 19, 21, 22, 24, and co-author of Sections 1 and 25 to 27 of the report titled “Technical Report and Preliminary Economic Assessment for the Lamaque Project (according to National Instrument 43-101 and Form 43-101F1)” (the “Technical Report”), effective date of February 28, 2014 and signature date of April 25, 2014, prepared for Integra Gold Corp.


8. I visited the Lamaque Project site on July 9, 2013, accompanied by François Chabot of Integra Gold and Marie-Claire Dagenais of InnovExplo.


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

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


Laurent Roy, Eng. InnovExplo Inc.

www.innovexplo.com


CERTIFICATE OF AUTHOR – MICHEL GARON

I, Michel Garon, do hereby certify that:

1. I am an Engineer employed as Senior Mining Engineer by WSP at 152, Murdoch Avenue, Rouyn-Noranda, Quebec.

2. I received a Bachelor’s Degree in Applied Sciences from the Université de Montréal (Montreal, Quebec) in 1975 and a Master’s Degree from the same university in 1976.

3. I am a registered member of the Ordre des Ingénieurs du Québec (OIQ member no. 28151).

4. I have over 30 years of experience as an engineer in the mining industry. My experience has been acquired mostly with Noranda. I have been working for WSP since June 2006 as Senior Mining Engineer.

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

6. I had prior involvement with the property that is the subject of the Technical Report. I did not visit the Lamaque property.

7. I am responsible for the preparation of Mineral and Metallurgical section (sections 13 and 17) and co-author of sections 1 and 25 to 27 of the Technical Report entitled “Technical Report and Preliminary Economic Assessment for the Lamaque Project (according to National Instrument 43-101 and Form 43-101F1)”, effective date of February 28, 2014 and signature date of April 25, 2014, prepared for Integra Gold Corp.

8. I am “independent” (as such term is defined in Section 1.5 of R 43-101 or NI 43-101) of Integra Gold Mining.

9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10. I have read R 43-101, Appendix 43-101F1 and the Technical Report which has been prepared in compliance with that instrument and form.

11. I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, or presentation of excerpts or a summary, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated in Val-d’Or, Quebec, this 25th day of April 2014.

(Original signed and sealed)

Michel Garon, Eng. WSP

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 1994.

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 nineteen (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 responsible for supervising the preparation of Sections 1,2,3,15,16, 18, 19, 21, 22, 24, 25, 26, 27 and co-author of Sections 1 and 25 to 27 of the report titled “Technical Report and Preliminary Economic Assessment for the Lamaque Project (according to National Instrument 43-101 and Form 43-101F1)” (the “Technical Report”), effective date of February 28, 2014 and signature date of April 25, 2014, prepared for Integra Gold Corp.


8. I did not visit the Lamaque property.

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 makes the Technical Report misleading.

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

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


Sylvie Poirier, Eng. InnovExplo Inc.

www.innovexplo.com


CERTIFICATE OF AUTHOR – Christian D’Amours

I, Christian D’Amours, residing at 895, rue Lévis, Val-d’Or, Québec, do hereby certify that: 1. I am an independent geologist with the consulting firm, GeoPointCom, located at 895 rue

Levis, Val d'Or, Québec, Canada, J9P 4B8.

2. I graduated in geology, as a professional geologist, from the University of Québec in Montréal.

3. I have been practising the profession of geologist on an ongoing basis since May 1985

4. From 1985 to 1994 the practice of my profession was mainly oriented towards exploration. From 1994 to 1999 I worked primarily in the field of mining. Since 1999, I have been working predominantly in the evaluation of resources, reserves and geostatistics.

5. I am a member of the Order of Geologists of Québec (#226);

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

7. I am responsible for the preparation of Section 14 and co-author of sections 1 and 25 to 27 of the report titled “Technical Report and Preliminary Economic Assessment for the Lamaque Project (according to National Instrument 43-101 and Form 43-101F1)” (the “Technical Report”), effective date of February 28, 2014 and signature date of April 25, 2014, prepared for Integra Gold Corp.

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

9. I did not visit the Lamaque property of Integra Gold Corp.


11. I am independent of the owners of the lands covered by this report within the meaning of section 1.5 of National Instrument 43-101 Standards of Disclosure for Mineral Properties (“NI 43-101”).

12. I have read the NI 43-101 and Form 43-101F1, and hereby certify that this report has been prepared in compliance with NI 43-101 and Form 43-101F1. The report gives a true picture of the state of scientific and technical knowledge as of February 28, 2014.


Christian D’Amours, P. Geo., OGQ GéoPointCom Inc

www.innovexplo.com


CERTIFICATE OF AUTHOR – STEPHAN BERGERON

I, Stephan Bergeron, P. Geo., M. Eng., as an author of this report, do hereby certify that: 1. I am the Director of the Environment Department with the Dorval, Quebec, office of AMEC

Environment and Infrastructure, a division of AMEC.

2. I am a graduate of University du Québec à Montréal in 1997 with a degree in Physical Geography, with a geomorphology specialization, and a Masters degree in Mineral Engineering, with an hydrogeological specialization from the École Polytechnique de Montréal in 1999. In addition I have taken specialist training in environmental site assessment and auditing from University of Sherbrooke in 2001.

3. I am a certified geologist, registered with the Ordre des Géologues du Québec (OGQ #787). My relevant experience for the purpose of the Technical Report is: Hydrogeology and surface geology; Environmental Regulations (permitting); Environmental management and monitoring programs; Environmental auditing; Management of multi-disciplinary environmental investigations.

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

5. A team of environmental specialists from AMEC Environment and Infrastructure visited the Lamaque Gold Project Site on a number of occasions during the spring to fall seasons of 2013 and reported to the QP for the purpose of the environmental baseline study which forms the basis of Item 20 of this report.

6. I am responsible for overall preparation of the section ''Environmental Considerations'' within Item 20 of this PEA report and co-author of sections 1 and 25 to 27, prepared for Integra Gold Corp., entitled “Technical Report and Preliminary Economic Assessment for the Lamaque Project (according to National Instrument 43-101 and Form 43-101F1)” (the “Technical Report”), effective date of February 28, 2014 and signature date of April 25, 2014.

7. I am independent of the Issuer applying the test set out in Section 1.5 of National Instrument 43-101 and Regulation 43-101.

8. I have had no prior involvement with the property that is the subject of this PEA report.

9. I have read National Instrument 43-101 and the Technical Report has been prepared in compliance with National Instrument 43-101 (Regulation 43-101) and Form 43-101F1.

10. To the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.


Stephan Bergeron, P.Geo., M.Eng.

AMEC Environment and Infrastructure

www.innovexplo.com


TABLE OF CONTENTS

1. SUMMARY ....................................................................................................................... 17

2. INTRODUCTION .............................................................................................................. 31

2.1. Terms of Reference and Scope of Work ....................................................................... 31 2.2. Principal Sources of Information ................................................................................... 31 2.3. Qualified Persons and Inspection on the Property ........................................................ 33 2.4. Units and Currencies .................................................................................................... 34

3. RELIANCE ON OTHER EXPERTS ................................................................................. 35

4. PROPERTY DESCRIPTION AND LOCATION ................................................................ 36

4.1. Location and Claims ..................................................................................................... 36 4.2. Agreements and Encumbrance .................................................................................... 41 4.3. Environmental Obligation .............................................................................................. 41 4.4. Risks and Uncertainties ................................................................................................ 42

5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ............................................................................................................ 43

6. HISTORY .......................................................................................................................... 44

6.1. Before 2003 .................................................................................................................. 44 6.2. From January 2003 to June 2013 ................................................................................. 46

7. GEOLOGICAL SETTING AND MINERALIZATION ........................................................ 48

7.1. Regional and Local Geology ......................................................................................... 48 7.2. Local and Property Geology ......................................................................................... 50 7.3. Mineralization ................................................................................................................ 53

7.3.1. North Cluster .......................................................................................................... 54 7.3.1.1. Fortune Zone (previously known as Forestel Zone) .................................................... 54 7.3.1.2. Parallel Zone (including Vein No. 10) .......................................................................... 54 7.3.1.3. No. 5 Plug .................................................................................................................... 55 7.3.1.4. No. 3 Mine (No. 1 Vein) ............................................................................................... 55

7.3.2. South Cluster ......................................................................................................... 55 7.3.2.1. No. 4 Plug .................................................................................................................... 55 7.3.2.2. Triangle Zone ............................................................................................................... 56 7.3.2.3. South Triangle Zone .................................................................................................... 56

7.3.3. West Cluster .......................................................................................................... 57 7.3.3.1. Vein No. 6 .................................................................................................................... 57 7.3.3.2. Sixteen Zone ................................................................................................................ 57

7.3.4. Other Mineralized Zones ....................................................................................... 58

8. DEPOSIT TYPES ............................................................................................................. 59

9. EXPLORATION ................................................................................................................ 61

10. DRILLING ......................................................................................................................... 62

10.1. South Triangle ............................................................................................................... 62 10.2. No. 3 Mine ..................................................................................................................... 67 10.3. Parallel Zone ................................................................................................................. 71 10.4. No. 6 Vein And Sixteen Zone ........................................................................................ 82 10.5. Additional Resource Potential ....................................................................................... 86

11. SAMPLE PREPARATION, ANALYSES AND SECURITY .............................................. 87

11.1. Results of Quality Control ............................................................................................. 90

www.innovexplo.com


11.1.1. Blanks ................................................................................................................ 90 11.1.2. Certified Reference Material (Standards) ........................................................... 90 11.1.3. Duplicates .......................................................................................................... 91

12. DATA VERIFICATION ..................................................................................................... 92

13. MINERAL PROCESSING AND METALLURGICAL TESTING ....................................... 93

13.1. Initial Testwork .............................................................................................................. 93 13.1.1. Samples Description .......................................................................................... 93 13.1.2. Sample Preparation ........................................................................................... 94 13.1.3. Sample Characterization .................................................................................... 94

13.1.3.1. Comminution Testwork ................................................................................................ 94 13.1.3.2. Mineralogy and Chemical Content .............................................................................. 95

13.1.4. Metallurgical Test Program ................................................................................ 96 13.1.4.1. Gravity and Cyanide Leach ......................................................................................... 96 13.1.4.2. Rougher Flotation ........................................................................................................ 97 13.1.4.3. Flotation, Regrind and Cyanidation ............................................................................. 98

13.2. Potential Processing Facility – Secondary Testwork .................................................... 98 13.2.1. Sample Description and Preparation ................................................................. 99 13.2.2. Properties of the Four (4) Zone Composites ...................................................... 99 13.2.3. Metallurgical Performance ................................................................................. 99

13.2.3.1. Flowsheet 1: Gravity and Carbon-in-Leach (CIL) ........................................................ 99 13.2.3.2. Flowsheets 2 and 3: Whole Sample Cyanidation and Carbon-in-Leach (CIL) .......... 100 13.2.3.3. Flowsheet 4: Flotation with Cyanidation of Concentrate ........................................... 101

13.3. Metallurgical tests summary ....................................................................................... 101

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

14.1. Resource Estimate – Fortune Zone ............................................................................ 104 14.1.1. Methodology ..................................................................................................... 105 14.1.2. Drill hole Sample Database .............................................................................. 105 14.1.3. Interpretation of Mineralized Zones .................................................................. 105 14.1.4. High Grade Capping ........................................................................................ 106 14.1.5. Compositing ..................................................................................................... 108 14.1.6. Variography ...................................................................................................... 109 14.1.7. Bulk Density ..................................................................................................... 111 14.1.8. Block Model Geometry ..................................................................................... 111 14.1.9. Mineralized Zone Block Model ......................................................................... 111 14.1.10. Grade Block Model ........................................................................................... 111 14.1.11. Resource Categories ....................................................................................... 111 14.1.12. Minimum cut-off Value ..................................................................................... 112 14.1.13. Mineral Resource Estimate Results ................................................................. 113 14.1.14. Comparison to Previous Mineral Resource Estimates ..................................... 114

14.2. Resource Estimate – Parallel Zone ............................................................................ 115 14.2.1. Methodology ..................................................................................................... 115 14.2.2. Drill Hole Sample Database ............................................................................. 115 14.2.3. Interpretation of Mineralized Zones .................................................................. 115 14.2.4. High Grade Capping ........................................................................................ 117 14.2.5. Compositing ..................................................................................................... 120 14.2.6. Variography ...................................................................................................... 120 14.2.7. Bulk Density ..................................................................................................... 121 14.2.8. Block Model Geometry ..................................................................................... 121 14.2.9. Mineralized Zone Block Model ......................................................................... 122 14.2.10. Grade Block Model ........................................................................................... 122

www.innovexplo.com


14.2.11. Resource Categories ....................................................................................... 122 14.2.12. Minimum cut-off Value ..................................................................................... 123 14.2.13. Mineral Resource Estimate Results ................................................................. 124 14.2.14. Comparison to Previous Mineral Resource Estimates ..................................... 127

14.3. Resource Estimate No. 4 Plug .................................................................................... 127 14.3.1. Methodology ..................................................................................................... 127 14.3.2. Drill Hole Database .......................................................................................... 127 14.3.3. Interpretation of Mineralized Zones .................................................................. 128 14.3.4. High Grade Capping ........................................................................................ 130 14.3.5. Compositing ..................................................................................................... 131 14.3.6. Variography ...................................................................................................... 132 14.3.7. Bulk Density ..................................................................................................... 134 14.3.8. Block Model Geometry ..................................................................................... 134 14.3.9. Mineralized Zone Block Model ......................................................................... 135 14.3.10. Grade Block Model ........................................................................................... 135 14.3.11. Resource Categories ....................................................................................... 135 14.3.12. Minimum cut-off Value ..................................................................................... 136 14.3.13. Mineral Resource Estimate Results ................................................................. 138 14.3.14. Comparison to Previous Mineral Resource Estimates ..................................... 140

14.4. Resource Estimate of the Triangle Zone .................................................................... 140 14.4.1. Methodology ..................................................................................................... 140 14.4.2. Drill hole Sample Database .............................................................................. 140 14.4.3. Interpretation of Mineralized Zones .................................................................. 140 14.4.4. High Grade Capping ........................................................................................ 142 14.4.5. Compositing ..................................................................................................... 144 14.4.6. Variography ...................................................................................................... 144 14.4.7. Bulk Density ..................................................................................................... 145 14.4.8. Block Model Geometry ..................................................................................... 145 14.4.9. Mineralized Zone Block Model ......................................................................... 145 14.4.10. Grade Block Model ........................................................................................... 145 14.4.11. Resource Categories ....................................................................................... 146 14.4.12. Minimum cut-off Value ..................................................................................... 147 14.4.13. Mineral Resource Estimate Results ................................................................. 147 14.4.14. Comparison with Previous Mineral Resource Estimates ................................. 149

14.5. Resource Estimate – Vein No. 6 ................................................................................. 149 14.5.1. Methodology ..................................................................................................... 150 14.5.2. Drill hole sample database ............................................................................... 150 14.5.3. Interpretation of Mineralized Zones .................................................................. 150 14.5.4. High grade capping .......................................................................................... 151 14.5.5. Compositing ..................................................................................................... 155 14.5.6. Variography ...................................................................................................... 155 14.5.7. Bulk Density ..................................................................................................... 156 14.5.8. Block Model Geometry ..................................................................................... 156 14.5.9. Mineralized Zone Block Model ......................................................................... 156 14.5.10. Grade block model ........................................................................................... 157 14.5.11. Resource categories ........................................................................................ 157 14.5.12. Minimum cut-off value ...................................................................................... 158 14.5.13. Mineral Resource Estimate Results ................................................................. 160 14.5.14. Comparison to previous mineral resource estimates ....................................... 160

14.6. Resource Estimate of the Sixteen Zone ..................................................................... 161

www.innovexplo.com


14.6.1. Methodology ..................................................................................................... 161 14.6.2. Drill hole sample database ............................................................................... 161 14.6.3. Interpretation of mineralized zones .................................................................. 161 14.6.4. High Grade Capping ........................................................................................ 163 14.6.5. Compositing ..................................................................................................... 165 14.6.6. Variography ...................................................................................................... 165 14.6.7. Bulk Density ..................................................................................................... 167 14.6.8. Block Model Geometry ..................................................................................... 167 14.6.9. Mineralized Zone Block Model ......................................................................... 168 14.6.10. Grade block model ........................................................................................... 168 14.6.11. Resource categories ........................................................................................ 168 14.6.12. Minimum cut-off value ...................................................................................... 169 14.6.13. Mineral Resource Estimate Results ................................................................. 169 14.6.14. Comparison to previous mineral resource estimates ....................................... 170

15. MINERAL RESERVE ESTIMATES ............................................................................... 171

16. MINING METHODS ........................................................................................................ 172

16.1. Caution to the reader .................................................................................................. 172 16.2. Introduction ................................................................................................................. 172 16.3. Mineral resources considered in the mining plan ........................................................ 172 16.4. Preliminary geotechnical assessment ......................................................................... 174

16.4.1. Rock quality designation (RQD) ....................................................................... 174 16.4.2. Total core recovery (TCR) ................................................................................ 175 16.4.3. ISRM field hardness ......................................................................................... 175 16.4.4. Fracture frequency and fracture spacing ......................................................... 176 16.4.5. Crown pillar ...................................................................................................... 177 16.4.6. Typical ground support patterns ....................................................................... 177 16.4.7. Summary of geotechnical data ......................................................................... 179

16.5. Mining method ............................................................................................................ 180 16.5.1. Long-hole method ............................................................................................ 180 16.5.2. Room and pillar ................................................................................................ 183

16.6. Existing mine infrastructure ......................................................................................... 183 16.7. Dewatering .................................................................................................................. 185 16.8. Underground mine design ........................................................................................... 185

16.8.1. North Ramp development ................................................................................ 185 16.8.2. North Ramp sequence ..................................................................................... 188 16.8.3. South Ramp development ................................................................................ 190 16.8.4. South Ramp sequence ..................................................................................... 191

16.9. Mining dilution and recoveries .................................................................................... 192 16.10. Mining rate .................................................................................................................. 192 16.11. Mine plan schedule criteria ......................................................................................... 192 16.12. Equipment ................................................................................................................... 193 16.13. Manpower requirements ............................................................................................. 194 16.14. Development and production schedule ....................................................................... 196 16.15. Mining services ........................................................................................................... 198

16.15.1. Ventilation for North and South ramps ............................................................. 198 16.15.2. Dewatering ....................................................................................................... 199

16.15.2.1. Dewatering: North Ramp ........................................................................................... 199 16.15.2.2. Dewatering: South Ramp .......................................................................................... 200

16.15.3. Compressed air ................................................................................................ 201 16.15.4. Industrial water ................................................................................................. 201

www.innovexplo.com


16.15.5. Underground power distribution ....................................................................... 202

17. RECOVERY METHODS ................................................................................................ 203

18. PROJECT INFRASTRUCTURE .................................................................................... 206

18.1. Plant and site layout .................................................................................................... 206 18.2. Lamaque Project 25 kV surface distribution ............................................................... 208 18.3. Site access .................................................................................................................. 208 18.4. Camp .......................................................................................................................... 208 18.5. Mine site entrance/guardhouse ................................................................................... 208 18.6. Office building and dry complex .................................................................................. 208 18.7. Service buildings ......................................................................................................... 209 18.8. Site Roads .................................................................................................................. 209 18.9. Compressor building ................................................................................................... 209 18.10. Fuel storage ................................................................................................................ 209 18.11. Site fencing ................................................................................................................. 209 18.12. Water systems ............................................................................................................ 209 18.13. Communication system ............................................................................................... 209 18.14. Sewage ....................................................................................................................... 210 18.15. Water treatment plant and settling pond ..................................................................... 210 18.16. Mineralized material stockpile ..................................................................................... 210 18.17. Waste stockpile ........................................................................................................... 210 18.18. Overburden stockpile .................................................................................................. 211 18.19. Project implementation schedule ................................................................................ 211

19. MARKET STUDIES AND CONTRACTS ....................................................................... 212

19.1. Market studies ............................................................................................................. 212 19.2. Contracts ..................................................................................................................... 212

20. ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT ......................................................................................................................... 213

20.1. Regulations and Permitting ......................................................................................... 213 20.2. Environmental Baseline Study .................................................................................... 215 20.3. Operations .................................................................................................................. 217 20.4. Reclamation ................................................................................................................ 218 20.5. Socio-Economic Setting .............................................................................................. 221

21. CAPITAL AND OPERATING COSTS ........................................................................... 223

21.1. Capital costs ............................................................................................................... 223 21.1.1. Surface infrastructure ....................................................................................... 223 21.1.2. Mobile equipment ............................................................................................. 225 21.1.3. Development and capitalized operating costs .................................................. 225 21.1.4. Owner’s costs ................................................................................................... 225 21.1.5. Capitalized revenue ......................................................................................... 226

21.2. Operating costs ........................................................................................................... 226 21.2.1. Definition drilling ............................................................................................... 226 21.2.2. Stope development .......................................................................................... 227 21.2.3. Mining costs ..................................................................................................... 227 21.2.4. Integra Gold staff .............................................................................................. 227 21.2.5. Energy cost ...................................................................................................... 228 21.2.6. Milling and transportation ................................................................................. 228 21.2.7. Environment ..................................................................................................... 228 21.2.8. Capitalized Opex .............................................................................................. 229

www.innovexplo.com


21.2.9. Taxes and Royalties ......................................................................................... 229

22. ECONOMIC ANALYSIS ................................................................................................. 230

22.1. Financial Analysis ....................................................................................................... 230 22.2. Sensitivity Analysis ..................................................................................................... 233

23. ADJACENT PROPERTIES ............................................................................................ 237

24. OTHER RELEVANT DATA AND INFORMATION ........................................................ 240

25. INTERPRETATION AND CONCLUSIONS .................................................................... 241

25.1. Opportunities & Risks ................................................................................................. 247

26. RECOMMENDATIONS .................................................................................................. 249

27. REFERENCES ............................................................................................................... 251

LIST OF FIGURES Figure 4.1 – Regional setting of the Lamaque Gold Project ........................................................ 37 Figure 4.2 – Detailed location map of the Lamaque Property ..................................................... 38 Figure 4.3 – Map of mining concessions and claims constituting the Lamaque Project ............. 40 Figure 7.1 – Regional Geology .................................................................................................... 49 Figure 7.2 – Property Geology and Mineralized Zones ............................................................... 51 Figure 10.1 – Diamond drilling program on the South Triangle Zone ......................................... 63 Figure 10.2 – Diamond drilling program in the No. 3 Mine area .................................................. 68 Figure 10.3 – 2013 diamond drilling program on the Parallel Zone ............................................ 73 Figure 10.4 – 2010 diamond drilling program on the Sixteen Zone ............................................ 84 Figure 10.5 – Diamond Drilling Program 2012 on No. 6 Vein ..................................................... 85 Figure 14.1 – Wireframe solids of the Fortune Zone (formerly the Forestel Zone). .................. 106 Figure 14.2 – Comparative histogram. ...................................................................................... 107 Figure 14.3 – Probability plot ..................................................................................................... 108 Figure 14.4 – Omnidirectional variography. .............................................................................. 110 Figure 14.5 – Sensitivity of estimates to different cut-off values ............................................... 113 Figure 14.6 – Changes to the interpretation from March 2011 to May 2012. ............................ 116 Figure 14.7 – Wireframe solids of the Parallel Zone. ................................................................ 117 Figure 14.8 – Comparative histogram ....................................................................................... 119 Figure 14.9 – Capping effect on total ounces. ........................................................................... 119 Figure 14.10 – Directional variography ..................................................................................... 121 Figure 14.11 – Wireframe solids from No. 4 Plug . ................................................................... 129 Figure 14.12 – Grade distribution within the High Probability Ore. ........................................... 131 Figure 14.13 – Grade versus length. ......................................................................................... 132 Figure 14.14 – Directional variography of the Gauss transformed data. ................................... 134 Figure 14.15 – Selectivity capability according to SMU size ..................................................... 138 Figure 14.16 – Wireframe solids from the Triangle Zone. ......................................................... 141 Figure 14.17 – Probability plot. .................................................................................................. 143 Figure 14.18 – Effect of capping in resource estimation. .......................................................... 143 Figure 14.19 – Omnidirectional variography. ............................................................................ 145 Figure 14.20 – Wireframe solids from Vein No. 6 ..................................................................... 151 Figure 14.21 – Probability plot (Vein No. 6) .............................................................................. 153 Figure 14.22 – Effect of capping on resource estimation (Vein No. 6) ...................................... 154 Figure 14.23 – Omnidirectional variography (Vein No. 6) ......................................................... 156 Figure 14.24 – Wireframe solids of the Sixteen Zone. .............................................................. 162 Figure 14.25 – Capping grade distribution versus % ounces lost (Sixteen Zone) ..................... 164

www.innovexplo.com


Figure 14.26 – Grade distribution within the High Probability Ore (Sixteen Zone) .................... 165 Figure 14.27 – Omnidirectional variography of the composites (Sixteen Zone) ........................ 167 Figure 16.1 – Histograms showing distribution and cumulative percentage RQD for a)

Intrusive Rock, and b) Volcanic Rock .............................................................. 175 Figure 16.2 – Standard requirement for ramp ground support .................................................. 178 Figure 16.3 – Standard support pattern as a function of excavation dimensions. ..................... 179 Figure 16.4 – Typical isometric stope view of North Ramp and Triangle zones. ...................... 181 Figure 16.5 – Typical longitudinal stope view of North Ramp and Triangle zones .................... 182 Figure 16.6 – Typical drilling patterns in the No. 4 Plug ............................................................ 182 Figure 16.7 – Typical room and pillar design ............................................................................ 183 Figure 16.8 – Plan view of Lamaque Project existing infrastructure ......................................... 184 Figure 16.9 – Longitudinal view of Parallel Zone (incl. No. 7 veins) and Fortune Zone,

looking north. ................................................................................................... 186 Figure 16.10 – Plan view of the Parallel-Fortune connection. ................................................... 187 Figure 16.11 – Longitudinal view of the Parallel Zone without the No. 7 veins, looking north. . 189 Figure 16.12 – Longitudinal view of the Triangle Zone and No. 4 Plug, looking west. .............. 191 Figure 16.13 – Ventilation network for the North Ramp ............................................................ 198 Figure 16.14 – Ventilation network for the South Ramp ............................................................ 199 Figure 16.15 – Longitudinal view of the dewatering system for the Parallel and Fortune

zones, looking west. ........................................................................................ 200 Figure 16.16 – Longitudinal view of the Triangle Zone and No. 4 Plug dewatering system,

looking west. .................................................................................................... 201 Figure 17.1 – Typical Merrill-Crowe flowsheet .......................................................................... 204 Figure 17.2 – Typical CIP (Carbon-in-Pulp) flowsheet .............................................................. 205 Figure 18.1 – Surface view of the Lamaque Project. ................................................................ 207 Figure 22.1 – Sensitivity analysis of economical parameters, pre-tax NPV at 5% (millions $).. 233 Figure 22.2 – Sensitivity analysis of grade, pre-tax NPV at 5% (millions $) .............................. 234 Figure 22.3 – Sensitivity analysis of economical parameters, pre-tax IRR ............................... 235 Figure 22.4 – Sensitivity analysis on grade, pre-tax IRR .......................................................... 236 Figure 23.1 – Adjacent Properties ............................................................................................. 239 LIST OF TABLES Table 4.1 – Mining titles comprising the Lamaque Project .......................................................... 39 Table 6.1 – Summary of Production for Historical Lamaque Property ........................................ 45 Table 6.2 – January 2003 to June 2013: drilling & exploration ................................................... 47 Table 7.1 – Regional Stratigraphy ............................................................................................... 48 Table 10.1 – Table of technical parameters on the South Triangle Zone ................................... 62 Table 10.2 – Table of significant assay results from the 2013 drilling program on the South

Triangle Zone ..................................................................................................... 64 Table 10.3 – 2013 technical parameters in the No. 3 Mine area ................................................. 67 Table 10.4 – Table of significant assay results for the drilling campaign in the No. 3 Mine

area .................................................................................................................... 69 Table 10.5 – 2013 technical parameters on the Parallel Zone .................................................... 71 Table 10.6 – Table of significant assay results for the 2013 drilling campaign on the Parallel

Zone ................................................................................................................... 74 Table 10.7 – 2010 technical parameters on the Sixteen Zone .................................................... 82 Table 10.8 – 2012 technical parameters on the No. 6 Vein ........................................................ 82 Table 13.1 – Sample Origin and Weight ..................................................................................... 94 Table 13.2 – Ball Mill Work Index ................................................................................................ 94

www.innovexplo.com


Table 13.3 – Head Assays Summary .......................................................................................... 95 Table 13.4 – High Grade Trace Mineral Search .......................................................................... 96 Table 13.5 – Gravity and Cyanide Leach – Conditions and Recoveries ..................................... 97 Table 13.6 – Summary of Flotation Gold Recoveries Obtained .................................................. 98 Table 13.7 – Flotation, Regrind and Cyanide Leach – Recoveries Obtained ............................. 98 Table 13.8 – Summary of Chemical Assays ............................................................................... 99 Table 13.9 – Flowsheet 1 – Recoveries Obtained .................................................................... 100 Table 13.10 – Flowsheet 2 – Recoveries Obtained .................................................................. 100 Table 13.11 – Flowsheet 3 – Recoveries Obtained .................................................................. 100 Table 13.12 – Flowsheet 4 – Recoveries Obtained .................................................................. 101 Table 14.1 – Total Indicated Resource Estimate by zone using a 3.00 g/t Au cut-off ............... 103 Table 14.2 – Total Inferred Resources Estimate by zone using a 3.00 g/t Au cut-off ............... 104 Table 14.3 – Key parameters for the 2013 and 2014 Mineral Resource Estimates by zone .... 104 Table 14.4 – Fortune Zone Mineral Resources ......................................................................... 114 Table 14.5 – Model variogram ................................................................................................... 122 Table 14.6 – Sensitivity of Resource to different cut-off values ................................................ 124 Table 14.7 – Mineral Resources in the Parallel Zone per vein .................................................. 125 Table 14.8 – Mineral Resources in the Parallel Zone (all veins) per elevation ......................... 126 Table 14.9 – Sensitivity of Indicated and Inferred categories to different cut-off values ........... 137 Table 14.10 – Mineral Resources of No. 4 Plug ........................................................................ 139 Table 14.11 – Sensitivity of the resource estimation to different cut-off values ........................ 147 Table 14.12 – Mineral Resources of the Triangle Zone ............................................................ 148 Table 14.13 – Cut-off sensitivity on resource estimation (Vein No. 6) ...................................... 159 Table 14.14 – Resources Estimates for Vein No. 6 calculated using progressive capping

above 40 g/t (cut-off grade of 3 g/t).................................................................. 160 Table 14.15 – Cut-off sensitivity on the resource estimation (Sixteen Zone) ............................ 170 Table 16.1 – Resources available to produce a preliminary mine plan with a 3.00 g/t Au cut-

off ..................................................................................................................... 173 Table 16.2 – Resources considered for the No. 7 veins to produce a preliminary mine plan ... 173 Table 16.3 – Resources used to prepare a preliminary mine plan ............................................ 174 Table 16.4 – Statistics for RQD data ......................................................................................... 174 Table 16.5 – Statistics for TCR data ......................................................................................... 175 Table 16.6 – Statistics for ISRM field hardness ........................................................................ 176 Table 16.7 – Approximate UCS ranges by ISRM field hardness (modified from ISRM, 1978) . 176 Table 16.8 – Statistics for fractures per metre of core .............................................................. 176 Table 16.9 – Typical ground support bolt length. ...................................................................... 177 Table 16.10 – Mine development quantities for the North Ramp .............................................. 185 Table 16.11 – Operational work place for the North Ramp ....................................................... 188 Table 16.12 – Mine development quantities for the South Ramp ............................................. 190 Table 16.13 – Operational work place for the South Ramp ...................................................... 192 Table 16.14 – Mining equipment for the Lamaque Project ........................................................ 193 Table 16.15 – Manpower requirements – Administration and Surface Services ....................... 194 Table 16.16 – Manpower requirements – Technical Services, Maintenance, Supervision

and Operations ................................................................................................ 195 Table 16.17 – Mine plan tonnage distribution ........................................................................... 196 Table 16.18 – Conceptual mining plan, yearly tonnage distribution for the North and South

ramps ............................................................................................................... 197 Table 16.19 – Airflow requirements ........................................................................................... 198 Table 17.1 – Potential plants for custom milling ........................................................................ 203

www.innovexplo.com


Table 20.1 – Cost Estimate for Restoration and Financial Guarantee - Lamaque Property Project, Integra Gold ........................................................................................ 220

Table 21.1 – Capital cost estimate ............................................................................................ 223 Table 21.2 – Surface infrastructure cost estimate ..................................................................... 224 Table 21.3 – Mining infrastructure cost estimate ....................................................................... 224 Table 21.4 – Development and capitalized operating costs ...................................................... 225 Table 21.5 – Owner’s costs ....................................................................................................... 226 Table 21.6 – Summary of total operating costs ......................................................................... 226 Table 21.7 – Mining costs ......................................................................................................... 227 Table 21.8 – Yearly energy cost (average for Years 1-4) ......................................................... 228 Table 21.9 – Yearly environment cost ....................................................................................... 229 Table 22.1 – Cash flow analysis summary ................................................................................ 231 Table 22.2 – Economic analysis for the Lamaque Project (figures in Canadian dollars) .......... 232 Table 22.3 – Sensitivity analysis of economical parameters, pre-tax NPV at 5% (millions $) ... 233 Table 22.4 – Sensitivity analysis of grade, pre-tax NPV at 5% (millions $) ............................... 234 Table 22.5 – Sensitivity analysis of economical parameters, pre-tax IRR ................................ 235 Table 22.6 – Sensitivity analysis on grade, pre-tax IRR ............................................................ 236 Table 23.1 – Compliant Reserve & Resources – Updated June 2009 ...................................... 237 Table 23.2 – Historical production statistics from Sigma mine .................................................. 237 Table 26.1 – Proposed work program and budget .................................................................... 250 LIST OF APPENDICES APPENDIX I – UNITS, CONVERSION FACTORS, ABBREVIATIONS .................................... 256

www.innovexplo.com


1. SUMMARY

At the request of Integra Gold Corp. (“Integra Gold”), InnovExplo has been retained to produce a Preliminary Economic Assessment (the “PEA”) and Technical Report (the “Report”) for the Lamaque Project (the “Project”) according to National Instrument 43-101 and Form 43 101F1. Integra Gold is a Canadian mineral exploration company trading publicly on the TSX Venture Exchange in Canada (TSXV: ICG). InnovExplo is an independent mining and exploration consulting firm based in Val-d’Or (Québec). The Report presents the results of the PEA for the Lamaque Project. The PEA is based on the Mineral Resource Estimate presented in an earlier report entitled “NI 43 101 Technical Report on the Lamaque Property” by Geologica Groupe-Conseil Inc. and GéoPointCom Inc., published in November 2013. The 2013 Mineral Resource Estimate is compliant with the Canadian Securities Administrators National Instrument 43 101 Standards of Disclosure for Mineral Projects (“NI 43 101”). In addition to Sylvie Poirier, Eng. (OIQ #112196) and Laurent Roy, Eng. (OIQ #109779) of InnovExplo, the other qualified persons responsible for the preparation of the Report are: Daniel Gaudreault, Eng. (OIQ #39834) of Geologica Groupe-Conseil Inc. (“Geologica”); Christian D’Amours, P.Geo. (OGQ #226) of GéoPointCom Inc. (“GéoPointCom”); Michel Garon, Eng. (OIQ #28151) of WSP Canada Inc. (“WSP); and Stephan Bergeron, P.Geo., M.Eng. (OGQ #787) of AMEC Environment & Infrastructure (“AMEC”). The Mineral Resource Estimate herein was prepared by Geologica and GéoPointCom. The environmental studies herein were completed by AMEC, and the review of the custom milling option for the Lamaque ore was completed by WSP. Property Description and Ownership

The Lamaque Property (the “Property”) is located in the Val-d’Or gold camp of central-northwestern Québec. It is lies within the Bourlamaque Township, partly within the municipality of Val-d’Or, about 550 km northwest of the city of Montréal. The coordinates for the approximate centre of the Property are 48°05’N latitude and 77°46’W longitude on NTS map sheet 32C/04 (National Topographic System quadrangle). The Property consists of parts of four contiguous mining concessions (“CM”) and 20 mining claims. The Property is registered 100% to Integra Gold for a total of 1,459.42 ha. None of the claims are within park or forest reserves that are restricted from exploration and mining. The PEA presented in this Report mainly concerns mining claims C002091, C002092, C002093, 3691171, 5275588 and 5275589. The current Lamaque Property includes much of what constituted the former mining property of Lamaque Gold Mines Ltd, which for parent company Teck Hughes Gold Mines Ltd (later Teck Corporation, now Teck Resources Ltd), produced a total of 4,554,167 ounces of gold between 1935 and 1985. The current Lamaque Property does not include the four main mines that provided the majority of the ore for the former property owners

www.innovexplo.com


Following the closure of the Lamaque mine, Teck Corporation (“Teck”) and Golden Pond Resources Ltd. (“Golden Pond”) formed the Teck-Golden Pond joint venture (“JV”) and Teck and Tundra Gold Mines Inc. (“Tundra”) formed the Teck-Tundra JV to explore a portion of the historical Lamaque mine property in 1985. The Golden Pond JV and some of the Tundra JV covered most of the ground now owned 100% by Integra Gold. In June 2003, Kalahari Resources Inc. (“Kalahari”) entered into an option agreement with Teck Cominco Ltd (“Teck”) to earn an interest of 50% to 53% (depending on the claims) in approximately 1,244 hectares called the Lamaque Project. On September 22, 2009, Kalahari entered into an option agreement with Alexandria Minerals Corp. to earn a 100% interest in the Roc d’Or East Extension property. Over a 3-year period, consideration was $25,000 cash (paid) and 500,000 shares (issued). There is a 2% NSR payable on the property, of which half (1%) may be purchased for $1,000,000. This claim group is adjacent to the Issuer’s 100% owned Roc d’Or East claims, which are now part of its Lamaque Property. In October 2009, Kalahari entered into separate agreements with Tundra and Golden Pond, joint venture parties at that time on the Lamaque Property, to purchase their interests in order to consolidate a 100% ownership of the Property and allow for more advanced exploration to be initiated. This meant issuing 9,593,128 shares to Tundra and 2,902,861 shares to Golden Pond, which asset could then be distributed in either cash or shares to shareholders, pro rata as determined by the boards of directors of Tundra and Golden Pond. This share consideration amounted to approximately one share of Kalahari for every dollar spent by Tundra and Golden Pond on the Property. Kalahari changed its name to Integra Gold Corp. in December 2010 and now owns 100% of the Property. Integra Gold also acquired 100% of Teck’s interest in the adjacent Roc d’Or East and Roc d’Or West claims. There is a 2% NSR payable to Teck on the Property, of which half (1%) may be purchased for $2,000,000 at any time within one year of commercial production. In December 2010, Integra Gold acquired an option to earn a 100% interest in the Bourlamaque property in Bourlamaque Township, Québec, adjacent to the Lamaque Property. Consideration for the property acquisition consisted of $3,500 cash (paid) and 10,000 shares (issued). Integra Gold purchased the entire NSR for $5,000 on April 30, 2013 (no outstanding NSR). An agreement is in place between Teck and Integra Gold concerning the tailings of the Old Lamaque mine. Integra Gold must provide a reclamation plan for the historical tailings and pay for the required work in order for Teck to be potentially liberated from its obligation, pending approval from the (MRN).

www.innovexplo.com


Geology and Mineralization

The Property is located in the Malartic Composite Block, or Malartic Group, in the southeastern section of the Abitibi Greenstone Belt, a 750 km long by 250 km wide Island Arc volcanic complex, of the Archean Superior Province, Canadian Shield. The Grenville Front is located about 50 km to the southeast and the Cadillac-Larder Lake Fault Zone, deformation zone or "Break" is located about one km south of the southernmost edge of the Property. The Cadillac-Larder Lake Break runs for about 250 km and is the fundamental control for gold mineralization in the Val-d’Or, Malartic, Cadillac and Kirkland Lake camps. The Property is principally underlain by volcanic flows and volcaniclastics of the Val-d’Or Formation (2705-2703 Ma), intruded by a variety of intermediate to mafic plugs, dykes and sills. The northern part of the Property is underlain by lithologies comprising the lower part of the Val-d’Or Formation, while the mid to southern part of the Property is underlain by units of the upper Val-d’Or Formation. The extreme southern section of the Property, in the vicinity of the airport, is underlain by mafic volcanics of the Héva Formation. There are twelve (12) zones of gold mineralization, including six (6) gold deposits with known resources on the Lamaque Property. They consist of various simple vein structures, vein complexes and stockwork zones. They are: • Parallel Zone (including historical No. 10, No.5 and No.7 Veins); • Fortune Zone (previously known as Forestel); • No. 4 Plug; • No. 5 Plug (including No. 35 Veins); • No. 3 Mine (including No. 1 and 2 Veins); • Triangle Zone; • South Triangle Zone; • Mylamaque Zone; • No.4 Vein; • No. 6 Vein; • Sixteen Zone; and • Sigma Zone. The six (6) presently known gold deposits are grouped into the North, South and West clusters. The North cluster consists of the Parallel and Fortune Zones but also the exploration targets known as the No. 5 Plug and No. 3 Mine. The South cluster consists of the No. 4 Plug, Triangle, and South Triangle zones. The West cluster consists of the Sixteen Zone and No. 6 Vein. For clarity, the West Cluster zones are not included in the PEA. Mineral Resource Estimate

The 2013 Mineral Resource Estimates for the Fortune, Parallel, No. 4 Plug and Triangle zones of the Lamaque Project, presented herein, were completed by GéoPointCom, using all available results as per the effective date of each zone. The main objective was to publish revised mineral resource estimates for the above mentioned zones. The mineral resources presented herein are not mineral reserves as they have no demonstrable economic viability. The result is a Mineral Resource

www.innovexplo.com


Estimate with Indicated and Inferred resources for each of the four mineralized zones, modelled for underground mining. The resource estimates for the abovementioned zones are part of a NI 43-101 Technical Report prepared and supervised by Geologica Inc. and filed on SEDAR in November 2013. The effective dates of the Mineral Resource Estimates vary from zone to zone but the common publication date by news release is September 25, 2013. All details for these resource calculations are presented in Tables 14.1, 14.2 and 14.3. The November 2013 mineral resources report for the four zones were included in the current PEA. The 2014 Mineral Resource Estimates for the No. 6 Vein and Sixteen Zone, presented herein, were completed by GéoPointCom, using all available results as per the effective date of each zone. The publication date by press release is January 28, 2014. The result is Mineral Resource Estimates with Indicated and Inferred Resources (see Tables 14.1, 14.2 and 14.3). The mineral resources are not mineral reserves as they have no demonstrable economic viability. These mineral resources were not included in the current PEA. Total Indicated Resource Estimate by zone using a 3.00 g/t Au cut-off (Table 14.1)

Gold Deposit Name Metric Tonnes Grade (g/t Au) Ounces

No. 4 Plug 1,325,100 5.6 237,450

Fortune Zone 125,500 5.8 23,600

Parallel Zone 793,900 8.2 209,570

Triangle Zone 599,700 9.9 190,670

No. 6 Vein 389,400 6.4 79,550

Sixteen Zone 91,700 5.2 15,440

Total Indicated 3,325,300 7.1 756,280

Total Inferred Resources Estimate by zone using a 3.00 g/t Au cut-off (Table 14.2)

Gold Deposit Name

Metric Tonnes Grade (g/t Au) Ounces

No. 4 Plug 0 0.0 0

Fortune Zone 252,300 5.6 45,220

Parallel Zone 153,400 17.5 86,050

Triangle Zone 332,300 12.9 137,600

No. 6 Vein 111,600 6.9 24,590

Sixteen Zone 1,800 4.2 250

Total Inferred 851,400 10.8 293,710

www.innovexplo.com


Key parameters for the 2013 and 2014 Mineral Resource Estimates by zone (Table 14.3)

Zone Estimator Cell dimensions

Min. search radius

Max. search radius

Min. no. samples

Max. no. samples

Capping grade

Cut Off grade

Min. true thickness

Compo-site

length

Effective date

Fortune* Ordinary Kriging

10X10X15m 25m 90m 3 10 None 3 2m 0.5m 2012 11/08

Parallel Ordinary Kriging

5X5X5m 50m 50m 3 15 100

[progressive] 3 2m 0.6m

2012 05/24

No. 4 Plug

Ordinary Kriging

10X10X10m 35X50X

16m 60X60X

16m 3 10 300 3 ** 1.0m

2013 03/19

Triangle Inverse squared distance

5X5X5m 25m 50m 2 10 80

[progressive] 3 2m 1m

2013 04/24

No. 6 Vein

Ordinary Kriging

10X10X10m 50X50X

50 100X100

X100 4 8

40 [progressive]

3 2m 1.0m 2012 08/17

Sixteen Ordinary Kriging

10X10X10m 15X15X

15 60X60X

60 5 10 35 3 ** 0.7m

2013 11/18

* True thickness, ore grade and dilution grade were estimated for all cells and recombined at the end.

** Not constrained to vein. The selection is based on "High Probability Ore" within the dioritic intrusive.

Mining

Mineralization at the Lamaque Project would be accessed via two separate ramps, or declines, located in the Parallel Zone to the north (the “North Ramp”) and in the Triangle Zone to the south (the “South Ramp”), approximately 2 km apart. Material would then be transported to an off-site mill for toll processing, thereby eliminating the need for the construction and permitting of a new mill and tailings facilities. The mining plan for the Lamaque Project calls for a combination of conventional and mechanized mining. Two mining methods are proposed based on the vein geometry of the four deposits: long-hole and room and pillar. The approach in this study has been to force the application of long-hole mining where applicable. Waste material generated from drift development will be used to backfill part of the long-hole open stopes. The mining methods are proposed to accommodate the geometry of the mineralization. The mining methods were selected according to vein geometry and common practices for comparable mining operations in the region, an area with an extensive history of underground mining. For mineralized zones dipping less than 45°, a room and pillar mining method is proposed, and sublevel long-hole retreat is proposed for zones dipping more than 45°. An administration and mine service hub would be located on Highway 117, part of the Trans-Canada highway system. The service hub would be served by a 25 KV power line, natural gas and municipal services. There will be two production centers, each with a ramp to access resources (the “North Ramp” and “South Ramp”), and both will include basic surface infrastructure.

www.innovexplo.com


The overall life of the Project is expected to be approximately 6.25 years, including a 2-year preproduction period, followed by 4.25 years of production. Once mining operations are completed, 1 year will be required to complete the mine closure work. The expected production rate will start at an average of 690 tpd from the North sector in the third quarter of preproduction Year 0, and will slowly ramp up to an average of 750 tpd in Year 1. The South sector production will start in last quarter of preproduction Year 0, to bring the production level to 790 tpd. The preliminary conceptual mine plan extends over a period of 4.25 years, including a 2-year preproduction period. The average production level for North and South Ramp is 1,480 tpd (312 days/year). The mine will operate seven (7) days a week, excluding night shifts on Fridays and Saturdays. This schedule is equivalent to 312 days per year of operation. Processing

At this early stage of the Project, several options can be considered for the processing plant. However, these assumptions would have to be re-examined and optimized during a pre-feasibility study. One of the main underlying assumptions of this PEA is that the mineralized material would be transported and processed off-site. The metallurgical testwork achieved to-date, has demonstrated the amenability of the Lamaque mineralized material to the gravity, leaching and flotation processes, although further work is required to better determine the specific flowsheet that will optimize the metallurgical performance. For the purpose of the preliminary economic assessment (PEA) of the project, WSP recommends to use recovery numbers that were obtained from retention times of 48 hours because the plants that have been looked at for the various milling options, might offer retention times of about that order of magnitude or likely less. Based on the results obtained from tests KM4025-05 to 08 (ALS Metallurgy Kamloops, Roulston D., Johnston H., Metallurgical Testwork on the Lamaque Deposit, January 16, 2014) and after adjusting leaching recoveries for a 48 hrs retention time, using the leach kinetic curves, the following recoveries could be used if the only process used is cyanidation:

Fortune: 93.8% (represents 98.1% of recovery obtained with a retention time of 96 hrs)

Parallel: 95.5% (represents 98.3% of recovery obtained with a retention time of 96 hrs)

Triangle: 89.7% (represents 96.5% of recovery obtained with a retention time of 96 hrs)

No. 4Plug: 82.0% (represents 98.5% of recovery obtained with a retention time of 96 hrs)

The percentage of the 96 hrs recovery for the Triangle Zone was lower than the one obtained with the others zones after 48 hrs (96.5% vs. >98%). To validate that, the

www.innovexplo.com


ALS Kamloops laboratory redid a cyanidation test with the same Triangle Zone composite and similar test conditions (KM4232). The results obtained, indicate a gold recovery of 89.5 % after a leaching time of 48 hrs, which represents 98.5% of the gold recovered after 96 hrs. Always based on the same report, if the process involves gravity separation with cyanidation of the gravity tails, the results obtained in KM4015-01 to 04 (gravity testwork) and KM4025-25 to 28 (cyanidation of gravity tails), were used after calculating leaching recovery for a 48 hrs retention time, using the factors above, derived from the leach kinetic curves of tests 05 to 08, to get the following recoveries that could be used for the PEA:

Fortune Leach: 68.2% Gravity: 26.8% Total: 95.0% Parallel Leach: 49.1% Gravity: 47.6% Total: 96.7% Triangle* Leach: 73.5% Gravity: 17.6% Total: 91.1% No. 4 Plug Leach: 72.4% Gravity: 13.7% Total: 86.1%

*The average of both tests (Reports KM4025 and 4232) was used to calculate the recovery factor for the Triangle Zone (89.6% recovery after 48 hrs which is 97.5% of the 91.9% recovery after 96 hrs). However the results obtained to-date, from the series of laboratory tests performed by ALS Metallurgy Kamloops, are preliminary and further testwork will be needed to better study gravity concentration, determine the fineness of grind and reagents consumption that will enhance recoveries, and define the best flowsheet that could be used to optimize metallurgical performance as well as economics. A bulk sample run in the plant selected could be extremely useful to adjust the process developed in the lab to a plant scale process.

Environment

Part of the Property is covered by tailings from the previous mining operations at the Lamaque Mine from 1935 to 1985 (Figure 4.2). The tailings are largely confined by a tailings dam. The dam and tailings are stable and fairly dry and support grasses and herbaceous growth. Teck (and its predecessors) previously planted conifers on various parts of the tailings and these trees appear to be doing well. The tailings do not appear to have the potential to generate acid mine drainage. Teck still retains legal responsibility for the tailings. Depending on season and weather, parts of the tailing are wet, with standing water and intermittent creeks. Excess water exits the tailing area through runoff over two open spillways. No environmental monitoring is required. In order to conduct exploration work, Integra Gold must respect all laws relative to exploration, and request all appropriate forestry intervention permits from the MRN for all drilling and trenching related activities. To drill on the historical Lamaque tailings pond, a reclamation plan must be submitted to the MRN. The federal and provincial environmental acts and regulations were reviewed. An Environmental Impact Assessment (EIA) will be required under the Canadian

www.innovexplo.com


Environmental Assessment Act (2012) because the proposed production will be greater than 600 tpd. A provincial EIA accompanied by a Bureau d’audiences publiques sur l’environnement (“BAPE”) public hearing may also be required, if production is 2,000 tpd or more. Other specific regulatory approvals will also be required, such as provincial Certificates of Authorization (“CofA”) for the planned water treatment facility, sinking of ramps and shafts, and others. An environmental baseline study of the Site was undertaken on behalf of Integra Gold by AMEC Environment & Infrastructure (“AMEC”) during 2013. The study reviewed available information across a number of disciplines, including geology and soils, hydrogeology, hydrology, air quality and noise, flora and fauna, socio-economic setting, and archaeology. There will be no tailings impoundment area on the Site, as the plan is to transport the ore off-site to an existing processing facility where tailings will be managed. Waste rock will be disposed of on-site in a dedicated storage pile, with control and capture of run-off from the waste rock to treat as may be indicated prior to discharge to the environment. Mine water will require a wastewater treatment system prior to discharge to the receiving environment, as indicated by the results of the hydrogeological investigation described above. Finally, the overall Site drainage will require a management plan in order to ensure that there is no contamination of water flows resulting from contact with the on-Site transportation network including vehicle fueling stations, waste rock disposal piles, on-Site ore stockpiles, and other work areas. At the operational phase of the Project, the outfalls of wastewater treatment systems and any other water discharged to the receiving environment will need to be monitored in keeping with the requirements of the federal Metal Mining Effluent Regulations, and the provincial Directive 019, to ensure that concentrations of specified contaminants are in compliance with acceptable levels. In addition, Environmental Effects Monitoring (“EEM”) will need to be conducted, again in keeping with the requirements of the MMERs. These regulations are currently under review (see point (c) below). A targeted water quality monitoring program will be developed as an integral part of the mine closure plan and the site reclamation plan. It will be consistent with the requirements of the MMERs and Directive 019. This program will be intended to demonstrate the effectiveness of the implemented closure and reclamation measures in preventing any contamination from the mine and disposal area workings from moving into the surrounding environment. The program will concentrate on the receiving waters of the principle stream crossing the Site, and will encompass both upstream (control) and downstream sampling stations of those areas which conceivably could contribute some form of contaminant, should the closure and reclamation measures be to some degree lacking in effectiveness. The program will be implemented over a minimum period of five years following reclamation. A plan for rehabilitation and restoration will be submitted to the MRN in order to obtain authorization for any of the exploration or development work mentioned above. This plan must be prepared according to the “Guidelines for preparing a mining site rehabilitation plan and general mining site rehabilitation requirements”

www.innovexplo.com


(1997). The plan must be reviewed every 5 years, but significant changes to the Project might also trigger the need for update, at the request of MRN. The entire cost of the proposed plan must be guaranteed by Integra Gold. An investigation of socio-economic information was carried out as part of the environmental baseline study referenced above. This investigation relied solely upon information available in the public domain and did not collect any primary information. The design of the Project is consistent with the zoning bylaws of Val d’Or. No protected areas or areas of archaeological interest are affected by the Project. A stakeholder and First Nations engagement and consultation plan has been prepared and is being implemented by Integra Gold This will eventually include consultations regarding the closure plan. Stakeholder issues raised to date include noise, vibration, traffic circulation, environmental protection and visual impact. Capital and Operating Cost

Capital costs

The preproduction costs are estimated at $69.2M, net of production revenue received during the second year of the preproduction period ($37.4M) (Table 21.1). Preproduction capital costs include surface infrastructure (site preparation, roads, power lines, and water lines), installation of modular buildings for offices and garages (mechanical and electrical shops, stockroom), mining infrastructure at the North and South sites, mobile equipment, development and capitalized operating costs, owner costs (closure costs in line with required financial guarantees, company staff costs, and indirect costs) as summarized in the following tables. Preproduction capital costs are minimal given that there is no need to build processing and tailings facilities, and that mineralization is spatially close to surface. 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.

Capital cost estimate (Table 21.1) Description Preproduction Sustaining Total *

Surface infrastructure 12.9 M$ 4.7 M$ 17.7 M$ Mining infrastructure 6.9 M$ 3.1 M$ 10.1 M$ Mobile equipment 14.8 M$ 17.3 M$ 32.2 M$ Develop. & capitalized operating costs 55.6 M$ 39.0 M$ 94.6 M$ Owner’s costs 16.3 M$ 2.6 M$ 18.9 M$ Offsetting capitalized revenue* (37.4) M$ (37.4) M$

Total 69.2 M$ 66.8 M$ 136.0 M$

www.innovexplo.com


Operating costs

Operating costs are summarized below for the production period (Table 21.6). Given that this PEA presents a toll milling scenario and that Integra Gold has the ability to process mineralized material recovered during the preproduction and development stage, revenue generated from these ounces has been included in forecasted cash flows. A total of approximately 28,000 ounces are anticipated to be produced during Year 2 of the preproduction phase.

Summary of total operating costs (Table 21.6)

Financial Analysis

An after-tax model was developed for the Lamaque Project. All costs are in 2013 Canadian dollars with no allowance for inflation or escalation. The Lamaque Project is subject to the following taxes:

- Québec mining tax rate of 16% (2014 rate); - Income tax rate of 26.9% (federal and provincial). The Lamaque Property is subject to a royalty in favor of Teck, equal to 2% of NSR and the buyout of 1% of the NSR for $2M is included in the economic model. 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 (Table 22.1):

• An average gold price of $US1,275/oz and an exchange rate of 1.05CAD/1USD (lower than 3-yr trailing average as of February 28, 2014).

• Milling recovery: o Parallel Zone: 97% o Fortune zone: 95%

Description Total cost Unit cost (Years 3-7) ($/t) ($/oz)

Definition drilling 11.7 M$ 6.00 24.58 Stope development 34.9 M$ 17.86 73.16 Mining cost 115.3 M$ 58.94 241.51 Integra Gold Staff 38.8 M$ 19.82 81.21 Energy cost 15.0 M$ 7.64 31.31 Milling and transportation 89.4 M$ 45.69 187.21 Environment 4.1 M$ 2.08 8.54

Total 309.3 M$ 158.04 647.53

www.innovexplo.com


o Triangle zone: 90% o No. 4 Plug zone: 86% o Resulting in an average recovery of 92.2% for the entire Project.

• Royal Mint fees of $3/oz. • Royalty of 2% of Net Profit for preproduction, and a payment of $2M and a

royalty payment of 1% of Net Profit for the production period. • 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. The financial analysis indicates a payback period of 1.8 years. The after-tax Net Present Value (NPV) of the project is estimated at $88.5M with a discount rate of 5%. The after-tax Internal Rate of Return (IRR) is evaluated at 38%. The main results are summarized in the following table:

Cash flow analysis summary (Table 22.1)

Parameters Results

Gold price (US$/ounce): 1,275

Foreign exchange rate (CAD/USD): 1.05

Gold price (CA$/ounce): 1,339

Average Annual Gold Production (ounces/year): 112,400

Peak Annual Gold Production (ounces) 143,300

Preproduction Capital Costs (CA$) 69.2 M

LOM Sustaining Capital (CA$) 66.8 M

Preproduction Period (years) 2

Mine Life (years) 4.25

Cash Cost per Gold Ounce (CA$/oz) 665

Cash Costs and Sustaining Cost per Gold Ounce (CA$/oz) 805

PRE-TAX

Life of Mine NPV at 5% Discount Rate (CA$) 146.0 M

Internal Rate of Return (IRR) 51%

Payback period (years) 1.5

AFTER-TAX


IRR after-tax (%) 38%

Payback Period (years) 1.8

www.innovexplo.com


Opportunities & Risks

Opportunities to improve the Lamaque Project economics include the following:

• The Issuer is developing a scenario internally which would involve delaying the development of the South Ramp by 12 to 18 months in order to reduce up-front capital cost requirements and utilize cash flow from the North Ramp to fund development of the South Ramp.

• Acquisition of a mill instead of toll milling would likely reduce LOM operating costs and allow Integra Gold greater security in meeting its future milling requirements.

• The potential to utilize contract mining in order to reduce up-front capital requirements.

• Production outlined in the PEA is limited to a vertical depth of 620 m at the Triangle Zone. A 2013 drill program intersected multiple high-grade zones below this level, to vertical depths of up to 1,000 m. The Triangle Zone also remains open to the south, east and west.

• Drilling at the Triangle Zone intersected 13.89 g/t Au over 7.0 m, approximately 175 m down-dip from the Triangle Zone resource estimate. The ground in between the Triangle Zone has been subsequently tested in 2014 with assay pending.

• The PEA is based on a mineral resource database cut-off date of April 24, 2013 and does not include the subsequent drilling (either infill or expansion) of approximately 39,235 m that was completed in late February 2014. Another phase of drilling (4 rigs) is currently underway, and this is also not included in the PEA.

• The PEA does not include resources from the No. 6 Vein or the Sixteen Zone.

• Significant mineralization has been identified at the No. 3 Mine and the No. 5 Plug targets. Integra Gold expects to have resource estimates completed for those zones in the second half of 2014. Should a resource be defined at these targets, they could be potentially mined from the North Ramp infrastructure.

• Recent metallurgical testwork indicates a potential to further improve gold recoveries.

Risks requiring mitigation strategies include:

• The Issuer’s future financial success depends on the ability to raise additional capital from the issue of shares or the discovery of property which could be economically justifiable to develop. Such development could take years to complete and resulting income, if any, is difficult to determine. The sales value of any mineralization potentially discovered by the Issuer is largely dependent upon factors beyond the Issuer’s control, such as the market value of the products produced.

• The resource exploration industry is an inherently risky business with significant capital expenditures and volatile metals markets. The marketability of any minerals discovered may be affected by numerous factors that are beyond the Issuer’s control and which cannot be predicted, such as market fluctuations, mineral markets and processing equipment, and changes to

www.innovexplo.com


government regulations, including those relating to royalties, allowable production, importing and exporting of minerals, and environmental protection.

• This industry is intensely competitive and there is no guarantee that, even if commercial quantities are discovered, a profitable market will exist for their sale. The Issuer competes with other junior exploration companies for the acquisition of mineral claims as well for the engagement of qualified contractors. Metal prices have fluctuated widely in recent years, and they are determined in international markets over which the Issuer has no influence.

• Exploration and development on the Issuer’s Property are affected by government regulations relating to such matters as environmental protection, health, safety and labour, mining law reform, restrictions on production, price control, tax increases, maintenance of claims, and tenure. There is no assurance that future changes in such regulations would not result in additional expenses and capital expenditures, decreasing availability of capital, increased competition, title risks, and delays in operations.

• Management of construction/engineering and procurement schedules, costs, and cost containment.

• Operating risks related to recruitment and training and performance of the underground workforce, specifically room-and-pillar miners.

• Permitting risks. • Crown pillar thickness and stability evaluation through geo-mechanics

characterization and stability analysis. • Possibilities that the population does not accept the mining project.

Recommendations

The results from this prefeasibility study demonstrate that the Lamaque Project is technically and economically viable and InnovExplo recommends that Integra Gold continue to advance the project toward prefeasibility. InnovExplo recommends proceeding with the following steps in the continued development of the Lamaque Project.

1. Continue exploration and definition drilling at the Parallel, Triangle and Fortune zones in 2014 in order to upgrade as many resources as possible from the Inferred resource category to the Indicated resource category, while continuing to increase the resource base laterally and at depth.

2. Estimates for all zones included in the PEA using all new information generated since the latest database cut-off and evaluate their impact on Project economics (since April 24, 2013).

3. Complete exploration drilling and perform resource estimations on two of the Project’s advanced exploration targets, the No. 3 Mine and No. 5 Plug, in order to integrate these areas into the future economic evaluations for the Project.

4. Commence a prefeasibility study that includes:

• Hydrogeology study; • Rock mass characterization and stope design; • Crown pillar stability analysis; • Revised mining plan using new resources;

www.innovexplo.com


• Trade-off analysis; • Re-scheduling development of the 2 ramps to limit capital requirement; • Energy alternatives for underground air heating; • Mineralized material and waste handling alternatives; and • Access, possibly via shaft sinking, to the deeper part of the Triangle and

No. 4 Plug zones. • Finalize the connecting scenario to the Hydro-Québec grid; • Engineering for surface installation, electricity and mechanics installations; • Engineering for water treatment and management facilities; and • Updated economic evaluation of capital expenditures and operating costs.

5. Initiate and complete the permitting process for an underground exploration program. A complete prefeasibility study will likely require underground exploration, meaning a significant portion of Project permitting will be completed during the prefeasibility stage:

• Apply for Certificate of Authorization under Québec’s jurisdiction; • Apply for Project Description under Canada’s jurisdiction; • Mineralization and waste characterization; • Hydrology study; • Noise study; and • Biology study.

6. Conduct a fourth phase metallurgical study in order to further improve gold recoveries for the Triangle and No. 4 Plug zones.

7. Complete a formal information and consultation process in order to promote social acceptability of the Lamaque Project and the plans for its development.

To advance the project, InnovExplo estimates an exploration budget of approximately $6.9M is required as presented in Table 26.1.

Proposed work program and budget (Table 26.1) Item Cost

1. Exploration and definition drilling at the Parallel, Triangle and Fortune zones

3,500,000$

2. Update resource estimates 300,000$3. Complete exploration drilling and perform

resource estimations on two of the Lamaque Project’s advanced exploration targets, the No. 3 Mine and No. 5 Plug.

1,000,000$

4. Prefeasibility study 1,350,000$

5. Initiate and complete the permitting process for an underground exploration program.

350,000$

6. Conduct a fourth phase metallurgical study 150,000$7. Complete a formal information and

consultation process 250,000$

Total 6,900,000$

www.innovexplo.com


2. INTRODUCTION

At the request of Integra Gold Corp. (“Integra Gold”), InnovExplo has been retained to produce a Preliminary Economic Assessment (the “PEA”) and Technical Report (the “Report”) for the Lamaque Project (the “Project”) according to National Instrument 43-101 and Form 43-101F1. Integra Gold is a Canadian mineral exploration company trading publicly on the TSX Venture Exchange in Canada (TSXV: ICG). InnovExplo is an independent mining and exploration consulting firm based in Val-d’Or (Québec). The Report presents the results of the PEA for the Lamaque Project. The PEA is based on the Mineral Resource Estimate presented in an earlier report entitled “NI 43-101 Technical Report on the Lamaque Property” by Geologica Groupe-Conseil Inc. and GéoPointCom Inc., published in November 2013. The 2013 Mineral Resource Estimate is compliant with the Canadian Securities Administrators National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”).

2.1. Terms of Reference and Scope of Work

The Issuer requested a Preliminary Economic Assessment. The global objectives of the PEA are to:

Examine the potential economic viability of mining the Lamaque deposit; Propose a strategy and preliminary timetable to develop the Lamaque

Project; The PEA herein evaluates and/or provides:

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

rock mass conditions of the Lamaque deposit; The basic design for most of the required facilities and infrastructure to

access, develop and mine the mineralized zones; An estimate of the capital and operating costs; A preliminary cash flow model; An estimate of the cost and timeframe for the preproduction and production

periods; An analysis of the financial aspects of the Project; An estimate of resources potentially amenable to mining; Recommendations for additional work to advance the Project to the next

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

2.2. Principal Sources of Information

InnovExplo’s assessment of the Lamaque Project was based on published material as well as the data, professional opinions and unpublished material submitted by Integra Gold or requested by InnovExplo to complete the study. The following specialists provided information for various portions of the study:

www.innovexplo.com


Frédéric Bédard, Horizon SF, provided the budgetary quote for tree clearing (trunks >10 cm in diameter);

Joel Lavigne, independent contractor, provided the budgetary quote for tree clearing (wood <10 cm in diameter);

Vick Morin, Aménagement V.M. Inc., provided the budgetary quote for fencing;

François Goulet, Dynamitage Castonguay Ltée, provided the budgetary quote for drilling related to fence installation;

Derek Bertrand, project manager at Services Miniers Nord-Ouest, provided the budgetary quote for the roadwork and access ramps;

Michel Leduc, Atlas Copco, provided the budgetary quote for underground mobile equipment;

John Chomyshyn, account manager at Maclean Engineering & Marketing Co. Ltd, provided the budgetary quote for underground mobile equipment;

Guy Marseille and Stéphane Bond, mining and sales representatives at Kubota Pro-AbÉquipements (2003) Inc, provided the budgetary quote for underground service mobile equipment and the MegaDome;

Denis Lemieux, CMAC – Thyssen Mining Manufacturer Inc., provided the budgetary quote for long-hole drilling;

Jean-Yves Poitras, Corporation de Développement Industriel de Val-d'Or (CDIVD), provided the budgetary quote for lot acquisition;

Mario Massé, executive vice-president at VCC Massénor, provided the budgetary quote for the underground mobile equipment garage;

Maxime Provencher, sales representative at Pétroles J.C. Trudel, provided the budgetary quote for the fuel station;

Louis Martin, Les Industries Desjardins Ltée, provided the budgetary quote for used oil storage;

Benoit Décary, Conteneurs Experts S.D., provided the budgetary quote for the storage container;

François Lambert, Dyno Nobel, provided the budgetary quote for the surface powder-magazine;

Stéphane Castonguay, sales person at Protexplo, provided the budgetary quote for the heating system for the surface powder-magazine;

Jacques Grenier, sales director at Modulabec, provided the budgetary quote for the office, dry and gate house;

Patrick Bédard, manager at Stavibel, provided the budgetary quote for networking and communication;

Marc Turcotte, manager at ASDR Environnement, provided an assessment of the water treatment installation;

Jacques Bédard, Entreprises Larry Inc., provided the budgetary quote for the compressor;

Alain Fluet, project and operations director at Agrégat R-N Inc., provided the budgetary quote for the ramp portal;

Alain Grenier, account manager at Boart Longyear, provided the budgetary quote for drilling equipment;

Réal Bérubé, Location Dumco, provided the budgetary quote for the surface pipeline;

Ghislain Grondin, Équipements Miniers 2000, provided the budgetary quote for underground equipment;

www.innovexplo.com


Dominic Mailhot, representative at Pompaction, provided the budgetary quote for the grouting pump;

Jeremy Wilgos, Wilson Mining Products Inc., provided the budgetary quote for the diaphragm pump;

François Tremblay and Yves Laquerre, Hewitt Equipment Ltd, supplied the mobile fleet requirements and budgetary costs;

Emile P. Molgat, Business Support at Orica, provided the budgetary quote for explosive products and associated services;

Maxim Dupras, Xylem, provided the budgetary quote for the mine dewatering system;

François Chabot, manager of engineering operations at Integra Gold, provided the South Ramp planning, with the contribution of Alain Côté, LBP Mining Services, for technical drawing;

Richard Knight, KN Equipment, provided the budgetary quote for the mine ventilators and main fans installation;

Martin Vigneault, director of operations, Propane Nord-Ouest Inc., provided the budgetary quote for the surface propane installation:

Denis Cloutier, Eclipse Combustion Canada Inc., provided the budgetary quote for heating the underground ventilation system; and

Yoland Nolet, project leader, Meglab Inc., provided the budgetary quote for surface and underground electricity distribution.

In addition, InnovExplo estimated costs using quotes from contractors and suppliers, as well as data provided in the 2013 volume of Mining Cost Service (with a subscription to Cost Data Update Services) published by CostMine, a division of InfoMine USA Inc. InnovExplo conducted a review and appraisal of the information used to prepare the Report and to formulate its conclusions and recommendations, and believes that such information is valid and appropriate considering the status of the Project and the purpose for which the Report is prepared. The authors have fully researched and documented the conclusions and recommendations made in the Report.

2.3. Qualified Persons and Inspection on the Property

In addition to Sylvie Poirier, Eng. (OIQ #112196) and Laurent Roy, Eng. (OIQ #109779) of InnovExplo, the other qualified persons responsible for the preparation of the Report are: Daniel Gaudreault, Eng. (OIQ #39834) of Geologica Groupe-Conseil Inc. (“Geologica”); Christian D’Amours, P.Geo. (OGQ #226) of GéoPointCom Inc. (“GéoPointCom”); Michel Garon, Eng. of WSP Canada Inc. (“WSP); and Stephan Bergeron, P.Geo., M.Eng. (OGQ #787) of AMEC Environment & Infrastructure (“AMEC”). The Mineral Resource Estimate herein was prepared by Geologica and GéoPointCom. The environmental studies herein were completed by AMEC, and the review of the custom milling option for the Lamaque ore was completed by WSP. Technical support was provided by InnovExplo employees Marie-Claire Dagenais, Jr Eng., and Serge Morin. In addition, Bruno Turcotte, P.Geo., of InnovExplo, validated the compliance of the Report with NI 43-101 and Form 43-101F. InnovExplo wishes to acknowledge the diligent and professional technical assistance of Geologica and

www.innovexplo.com


GéoPointCom for this gold project, particularly to Josette Boucher, Jean St-Laurent, DAOs and Infograpphists, Laura Guillaume and Benjamin Blaise, Exploration Geologists, Drill Data Compilation and Validation, Daniel Gaudreault, P. Eng, Geo. and Alain- Jean Beauregard, P. Geo., Drill Program Planning and Supervision, Mineralized Zones Interpretations and Proof reading of this report, Christian D’Amour PGeo., resources estimation. Alain Coté of LBP Mining did the mine plan drawing for the Triangle and No. 4 Plug. Also, Mr. Hervé Thiboutot, P.Eng., P. Geo., Executive and Corporate Vice-President of Integra Gold Corp. played a key role by orienting the Lamaque Project advancing it towards its development and mining stages with the precious collaboration of Mr. François Chabot. P.Eng, P.Geo., Manager of Operations. The list below presents the sections for which each Qualified Person was responsible:

Laurent Roy: author of sections 2, 3, 15, 16, 18, 19, 21, 22, 24; co-authors of sections 1, and 25 to 27.

Sylvie Poirier: responsible for supervising the preparation of Sections 1,2,3,15,16, 18, 19, 21, 22, 24, 25, 26, 27 and co-author of Sections 1 and 25 to 27.

Daniel Gaudreault: author of sections 4, 5, 6, 7, 8, 9, 10, 11, 12, 23; co-author of sections 1 and 25 to 27.

Christian D’Amours: author of section 14; co-author of sections 1 and 25 to 27.

Stephan Bergeron: author of section 20; co-author of sections 1 and 25 to 27. Michel Garon: author of sections 13, 17; co-author of sections 1 and 25 to 27.

For the current PEA, Laurent Roy and Marie-Claire Dagenais, of InnovExplo, visited the Lamaque Project site on July 9, 2013, accompanied by François Chabot of Integra Gold.

2.4. Units and Currencies

All currency amounts are stated in Canadian Dollars ($, C$, CAD) or US dollars (US$, USD). Quantities are stated in metric units, as per standard Canadian and international practice, including metric tons (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 ton (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 the Report is provided in Appendix I.

www.innovexplo.com


3. RELIANCE ON OTHER EXPERTS

The authors, Qualified and Independent Persons as defined by Regulation 43-101, were contracted by the Issuer to study technical documentation relevant to the Report, to perform a PEA study on the Lamaque deposit, and to recommend a work program if warranted. The authors relied on reports and opinions as follows for information that is not within the authors’ fields of expertise:

Information about the mining titles and option agreements was supplied by Geologica. InnovExplo is not qualified to express any legal opinion with respect to property titles or current ownership and possible litigation.

Golder Associates Ltd (“Golder”) was retained by Integra Gold to provide professional services with respect to the Lamaque Property. The main objective was to determine the level of study required for the Project with respect to geomechanics, crown pillar, mining recovery rate and groundwater/geochemistry, of the Triangle Zone, the No. 4 Plug, and the Parallel and Fortune zones. The Golder report was used to elaborate part of section 16.3 (“Preliminary Geotechnical Assessment”).

Meglab Inc. provided the preliminary estimate for the electrical distribution requirements and costs.

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

Venetia Bodycomb of Vee Geoservices provided the linguistic editing on an earlier draft version of the Report.

The authors believe the information used to prepare the Report and 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 in the Report are in accordance with NI 43-101 and CIM technical standards.

www.innovexplo.com


4. PROPERTY DESCRIPTION AND LOCATION

4.1. Location and Claims

The Lamaque Property (the “Property”) is located in the Val-d’Or gold camp of central-northwestern Québec. It is lies within the Bourlamaque Township, partly within the municipality of Val-d’Or, about 550 km northwest of the city of Montréal (Figures 4.1 and 4.2). The coordinates for the approximate centre of the Property are 48°05’N latitude and 77°46’W longitude on NTS map sheet 32C/04 (National Topographic System quadrangle). The Property consists of parts of four contiguous mining concessions (“CM”) and 20 mining claims (“C”; Table 4.1 and Figure 4.3). The Property is registered 100% to Integra Gold for a total of 1,459.42 ha. None of the claims are within park or forest reserves that are restricted from exploration and mining. The PEA presented in this Report mainly concerns mining claims C002091, C002092, C002093, 3691171, 5275588 and 5275589. The four mining concessions (CM375, CM380, CM264 and CM314) grant surface and mineral rights, have no expiry dates, and remain in good standing provided a small amount of work is carried out or payment in lieu of work is made each year. The mining concessions have been legally surveyed. The status of the claims was verified using GESTIM, the Québec government’s claim management database system available on the Ministry of Natural Resources (“MRN”: Ministère des Ressources naturelles) website: https://gestim.mines.gouv.qc.ca. There are no surface rights related to the land holdings. The current Lamaque Property includes much of what constituted the former mining property of Lamaque Gold Mines Ltd, which for parent company Teck Hughes Gold Mines Ltd (later Teck Corporation, now Teck Resources Ltd), produced a total of 4,554,167 ounces of gold between 1935 and 1985. The current Lamaque Property does not include the four main mines that provided the majority of the ore for the former property owners.

www.innovexplo.com


Figure 4.1 – Regional setting of the Lamaque Gold Project

www.innovexplo.com


Figure 4.2 – Detailed location map of the Lamaque Property

www.innovexplo.com


Table 4.1 – Mining titles comprising the Lamaque Project

NTS Sheet

Type of Title

Title No Expiry Date Area (Ha)

Excess Work

Required Work

Titleholder(s) (name, number

and percent interest)

NTS 32C04

CL C002081 2017-01-27 23:59 21.00 $32,258.43 $1,000 Integra Gold Corp.

(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04

CL 9731 2015-08-09 23:59 16.00 $15,630.76 $1,000 Integra Gold Corp.

(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04

CL 1348121 2017-01-26 23:59 16.00 $0.00 $1,000 Integra Gold Corp.

(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04


(86793) (100%)

NTS 32C04

CL 5275588 2017-01-18 23:59 16.00 $2,750.00 $750.00 Integra Gold Corp.

(86793) (100%)

NTS 32C04

CL 5275589 2017-01-18 23:59 9.00 $2,750.00 $750.00 Integra Gold Corp.

(86793) (100%)

NTS 32C04

CM 375

324.13

$35.00 per hectare

Integra Gold Corp. (86793) (100%)

NTS 32C04

CM 380

195.75

$35.00 per hectare


NTS 32C04

CM 264PTA

246.65

$35.00 per hectare


NTS 32C04

CM 314PTA

392.49

$35.00 per hectare


www.innovexplo.com


Figure 4.3 – Map of mining concessions and claims constituting the Lamaque Project

www.innovexplo.com


4.2. Agreements and Encumbrance

Following the closure of the Lamaque mine, Teck Corporation (“Teck”) and Golden Pond Resources Ltd. (“Golden Pond”) formed the Teck-Golden Pond joint venture (“JV”) and Teck and Tundra Gold Mines Inc. (“Tundra”) formed the Teck-Tundra JV to explore a portion of the historical Lamaque mine property in 1985. The Golden Pond JV and some of the Tundra JV covered most of the ground now owned 100% by Integra Gold. In June 2003, Kalahari Resources Inc. (“Kalahari”) entered into an option agreement with Teck Cominco Ltd (“Teck”) to earn an interest of 50% to 53% (depending on the claims) in approximately 1,244 hectares called the Lamaque Project. On September 22, 2009, Kalahari entered into an option agreement with Alexandria Minerals Corp. to earn a 100% interest in the Roc d’Or East Extension property. Over a 3-year period, consideration was $25,000 cash (paid) and 500,000 shares (issued). There is a 2% NSR payable on the property, of which half (1%) may be purchased for $1,000,000. This claim group is adjacent to the Issuer’s 100% owned Roc d’Or East claims, which are now part of its Lamaque Property. In October 2009, Kalahari entered into separate agreements with Tundra and Golden Pond, joint venture parties at that time on the Lamaque Property, to purchase their interests in order to consolidate a 100% ownership of the Property and allow for more advanced exploration to be initiated. This meant issuing 9,593,128 shares to Tundra and 2,902,861 shares to Golden Pond, which asset could then be distributed in either cash or shares to shareholders, pro rata as determined by the boards of directors of Tundra and Golden Pond. This share consideration amounted to approximately one share of Kalahari for every dollar spent by Tundra and Golden Pond on the Property. Kalahari changed its name to Integra Gold Corp. in December 2010 and now owns 100% of the Property. Integra Gold also acquired 100% of Teck’s interest in the adjacent Roc d’Or East and Roc d’Or West claims. There is a 2% NSR payable to Teck on the Property, of which half (1%) may be purchased for $2,000,000 at any time within one year of commercial production. In December 2010, Integra Gold acquired an option to earn a 100% interest in the Bourlamaque property in Bourlamaque Township, Québec, adjacent to the Lamaque Property. Consideration for the property acquisition consisted of $3,500 cash (paid) and 10,000 shares (issued). Integra Gold purchased the entire NSR for $5,000 on April 30, 2013 (no outstanding NSR). An agreement is in place between Teck and Integra Gold concerning the tailings of the Old Lamaque mine. Integra Gold must provide a reclamation plan for the historical tailings and pay for the required work in order for Teck to be potentially liberated from its obligation, pending approval from the (MRN).

4.3. Environmental Obligation

Part of the Property is covered by tailings from the previous mining operations at the Lamaque Mine from 1935 to 1985 (Figure 4.2). The tailings are largely confined by a

www.innovexplo.com


tailings dam. The dam and tailings are stable and fairly dry and support grasses and herbaceous growth. Teck (and its predecessors) previously planted conifers on various parts of the tailings and these trees appear to be doing well. The tailings do not appear to have the potential to generate acid mine drainage. Teck still retains legal responsibility for the tailings. Depending on season and weather, parts of the tailing are wet, with standing water and intermittent creeks. Excess water exits the tailing area through runoff over two open spillways. No environmental monitoring is required. In order to conduct exploration work, Integra Gold must respect all laws relative to exploration, and request all appropriate forestry intervention permits from the MRN for all drilling and trenching related activities. To drill on the historical Lamaque tailings pond, a reclamation plan must be submitted to the MRN.

4.4. Risks and Uncertainties

The Issuer is subject to a number of risks and uncertainties due to the nature of its business. The Issuer’s exploration and development activities expose the Issuer to various financial and operational risks. Readers are advised to study and consider risk factors stressed below. The Issuer’s future financial success depends on the ability to raise additional capital from the issue of shares or the discovery of property which could be economically justifiable to develop. Such development could take years to complete and resulting income, if any, is difficult to determine. The sales value of any mineralization potentially discovered by the Issuer is largely dependent upon factors beyond the Issuer’s control, such as the market value of the products produced. The resource exploration industry is an inherently risky business with significant capital expenditures and volatile metals markets. The marketability of any minerals discovered may be affected by numerous factors that are beyond the Issuer’s control and which cannot be predicted, such as market fluctuations, mineral markets and processing equipment, and changes to government regulations, including those relating to royalties, allowable production, importing and exporting of minerals, and environmental protection. This industry is intensely competitive and there is no guarantee that, even if commercial quantities are discovered, a profitable market will exist for their sale. The Issuer competes with other junior exploration companies for the acquisition of mineral claims as well for the engagement of qualified contractors. Metal prices have fluctuated widely in recent years, and they are determined in international markets over which the Issuer has no influence. Exploration and development on the Issuer’s Property are affected by government regulations relating to such matters as environmental protection, health, safety and labour, mining law reform, restrictions on production, price control, tax increases, maintenance of claims, and tenure. There is no assurance that future changes in such regulations would not result in additional expenses and capital expenditures, decreasing availability of capital, increased competition, title risks, and delays in operations.

www.innovexplo.com


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

The Lamaque Property is immediately adjacent to the town of Val-d’Or, partially overlapping its eastern urban limits. Val-d’Or is located along Highway 117, east of the towns of Malartic and Rouyn-Noranda. Highway 117 crosses the northern part of the Property and provides access to it. The Property is accessible from Val-d’Or via Highway 117 and other public roads, as well as bush roads suitable for 4-wheel drive vehicles. One of the most prominent and useful of these gravel roads is the one built on the tailings retention dykes. The Val-d’Or airport is located on the southern edge of the Property and has regularly scheduled flights to and from Montréal. Val-d’Or, with a population of approximately 35,000, is a modern city and one of a number of communities in the Abitibi region of Québec with a long and rich mining heritage. Both the Lamaque and Sigma mines are situated within the municipality of Val-d’Or. These two mines were the largest producers in the area. All necessities to support a mining operation, including hydro-electric power, are available in Val-d’Or and the surrounding area. On or near the Property is an ample supply of timber as well as enough gravel and water to supply a mining operation. There is a plentiful supply of skilled workers for all aspects of exploration and mining available in the area. A large swamp partially covers parts of the Property, while spruce forest and mixed deciduous and coniferous forest cover the eastern, western and southern extremities. The relief rarely exceeds 50 m, except where eskers and glacial deposits are found. The Property is at an elevation of about 320 m above sea level. The tailing retention area covers a large part of the central part of the Property. As described under section 4.3, Environmental Issues, this tailings area is generally populated with herbaceous growth, grasses and areas of small trees planted by Integra Gold’s predecessors. Spruce forest and mixed deciduous and coniferous forest cover much of the rest of the Property. Based on Environment Canada statistics, from 1971 to 2000, the region is characterized by a mean daily temperature of 12° Celsius (C). The month of July has an average temperature of 17.2° C, whereas the month of January averages -17.2° C. The extreme minimum recorded temperature was -43.9° C, whereas the highest recorded temperature was 36.1° C. There were 209 days recorded below freezing point. The average annual precipitation is 954 mm. The month of September receives the highest average precipitation with 101.5 mm of rain. However, July has the highest daily precipitation with 68 mm of rain. Snow falls from October to May with the highest snowfall between November and March. The monthly average precipitation (in mm of rain) for this six month period is 54 mm.

www.innovexplo.com


6. HISTORY

The production, resource and reserve estimates, indicated below in this Item, are of historical nature and do not comply with NI 43-101. However, the authors believe that this information gives a conceptual indication of the potential of the area, which is pertinent to the Report. The qualified persons have been unable to verify the information and the information is not necessarily indicative of the mineralization on the property that is the subject of the Report.

6.1. Before 2003

The early history of the Lamaque Mine is described by Wilson in 1948. He describes the original gold discovery by R. Clark in 1923 and subsequent staking of the original claims the same year followed by prospecting, stripping and trenching. In 1928, Read-Authier Mines Limited was formed to acquire the historical Lamaque property and in 1929, nineteen (19) surface diamond drill holes totalling 2,143.05 m (7,031 ft) were drilled. In 1932, Teck-Hughes acquired an option on the property and drilled five holes totalling 519.68 m (1,705 ft) to confirm previous results. The 1932 drilling returned positive results and Teck-Hughes exercised its option, forming Lamaque Gold Mines Limited (a wholly owned subsidiary of Teck-Hughes), to take over the original property and a number of the adjoining claims. Wilson reports that a shaft was sunk beginning in January 1933 and lateral work followed as well as construction on the original mill in the summer of 1934. The mill started operations at 250 short tons per day in April 1935; was increased to 500 short tons per day by the end of 1935 and to 1,000 tons per day by December 1937. Subsequently, further shafts were sunk adjacent to the Main Mine or No. 1 Mine. These shafts include the number 2, 3, 4, 5, 6 and 7 shafts and the development of the East Mine and West Mine areas. The No. 2 Mine, approximately 1,158.24 m (3,800 ft) northeast of the Main Mine area (not to be confused with the No. 2 Shaft located on the Lamaque Property), close to and on the extension of Sigma Mine structures, was developed in 1950-1951 to a depth of 410.56 m (1,347 ft) with nine levels developed. Production from the No. 2 Mine ceased on November 30, 1955. None of the Main, East or West mines (comprising the Main Mine and Mill Complex), or No. 2 Mines are on the property (Figures 4.2 and 4.3 for location); but they all contributed to the tailings that cover part of the property. In late 1955 a new discovery approximately 1,371.6 m (4,500 ft) southeast of the Main Mine was made and in late 1960/early 1961 a shaft was sunk on the new zone with a three compartment vertical shaft to 146.3 m (480 ft) with three levels. In the summer of 1961, development work for three zones in the No. 3 Mine took place. This shaft, known as the No. 3 Mine shaft, along with its mineralization, is situated underneath the actual tailings. Historical records indicate that the No. 3 Mine yielded 152,015 short tons grading 0.220 oz/t Au. Subsequently, drifts were extended toward the No. 4 Plug from the No. 3 Mine area to provide access at the 450-ft (137.16 m) and 700-ft (213.36 m) levels. The No. 4

www.innovexplo.com


Plug was mined and yielded 160,000 short tons grading 0.152 oz/t Au from workings above the 700-ft level. Production for the No. 3 Mine and No. 4 Plug areas extended from July 1961 to the end of 1967. On the 300-ft (91.44 m) level the No. 3 Mine is also connected with the upper part of the No. 5 Plug. In May 1985, all production at the Lamaque Mine ceased. The Lamaque mill was kept on a stand-by basis for custom milling until 1986. Total production from the historical Lamaque Mine from 1935 to 1985 totalled 4,554,167 ounces of gold distributed over nine (9) mineralized zones (Table 6.1). Table 6.1 – Summary of Production for Historical Lamaque Property

SUMMARY OF PRODUCTION FOR HISTORICAL LAMAQUE PROPERTY

Zones not located on the property

Zone Tons Milled Grade Au oz/ton Total ounces % of Total

Main 20,025,627 0.185 3,695,194 81.14

East 2,999,842 0.115 343,827 7.55

West 1,644,606 0.133 219,014 4.81

No.2 Mine 1,634,488 0.145 237,596 5.22

Zones located on the property

No. 4 Plug 160,973 0.152 24,497 0.54

No.3 Mine 152,015 0.22 33,423 0.73

No. 5 Plug 5,572 0.111 616 0.01

Total: 26,623,123 0.17 4,554,167 100.00

Note: Grade average is weighted average by tons milled

Post-shutdown, Teck and Golden Pond Resources Ltd. (“Golden Pond”) formed the Teck-Golden Pond joint venture (“JV”) and Teck and Tundra Gold Mines Inc. (“Tundra”) formed the Teck-Tundra JV to explore a portion of the historical Lamaque property. The Golden Pond JV and some of the Tundra JV covered most of the ground now owned 100% by Integra Gold (excluding the Lamaque Mine area under the Tundra JV), but in addition included two small claims on the southern limit of the Villemaque Block (claims previously identified as 422883-2 and 421475-2). The Tundra JV also included two non-contiguous parcels: the first parcel of land centred on the No. 5 Plug and the second parcel centred on the No. 4 Plug. Teck was the operator for both the Golden Pond and Tundra JV programs. Subsequently, in December 1988, Tundra signed an agreement with Teck to acquire a 100% interest in all of Teck’s assets at Lamaque. The assets to be acquired included the Main Mine property, all surface structures including the mill, surface and underground equipment, and Teck’s interest in the Tundra, Golden Pond and Roc d’Or Mines agreements. The purchase price for the assets was $8,000,000. Tundra was also required to complete an exploration program and sink an exploration shaft to 304.8 m (1,000 ft) on the No. 4 Plug. Preliminary work was initiated to meet the obligations of the agreement but a downturn in the industry made funding difficult and the 1988 option was never exercised, leaving Teck with a 100% interest in the

www.innovexplo.com


Main Mine and mill area, which was eventually optioned and purchased by Placer Dome Inc. Subsequently Tundra’s and Golden Pond’s interest in the Tundra and Golden Pond JV properties was diluted to 50% due to non-payment of their respective portions of lease rentals, assessment filings and taxes. No exploration was conducted on the Tundra and Golden Pond JV properties between 1990 and 2003, when Kalahari and Teck Cominco signed their agreement providing Kalahari the option to earn Teck’s interest in the JV properties. In 2009, Kalahari purchased the remaining Tundra and Golden Pond interests in the properties through a share swap. Kalahari changed its name to Integra Gold Corp. in December 2010 and now owns 100% of the Property.

6.2. From January 2003 to June 2013

During the period between January 2003 and June 2013, exploration work was completed on the Lamaque Property, mainly via drilling campaigns. A total of over 115,326 m of drilling was completed, essentially distributed on the Fortune (previously known as Forestel), Parallel (including No.10 Vein), Triangle, No. 6 Vein, No. 4 Plug, No. 5 Plug, Sigma, Mylamaque, Sixteen Zone, and various geochemical and geophysical targets (Table 6.2). The various drilling programs, and their results, were discussed in detail in previous version of NI 43-101 Technical Reports, all of which were filed on SEDAR. The drilling was completed by Orbit-Garant Drilling from Val-d’Or, Québec. Analyses were completed by Bourlamaque Assay Laboratory and ALS Canada in Val-d’Or. Supervision of the exploration work completed between 2003 and 2008 was conducted by Don Cross and Terrence Coyle; between 2009 and to the end of 2013, exploration orientation, geoscientific compilation, drill planning, core logging, data validation and geological mineralized zone interpretation of cross-sections and longitudinal sections were conducted and supervised by Geologica Groupe-Conseil Inc. From 2009, drill supervision, core logging and QA/QC sampling protocols were designed and followed by Geologica personnel with duplicates, blanks and standards inserted for each drill hole. A re-sampling program of 2003-2008 diamond drill cores was conducted by Geologica, beginning in 2009. This re-sampling includes the addition of QA/QC samples and to date has included a total of 1,654 samples along with 319 QA/QC samples on 121 diamond drill holes.

www.innovexplo.com


Table 6.2 – January 2003 to June 2013: drilling & exploration Target Goal Drill Holes Meters

Fortune Zone (Formerly Forestel)

Confirm east-west extension of mineralized auriferous, vein clusters

43 12,208

Parallel Zone including #10 Vein

Confirm east-west extension of mineralized auriferous, vein clusters

88 22,208

Sixteen Zone Verify the lateral extension of drill holes completed in 2003 – 2004

56 12,823

Triangle Zone Verify east-west strike and depth extensions 55 20,014

Sigma Test the potential extension of the quartz carbonate-tourmaline veins intersected by Century Mining in the Sigma Mine project

10 2,862

No.4 Plug Verify and validate past drill hole results 46 24,417

No.5 Plug Verify and validate past drill hole results 12 5,014

Mylamaque Verify and validate past drill hole results 25 6,347

No. 4 Vein Validate previous results 1 300

No. 6 Vein Validate previous results and extend the mineralized vein structures

30 7,555

Geophysical Targets

Verify two airborne electromagnetic anomalies 2 585

West Property Explore the southwest of the property 3 438

Other Zone School drill holes 2 555

Following the drilling program planning, drill core logging, the data validation and geological interpretation of the mineralized zones by Geologica between October 2009 and November 2013, resource calculations were completed in 2013 on the Fortune, Parallel, Triangle and No. 4 Plug zones by GéoPointCom (Christian D’Amours) with help from Geologica’s Val-d’Or team. The result was Indicated Resources of 2,844,200 tonnes at 7.2 g/t Au (661,290 oz Au) and Inferred Resources of 738,000 tonnes at 11.3 g/t Au (268,870 oz Au).

www.innovexplo.com


7. GEOLOGICAL SETTING AND MINERALIZATION

7.1. Regional and Local Geology

The Property is located in the Malartic Composite Block, or Malartic Group, in the southeastern section of the Abitibi Greenstone Belt, a 750 km long by 250 km wide Island Arc volcanic complex, of the Archean Superior Province, Canadian Shield. The Grenville Front is located about 50 km to the southeast and the Cadillac-Larder Lake Fault Zone, deformation zone or "Break" is located about one km south of the southernmost edge of the Property. The Cadillac-Larder Lake Break runs for about 250 km and is the fundamental control for gold mineralization in the Val-d’Or, Malartic, Cadillac and Kirkland Lake camps (Figure 7.1). In the Malartic/Val-d’Or area, the Cadillac-Larder Lake Break separates the dominantly volcanic Malartic Block from the Pontiac Metasedimentary Subprovince to the south. The Bourlamaque Batholith hosts a number of former gold producers (Bras d’Or, Belmoral, Stabell) and is located one mile northeast of the northeast corner of the Property while the East Sullivan Stock is located to the east. The regional stratigraphy for the area is summarized in the Table 7.1 below.

Table 7.1 – Regional Stratigraphy Unit Name Description Age

Pontiac Group Greywacke and mudstone with minor conglomerate and ultramafic rocks

2691-2682 Ma

Kewagama Group Greywacke

Bourlamaque Pluton

Tonalite-granodiorite 2701±1Ma

Louvicourt Group

Héva Formation Mafic to felsic volcaniclastics and volcanics 2702±2Ma

Val-d’Or Formation Intermediate to felsic flows intercalated with volcaniclastic rocks

Transition Zone 2704±2Ma

Malartic Group

Jacola Formation Mafic volcanics and hyaloclastites with some intercalated ultramafics

Upper Dubuisson Mafic and ultramafic volcanics

Lac Caste Sedimentary Rocks

Lower Dubuisson Mafic and ultramafic volcanics

La Motte-Vassan Predominantly ultramafic volcanics with volcaniclastic interbeds 2714±2Ma

www.innovexplo.com


Figure 7.1 – Regional Geology

www.innovexplo.com


7.2. Local and Property Geology

The Property is principally underlain by volcanic flows and volcaniclastics of the Val-d’Or Formation (2705-2703 Ma), intruded by a variety of intermediate to mafic plugs, dykes and sills. The northern part of the Property is underlain by lithologies comprising the lower part of the Val-d’Or Formation, while the mid to southern part of the Property is underlain by units of the upper Val-d’Or Formation. The extreme southern section of the Property, in the vicinity of the airport, is underlain by mafic volcanics of the Héva Formation. There are twelve (12) zones of gold mineralization, including six (6) gold deposits with known resources on the Lamaque Property. They consist of various simple vein structures, vein complexes and stockwork zones. They are:

Parallel Zone (including historical No. 10, No.5 and No.7 Veins); Fortune Zone (previously known as Forestel); No. 4 Plug; No. 5 Plug (including No. 35 Veins); No. 3 Mine (including No. 1 and 2 Veins); Triangle Zone; South Triangle Zone; Mylamaque Zone; No.4 Vein; No. 6 Vein; Sixteen Zone; and Sigma Zone.

The six (6) presently known gold deposits are grouped into the North, South and West clusters. The North cluster consists of the Parallel and Fortune Zones but also the exploration targets known as the No. 5 Plug and No. 3 Mine. The South cluster consists of the No. 4 Plug, Triangle, and South Triangle zones. The West cluster consists of the Sixteen Zone and No. 6 Vein. Figure 7.2 presents the location of the various zones. For clarity, the West Cluster zones are not included in the PEA.

www.innovexplo.com


Figure 7.2 – Property Geology and Mineralized Zones

www.innovexplo.com


The Val-d’Or Formation is characterized by interstratified and overlapping massive to pillowed lavas and volcaniclastics of andesitic to rhyolitic composition. Reports also mention one horizon identified as an exhalative chert horizon and in another instance, a horizon of magnetic agglomerate. Ground magnetic surveys and interpretations for the most part have failed to resolve volcanic units of contrasting magnetic susceptibility. Areas of enhanced magnetic susceptibility have generally been interpreted to represent various intermediate porphyritic intrusives in the host volcanics rather than volcanic units of contrasting magnetic susceptibility. Magnetic surveys to date have therefore been of little help in deciphering stratigraphic and structural relationships. According to Scott (2002) and based on yttrium-zirconium ratios, the volcanics from the lower Val-d’Or Formation are tholeiitic to calc alkalic transitional while up-section, to the south, the upper Val-d’Or units, are transitional tholeiitic to calc-alkalic. Because of their intimate spatial association with most of the known gold mineralization, the intermediate to mafic, mostly granodiorite to dioritic plugs and associated porphyritic dykes and sills that intrude the Val-d’Or Formation are of particular importance. Structural and temporal relationships, however, remain enigmatic for some of these porphyritic rocks because of poor outcrop exposure and dominance of the pervasive and obliterating tectonic fabric. Most of the gold production at the Lamaque Mine (immediately to the north of the Property) came from or was immediately adjacent to the Main Mine, largely in the Main Plug. Significant production also came from the East Plug at the east end of the Main Mine (not on the Property). The Main Plug, measuring only 800 ft by 380 ft (243.84 x 115.82 m), is a chimney-like body and concentrically compositionally zoned. The outer rim of the body, in contact with its host volcanics, is diorite. Its diorite rim grades into a porphyritic quartz diorite and the core of the pipe is granodiorite. U-Pb zircon ages of 2685±3 Ma and 2682 Ma have been obtained on the Main Plug (Jemielita et al, 1990). The West Plug, southwest of the Main Plug, is only developed on a few levels and is a coarse porphyritic granodiorite. Similar pipe-like porphyritic intrusions, the No. 4 and No. 5 Plugs, occur on the Property. The No. 6 Plug may be partly on the Property. The No. 4 Plug measures 350 to 400 ft (106.68 x 121.92 m) in diameter and extends to a depth in excess of 3,000 ft (914.4 m). It appears to be a composite intrusion with a western portion which is granodioritic and an eastern portion which is a fine- to medium-grained diorite. Burrows and Spooner (1989) reported that the Main, East and West Plugs are unusually sodic and enriched in barium and strontium and suggest these compositions are magmatic. Watts, Griffis and McOuat and Geologica consider that this sodium, barium and strontium enriched composition is more likely a reflection of gold-related hydrothermal alteration. As pointed out by Patton (1988) these plugs are foliated and have a consistent gross orientation with a steep (70o) northeastern plunge. U-Pb zircon geochronology dates the Main Plug at 2682 to 2685 Ma (Corfu, 1993) making them Timiskaming in age, post volcanic but syntectonic. More obscure are the various other porphyritic dykes and sills commonly intersected in drill holes and possibly concordant or sub-parallel with general stratigraphy, but perhaps also extensively cross cutting. These sills and dykes, felsic to intermediate in composition units, are also important because of their common association with gold mineralization. Many of these bodies are dykes and/or sills but given the difficulties of geological mapping, poorly defined deformation and pervasive hydrothermal alteration, we would not rule out the possibility that some of these rocks represent flows or synvolcanic intrusions.

www.innovexplo.com


The volcano-sedimentary rocks have been metamorphosed to lower greenschist facies (quartz + sericite + albite + chlorite + carbonate + actinolite + chloritoid). Scott (2002) reports that all flow units of the Val-d’Or Formation have undergone silicification, sericitization and carbonatization to some degree, probably as a function of their original porosity. Throughout the drill logs prepared by Teck for the Golden Pond and Tundra JVs, various types and extents of alteration are documented. In most cases the alteration that is described is immediately peripheral, or enveloping quartz-tourmaline veins in either volcanics or intermediate porphyritic rocks. The volcanics, on a regional scale, strike east-west to northeast-southwest and dip steeply to the north or south. Robert and Brown (1986) state for units south of the Bourlamaque Batholith, bedding is overturned with younging directions mainly to the south. At the regional scale and on the Property, an east-west, steeply dipping tectonic fabric superimposed on the volcanic rocks obscures bedding in the volcanics. Teck’s latest geological compilation maps by Madon (c. 1987) show very few stratigraphic orientations. Most of the measurements for orientation on outcrops are for schistosity, which is shown to be steeply dipping to vertical. Rarely are bedding attitudes shown. The few primary attitudes that have been mapped are contacts between volcanic rocks of contrasting texture and these often do not appear to parallel the strike of interpreted stratigraphy. Interpreted stratigraphy also does not always appear to parallel the trend of magnetic horizons. Patton (1988) suggested that the volcanic rocks may be tightly folded with the "plugs" located at the fold noses. Tight folding would be compatible with regional interpretation. Patton also suggested that the "plugs" are in fact folded, and thickened sills or dykes occurring in the fold noses. Burrows and Spooner (1989) also suggested that the plugs represented the coalescence of parallel dykes.

The effectiveness of exploration programs might be enhanced if more detailed interpretation of the stratigraphic and structural relationships of the various lithological units could be achieved.

7.3. Mineralization

The more significant zones of gold mineralization or deposits known on the Property contain various simple vein structures, vein complexes and stockwork zones. These zones are listed below:

Triangle Zone (South cluster); South Triangle Zone (South cluster); No. 4 Plug (South cluster); Fortune Zone (North cluster); No. 5 Plug (North cluster); Parallel Zone (North cluster, including historical No. 10, 7 & 5 Veins); No. 3 Mine area (North cluster, including No 1 & 2 Veins); No. 6 Vein (West cluster); No. 4 Vein; Sixteen Zone (West Cluster); Mylamaque Zone; and Sigma Vein.

www.innovexplo.com


From these mineralized zones and Plugs, the most significant were subdivided into three main clusters: North cluster (Fortune, Parallel, No. 3 Mine, No. 5 Plug); South cluster (No. 4 Plug, Triangle, South Triangle); West cluster (Sixteen, No. 6 Vein).

7.3.1. North Cluster

7.3.1.1. Fortune Zone (previously known as Forestel Zone)

The first drill holes were completed by Kalahari (now Integra Gold) in the area north of the T-06 Magnetic anomaly in 2003 and drilled further to the north but cut no significant mineralization. The first hole drilled in the area in 2004, T-06-04-01, was designated as the Forestel Zone discovery hole (now called Fortune). It intercepted a zone of 0.183 oz Au/ton over 16.5 ft (≈ 5 m). Subsequently, eight additional holes were drilled to the east and west on sections at 100-ft (30.48 m) intervals and above the up-dip projection of the original discovery. Several of these holes intersected several zones of mineralization comprising quartz-carbonate and quartz-tourmaline veins with fine pyrite in foliated or tuffaceous mafic volcanics. Mineralization in holes T-06-04-06, T-06-04-02 and sporadic mineralization in T-06-04-08 up-dip from the original discovery can perhaps be correlated with the original discovery intersection. Other correlations are more tenuous. Assuming that the Fortune Zone represents veins in a series of steeply dipping, east-west trending P and Riedel shears, true widths for the discovery zone (intersection length -16.5 ft or 5 m) might be in the order of 60% of this figure for about 10 to 12 ft (3.04 to 3.65 m).<

7.3.1.2. Parallel Zone (including Vein No. 10)

Parallel Zone and the No. 10 Vein complex are both located northwest of the No. 3 Mine (Figure 7.2). They consist of several sub-parallel en echelon pinch and swell veins, which are hosted within a narrow tabular shear zone or corridor that crosscuts all the lithological units. The vein itself generally ranges in width between 6 and 60 inches (15.24 and 152.4 cm) and attains a maximum thickness of 103 inches (approximately 261.62 cm) in drill hole G-88-8. The higher-grade portion of this vein extends from section 111E to section 156E, a strike length of 1,150 ft (350.52 m). It dips between 50° and 60° towards the south and has a known vertical extent of 400 ft (121.92 m) between 300 and 700 ft (91.44 and 213.36 m) below the surface. The majority of the mineralization is hosted within fine- to medium-grained diorite with generally 1 to 3% of disseminated pyrite, locally up to 5%. Some zones are within intermediate to mafic volcanics with 1 to 3% of disseminated pyrite. The veins are hosted by a series of andesitic to dacitic flows, crystal tuffs and lapilli tuffs. Rocks proximal to the vein are reported as silicified. The highly siliceous zones are dyke-like in shape and competent. The vein or vein segments are composed primarily of quartz and carbonate with lesser amounts of tourmaline and chlorite. Pyrite is the predominant sulphide mineral but it is rarely present in quantities greater than 5%. Typically in the wider veins the pyrite and gold are confined to the margins of the vein, while the bulk of the vein is made up of bull quartz, carbonate fragments and tourmaline.

www.innovexplo.com


The most recent interpretation of the No. 10 vein system for Teck by H. Hugon indicated that the veining is disrupted by a series of near-vertical normal faults. These faults strike approximately east-west and have displaced the vein between 10 to 20 ft (3.48 to 6.09 m). They do not appear to be genetically related to the mineralization since they are generally barren except in close proximity to the veining.

7.3.1.3. No. 5 Plug

The No. 5 Plug is located in the northwest part of the Property, approximately 500 m northeast of the Parallel Vein Zone and 790 m north of No. 3 Mine Shaft, to which it is connected by a drift on the 700-ft (213.36 m) level. The No. 5 Plug was partially mined and yielded 5,572 short tons grading 0.11 oz Au/ton (5,054 tonnes grading 3.8 g Au/t) from workings above the 3,200-ft (975 m) level. It is located 1.5 km from the Main Mine shaft. The No. 5 Plug main geological unit is a fine- to medium-grained diorite, 200 to 250 m in diameter with depth potential expansion. It is mixed with a granodioritic intrusion. The two intrusive units have similar competency and therefore should be equally good hosts for gold-bearing quartz veins; however, mining was restricted to the dioritic portion of the intrusive.

7.3.1.4. No. 3 Mine (No. 1 Vein)

No. 1 Vein is located 300 ft (91.44 m) north of the No. 3 Mine shaft, southeast of the No. 10 Vein complex (Figure 5). The vein is continuous over an east-west strike length of about 1,000 ft (304.8 m), extends from the 700 ft to 1,200 ft (213.36 to 365.76 m) levels, and dips to the south at 80°. It ranges in width from 30 to 42 inches (76.2 to 106.68 cm). Potential for additional mineralization exists at depth where the vein was intersected in four other holes. None of these, however, encountered economic values. The auriferous zone most probably follows a plunge towards the east as is often observed for gold concentrations in ore shoots of the camp in general. The vein is composed primarily of quartz and carbonate with lesser amounts of tourmaline. Pyrite is the predominant sulphide mineral but it is rarely present in quantities greater than 5 to 7%.

7.3.2. South Cluster

7.3.2.1. No. 4 Plug

The No. 4 Plug is located on the southeastern part of the tailing, approximately 3,500 ft (1,066.8 m) southwest of No. 3 Mine Shaft, to which it is connected by drifts on the 450 and 700-ft (137.16 m and 213.36 m) levels. The No. 4 Plug was mined and yielded 160,000 tons grading 0.152 oz Au/ton (145,136 tonnes grading 5.21 g Au/t) from working above the 700 ft (213.36 m) level. Production for the No. 3 Mine and No. 4 Plug areas extended from July 1961 to 1967. The No. 4 Plug is composed of an easterly portion of fine- to medium-grained diorite, 300 to 400 ft (106.68 to 121.92 m) in diameter. It is enveloped on the west side by a granodioritic intrusion, which extends to the west for an additional 275 ft (83.82 m). The two intrusive units have similar competency and therefore should be equally

www.innovexplo.com


good hosts for gold-bearing quartz veins; however, mining was restricted to the dioritic portion of the intrusive. The intrusions are cut by a series of near vertical shears striking 70° to 90° and dipping 70° to 85° to the south. The best developed of these shears is the No. 6 fault/shear. It has a dextral displacement of approximately 200 ft (60.96 m) and vertical displacement of 60 to 80 ft (18.28 to 24.38 m). The vertical shears have probably produced the brittle (Riedel) shears and tension fractures, which are believed to be the conduits along which the gold-bearing solutions migrated. The quartz veins are hosted by three dominant structural features:

a) ductile "P" Shears; b) brittle "Riedel" or "R" Shears; c) Tension Veins.

The best mineralization is interpreted as occurring in a series of stacked, tabular subhorizontal bodies of minimum mining width interpreted to be tension veins.

7.3.2.2. Triangle Zone

The Triangle Zone was discovered by drill testing an ovoid shaped magnetic anomalous zone which corresponded to a younger tonalite intrusion. The anomalous zone extends both east and west over a strike length of 393 metre and is 785 m wide. In drill core, the mineralized veins occur as part of a related vein cluster within the fractured and altered zones of the tonalite, which intrudes the surrounding intermediate to mafic volcanic rocks (mostly lapilli and blocky tuffs, undifferentiated pyroclastics and basalt). North of the Triangle Zone, two other similar ovoid shaped anomalous magnetic zones are believed to correspond to mineralized intrusions. Mineralized veins at Triangle generally occur within stacked, southerly dipping shear zones that are between 10 to 50 m from each other’s; at the moment a minimum of 12 stacked shear zones have been identified. Those shear zones, and the quartz / tourmaline / pyrite / gold veins they contain, extend for over 350 m along strike and 250 m down dip, and are generally constraint to the intrusive units although extension in intermediate to felsic pyroclastics has been recently recognized. Mineralized veins within the shear zones could be anywhere from 10 cm to 20 m in thickness.

7.3.2.3. South Triangle Zone

The South Triangle Zone was discovered by drill testing an ovoid shaped magnetic anomalous zone which corresponded to intermediate volcanic units with magnetite and pyrrhotite. The auriferous zone is not this ovoid anomaly but an auriferous quartz-carbonate-tourmaline-pyrite vein with free gold within blocky and lapilli tuff. This zone is presently recognized to extend some 100 m along strike (east-west) and for a minimum of 150 m at down dip.

www.innovexplo.com


7.3.3. West Cluster

7.3.3.1. Vein No. 6

Vein No. 6 is located approximately 2,000 ft (609.6 m) west of the No. 1 Main shaft (Figure 7.2). The vein has an east-west strike length of about 800 ft (243.84 m). Veins interpreted to be tension veins dip at about 50°. Steeper dipping veins (75°) are interpreted to be hosted in Reidel shears. The vein is hosted by a series of massive andesite or dacite and crystal tuffs and lapilli tuffs with several feldspar porphyry dykes. Generally a series of shear zones with vein injections cross-cut the dykes and tuffs. The vein itself generally ranges in width between 1 and 30 inches (2.54 and 76.2 m). The veins are composed primarily of quartz and carbonate with lesser amounts of tourmaline and chlorite. Pyrite is the predominant sulphide mineral but it is rarely present in quantities greater than 5%.

7.3.3.2. Sixteen Zone

The Sixteen Zone was named for the T-16 magnetic anomaly defined in a geophysical compilation from 1986. It is located in the western part of the Property, just north of the southern By-Pass road (Figure 7.2). This zone is oriented roughly east-west and has a strike length of 800 ft (243.84 m). The zone is a system of quartz/tourmaline/pyrite veins and associated alteration, within a steeply north-dipping, east-northeast trending, feldspar porphyry granodiorite dyke. The host dyke carries 15-20% medium grained, subhedral to euhedral, zoned feldspar phenocrysts in a sericitic light gray, aphanitic matrix. In general, foliation is weak but becomes strong to intense in the vicinity of the veins. This dyke, which can be up to 400 ft (121.92 m) wide, is one of several in a swarm that has intruded the surrounding lapilli to blocky tuff, and the dyke probably extends the length of the Property. Within the Property the mineralized zones are defined as mesothermal, structurally controlled deposits, which are typical of the Abitibi Greenstone Belt and abundant in other Archean greenstone belts around the world. The gold bearing quartz-tourmaline veins are young in geologic age and cut all rock-types, including irregular intrusive (diorite) bodies, which are affected by regional deformation and greenschist grade metamorphism. In the Sixteen zone, most of the gold intersection was in a porphyritic diorite and porphyritic granodiorite with 1 to 5% of pyrite in quartz tourmaline sericite carbonate veins. The quartz/tourmaline veins dip shallowly to the north and are usually confined to the core of the granodiorite dyke, although occasionally they are seen along the granodiorite/volcanic contact. Correlation between vein intersections in adjoining holes has proven difficult and it is not known if the veins are tabular and laterally extensive or if they are more cylindrical. The veins vary in true width from less than an inch to several ft. Tourmaline can be up to 60%, but most often comprises 15-20% of vein volume. Pyrite occurs as medium to very coarse pyrite cubes within the veins. Strong to intense sericite and silica alteration, plus or minus minor tourmaline and fuchsite, with gray to pink to beige selvage displaying strongly foliated, remnant, porphyritic, granodiorite texture

www.innovexplo.com


were observed. The selvages carry from 5 to 10% fine- to coarse-pyrite cubes. Gold is found along the contacts of tourmaline patches within the veins, as faint irregular gold-coloured patches in pyrite cubes, or as cryptic gold within the selvedges. It is possible that the veins and strongest alteration are concentrated in a pipe-like feature, the plunge of which is unknown. The occurrence of the zone within one granodiorite unit at this particular place is an enigma. Flexing and cracking of the granodiorite associated with a nearby controlling structure, acting as a conduit from a nearby plug or intrusive may be one explanation. However, at this time, all is conjecture and the relatively unexplored length of the granodiorite dyke remains an important drilling target.

7.3.4. Other Mineralized Zones

Several zones of mineralization (No. 35 Vein, North and South Shears) located north of, and/or within the No. 5 Plug were drilled by the Teck-Tundra JV from 1986 to 1988 (Figure 7.2). Numerous other zones of mineralization have been reported or intersected on the Property. Many of these zones were defined by intersections described in pre-1988 drill holes. In order to use this older data, a review of the older drill hole logs and a creation of a digital database is required so that the intersections can be plotted on plan and section. These zones, and new zones such as the Sigma Zone, will also be validated and verified with additional drilling and eventually evaluated with definition drilling permitting resource calculations to be completed.

www.innovexplo.com


8. DEPOSIT TYPES

Karvinen (1985) proposed a volcanogenic origin and control for the Lamaque gold mineralization suggesting that many of the porphyry units were volcanic and the associated gold mineralization was exhalative. Uranium-lead-zircon age dating of various porphyritic units and mineralization and structural relationships have since proven that the gold-bearing quartz-tourmaline veins are significantly younger than their host rocks. Burrows and Spooner (1989) proposed that the intrusive plugs (Main, West, and East), showed a primary, magmatic gold enrichment and inferred the ore bodies formed from the remobilization of this primary gold. According to Feng et al. (1993), the elevated gold values in samples from the intrusive are due to secondary enrichment in the halo of an ore-forming system. Original opinion (Wilson, 1948) was that the Lamaque mineralization represented structurally controlled deposits. Wilson describes mineralization at the Main Plug as controlled by a series of east-west trending stacked thrust faults, which occur one below the other and dip into the intrusive with the gold-bearing quartz-tourmaline veins in the thrust faults and subsidiary fractures. Wilson also describes how competency of the host rock controls the expression of deformation and foliation associated with the faults and fractures. Wilson reported that movement along the faults, in hard competent rock, resulted in considerable fracturing and brecciation and the development of numerous flatly dipping tension fractures hosting numerous gold-bearing veins. Similar descriptions to those related by Wilson are reported by Robert and Brown (1984 & 1986) in describing mineralization at the Sigma Mine, which represents the northern part of the same complex quartz-tourmaline vein system. The host rocks at Sigma, like those at Lamaque, are composed principally of andesitic flows, irregular intrusive bodies of porphyritic diorite and dykes of feldspar porphyry. Robert and Brown describe how the gold-bearing quartz tourmaline veins cut all rock-types, which include an irregular body of porphyritic diorite affected by regional deformation and greenschist grade metamorphism and younger feldspar porphyry dykes. Robert and Brown report that the gold-bearing veins were emplaced lately and were cut by diabase dykes and have not been subject to pervasive deformation and metamorphism. Two principal types of veins occur in the mine – subvertical veins and subhorizontal veins. Robert and Brown describe the subvertical veins as occurring within steeply dipping east-west trending ductile shear zones along which reverse subvertical displacements have taken place. The veins occupy openings created by irregularities during progressive movement within the shears. Subhorizontal veins occupy extensional fractures which formed between ductile shear zones, preferentially in the more competent host rocks. Neumayr et al (2000) present evidence that the mineralized zones and deposits at Lamaque and Sigma are second- and third-order structures related to deformation along the Larder Lake-Cadillac Fault Zone, which represents the first-order structure. Similar models are commonly expressed for most gold mineralization associated with both the Larder Lake-Cadillac Fault Zone and the Destor-Porcupine Fault zones where splay faults which represent second order structures off the principal fault

www.innovexplo.com


zones host the bulk of gold mineralization in the Timmins, Kirkland Lake and Val-d’Or Camps. Geologica agrees with Wilson (1948) and Robert and Brown (1986), that the mineralized zones are mesothermal structurally controlled deposits typical of the Abitibi Greenstone Belt and abundant in other Archean greenstone belts around the world. They believe, however, that in detail the property-scale controls for the zones of mineralization are not well understood. To Geologica’s knowledge there is no understanding and explanation for the clusters of gold-bearing veins in the No. 3 Mine, No. 10 Vein and No. 6 Vein, Fortune, Sixteen, Triangle or Mylamaque zones aside from junctions of structures, faults and fracture networks near and/or within younger altered intrusive plugs of intermediate to mafic composition.

www.innovexplo.com


9. EXPLORATION

Exploration work programs completed on the Property were described in detail in previous technical reports filed on SEDAR, and the reader is referred to those reports for that information:

NI 43-10 Technical Report on the Lamaque Property, June 23, 2011 - Geologica Groupe Conseil

NI 43-10 Technical Report on the Lamaque Property, Amended September

21, 2012 - Geologica Groupe Conseil

2013 NI 43-10 Technical Report on the Lamaque Property, November 1, 2013 - Geologica Groupe Conseil

2014 NI 43-10 Technical Report on the Lamaque Property, March 11, 2014 -

Geologica Groupe Conseil

www.innovexplo.com


10. DRILLING

Diamond drilling programs completed on the Property were described in detail in previous technical reports filed on SEDAR and the reader is referred to those reports for the information contained therein:

NI 43-10 Technical Report on the Lamaque Property, June 23, 2011 - Geologica Groupe Conseil

NI 43-10 Technical Report on the Lamaque Property, Amended September 21, 2012 - Geologica Groupe Conseil

2013 NI 43-10 Technical Report on the Lamaque Property, November 01, 2013 - Geologica Groupe Conseil

2014 NI 43-10 Technical Report on the Lamaque Property, March 11, 2014 - Geologica Groupe Conseil

During the period from June to December 2013, a total of 24,339.48 m of diamond drilling was completed on the South Triangle, Mine No. 3 and Parallel zones by Orbit-Garant Drilling of Val-d’Or, Québec. The drill core samples were assayed by Bourlamaque Assay Laboratory and ALS Chemex at their respective facilities in Val-d’Or, Québec. The planning, core logging, data validation and supervision of the 2013 drilling programs were completed by Geologica of Val-d’Or, Québec, with the help of Integra Gold technical assistance and logistics. Most of the drilling consisted of definition diamond drilling on the Parallel Zone and exploration drilling on the South Triangle Zone and No. 3 Mine targets.

10.1. South Triangle

During the June to December 2013 diamond drill program, Integra Gold completed thirteen (13) diamond drill holes on the South Triangle Zone for a total length of 6,965.67 m. The aim of this program was to verify a magnetometric anomaly and the depth extensions of some of the mineralized zones in the South Triangle Zone. Technical parameters are presented in Table 10.1 and drill hole locations are provided in Figure 10.1.

Table 10.1 – Table of technical parameters on the South Triangle Zone Drill hole East UTM North UTM Elevation Azimuth Dip Length(m)

TMS-13-01 296380.00 5327880.00 320.00 358.5 -48.2 440.90

TMS-13-02 296380.00 5327880.00 320.00 356.9 -64.2 541.12

TMS-13-03 296775.00 5327865.00 320.00 350.1 -55.3 575.64

TMS-13-04 296505.00 5327949.00 330.00 357.9 -55.6 434.42

TMS-13-05 296373.00 5328144.00 330.00 2.7 -51.2 1,079.76

TMS-13-06 296314.00 5328141.00 330.00 359.8 -56.1 1,163.10

TMS-13-07 296449.00 5328123.00 320.00 358.8 -56.1 1,138.36

TMS-13-08 296449.00 5328123.00 320.00 356.3 -72.5 468.20

TMS-13-09 296449.00 5328123.00 320.00 32.9 -54.9 15.11

TMS-13-09A 296449.00 5328123.00 320.00 19.7 -54.9 375.09

TMS-13-10 296449.00 5328123.00 320.00 19.1 -63.5 386.90

TMS-13-11 296448.00 5328129.00 320.00 350.5 -65.8 15.00

TMS-13-11A 296448.00 5328129.00 320.00 354.1 -66.4 332.07

www.innovexplo.com


Figure 10.1 – Diamond drilling program on the South Triangle Zone

www.innovexplo.com


The main lithologies reported in drill core from the South Triangle Zone are basalt, lapilli tuff and block tuff, with some minor sedimentary sequences. The volcanic and sedimentary units often alternate with small granodiorite-diorite or gabbro sills and/or dykes. Generally chloritized, sericitized and silicified, the units are sheared and fractured. Faults are locally observed. Several mineralized veins were noted; their length varies between 0.5 cm and 100 cm with trace to 5% pyrite (locally up to 10%), and very locally trace to 3% pyrrhotite. The veins are made up of quartz-carbonate-tourmaline-chlorite and/or sericite-feldspar; epidote and calcite are rare. Integra Gold used appropriate QA/QC protocols, employing duplicates, blanks and standards. A total of 3,249 core samples and 440 QA/QC control samples were collected; the total sampled length is 3,134.40 m (45% of total drill hole core length). Table 10.2 below presents the most significant intersections (complete results are available at the office of Integra Gold).

Table 10.2 – Table of significant assay results from the 2013 drilling program on the South Triangle Zone

Drill Hole #

From (m)

To (m) Length (m)

Sample #

Au (g/t)

TMS-13-03 427.00 428.00 1.00 180719 1.78

TMS-13-03 515.50 516.10 0.60 180803 4.35

MS-13-05 459.00 460.00 1.00 181186 1.84

TMS-13-05 460.00 461.00 1.00 181187 6.81

TMS-13-05 461.00 462.00 1.00 181188 3.97

TMS-13-05 462.00 462.80 0.80 181189 6.13

TMS-13-05 462.80 463.50 0.70 181191 5.15

TMS-13-05 562.00 563.00 1.00 181212 1.04

TMS-13-05 623.30 624.00 0.70 181227 3.6

TMS-13-05 624.00 625.00 1.00 181228 1.23

TMS-13-05 764.00 765.00 1.00 181306 1.33

TMS-13-05 765.60 766.60 1.00 181308 3.98

TMS-13-05 766.60 767.60 1.00 181309 4.85

TMS-13-05 768.50 769.50 1.00 181312 2.53

TMS-13-05 770.50 771.50 1.00 181314 1.12

TMS-13-05 815.00 816.00 1.00 181365 1.85

TMS-13-05 816.00 817.00 1.00 181366 1.34

TMS-13-05 852.00 853.00 1.00 181407 1.09

TMS-13-05 853.00 854.00 1.00 181408 1.25

TMS-13-05 898.00 899.00 1.00 181459 1.51

TMS-13-05 927.00 928.00 1.00 181494 1.18

TMS-13-05 929.00 930.00 1.00 181496 1.28

TMS-13-05 930.70 931.20 0.50 181498 3.77

TMS-13-05 931.20 931.70 0.50 181499 8.18

www.innovexplo.com


Drill Hole #

From (m)

To (m) Length (m)

Sample #

Au (g/t)

TMS-13-05 978.70 979.40 0.70 181557 6.88

TMS-13-05 996.30 997.00 0.70 181582 2.88

TMS-13-05 1002.00 1003.00 1.00 181589 1.12

TMS-13-05 1013.00 1014.00 1.00 181602 1.7

TMS-13-05 1015.00 1016.00 1.00 181604 1.14

TMS-13-05 1057.00 1058.00 1.00 181652 1.57

TMS-13-06 502.00 503.00 1.00 181881 4.26

TMS-13-06 504.00 505.00 1.00 181883 1.72

TMS-13-06 531.00 532.00 1.00 181887 3.3

TMS-13-06 765.00 766.00 1.00 181948 5.82

TMS-13-06 766.00 767.00 1.00 181949 12.05

TMS-13-06 767.00 768.00 1.00 181951 6.67

TMS-13-06 770.00 771.00 1.00 181954 1.4

TMS-13-06 771.00 772.00 1.00 181955 4.08

TMS-13-06 772.00 772.50 0.50 181956 4.29

TMS-13-06 902.00 903.00 1.00 182105 1.26

TMS-13-06 971.00 972.00 1.00 182184 16.13

TMS-13-06 974.00 975.00 1.00 182187 1.04

TMS-13-06 991.00 992.00 1.00 182206 2.25

TMS-13-06 998.00 999.00 1.00 182214 10.26

TMS-13-06 1008.00 1009.00 1.00 182226 1.38

TMS-13-06 1060.00 1061.00 1.00 182285 1.94

TMS-13-06 1061.00 1062.00 1.00 182286 1.1

TMS-13-06 1063.50 1064.30 0.80 182289 1.26

TMS-13-06 1064.30 1065.00 0.70 182291 2.55

TMS-13-06 1069.70 1070.50 0.80 182297 2.57

TMS-13-06 1071.00 1072.00 1.00 182299 9.26

TMS-13-07 289.00 290.00 1.00 182488 28.75

TMS-13-07 290.00 291.00 1.00 182489 22.57

TMS-13-07 291.00 292.00 1.00 182491 34.11

TMS-13-07 292.00 293.00 1.00 182492 3.01

TMS-13-07 294.80 295.30 0.50 182495 7.61

TMS-13-07 473.50 474.50 1.00 182535 2.85

TMS-13-07 762.00 763.00 1.00 182724 14.49

TMS-13-07 763.00 764.00 1.00 182726 4.5

TMS-13-07 786.00 787.00 1.00 182742 7.85

TMS-13-07 787.00 788.00 1.00 182743 4.41

www.innovexplo.com


Drill Hole #

From (m)

To (m) Length (m)

Sample #

Au (g/t)

TMS-13-07 788.00 789.00 1.00 182744 6.91

TMS-13-07 789.00 790.00 1.00 182745 1.54

TMS-13-07 799.00 800.00 1.00 182749 4.98

TMS-13-07 800.00 801.00 1.00 182751 26.69

TMS-13-07 801.00 802.00 1.00 182752 21.41

TMS-13-07 802.00 803.00 1.00 182753 3.03

TMS-13-07 803.00 804.00 1.00 182754 5.7

TMS-13-07 804.00 805.00 1.00 182755 1.43

TMS-13-07 945.00 945.50 0.50 182865 6.49

TMS-13-07 945.50 946.00 0.50 182866 4.94

TMS-13-07 946.00 946.50 0.50 182867 12.7

TMS-13-07 946.50 947.00 0.50 182868 1.09

TMS-13-07 948.00 949.00 1.00 182871 1.92

TMS-13-07 949.00 950.00 1.00 182872 1.37

TMS-13-07 952.00 953.00 1.00 182876 1.89

TMS-13-07 953.00 954.00 1.00 182877 2.72

TMS-13-07 958.00 959.00 1.00 182883 1.67

TMS-13-07 959.00 960.00 1.00 182884 1.12

TMS-13-07 979.00 980.00 1.00 182906 4.31

TMS-13-07 1004.00 1005.00 1.00 182935 1.42

TMS-13-07 1061.00 1061.50 0.50 182999 1.12

TMS-13-07 1062.50 1063.00 0.50 183503 2.96

TMS-13-07 1117.00 1118.00 1.00 183565 2.43

TMS-13-07 1134.00 1135.00 1.00 183584 6.54

TMS-13-07 1135.00 1136.00 1.00 183585 1.44

TMS-13-07 1136.00 1137.00 1.00 183586 1.4

TMS-13-08 368.00 369.00 1.00 183717 1.09

TMS-13-09A 285.50 286.50 1.00 184095 22.69

TMS-13-09A 286.50 287.50 1.00 184096 2.00

TMS-13-09A 312.20 312.70 0.50 184109 1.55

www.innovexplo.com


10.2. No. 3 Mine

Between June and end of December 2013, twelve (12) diamond drill holes were completed in the No. 3 Mine area for a total of 4,784.67 m, with the objective of exploring historical veins No. 1 and No. 2 at depth and confirming historical gold values reported by previous operators. Table 10.3 presents the technical parameters of each drill hole and Figure 10.2 shows the drill hole locations.

Table 10.3 – 2013 technical parameters in the No. 3 Mine area Drill hole # East UTM North UTM Elevation Azimuth Dip Length (m)

M3-13-01 295390.00 5329510.00 320 356.8 -73.5 180.14

M3-13-01A 295412.00 5329493.00 320 353.9 -72.5 89.82

M3-13-02 295550.00 5329490.00 320 355.2 -69.3 482.54

M3-13-03 295190.00 5329510.00 320 353.3 -73.2 84.05

M3-13-03A 295190.00 5329510.00 320 354.3 -70.8 563.13

M3-13-04 295190.00 5329510.00 320 358.1 -53.8 405.72

M3-13-05 295240.00 5329490.00 320 2.9 -69.8 501.00

M3-13-06 295240.00 5329490.00 320 358.5 -55.9 439.68

M3-13-07 295290.00 5329440.00 320 353.2 -50.5 461.23

M3-13-08 295344.00 5329454.00 320 354.8 -65.1 473.85

M3-13-09 295490.00 5329440.00 320 1.4 -49.5 492.76

M3-13-10 295490.00 5329440.00 320 359.9 -62 610.75

www.innovexplo.com

2014NI43‐101TechnicalReportontheLamaquepropertyApril25.2014IntegraGoldCorp. 68

Figure 10.2 – Diamond drilling program in the No. 3 Mine area

www.innovexplo.com


The main lithologies encountered in the No. 3 Mine area were diorite sills, with local occurrences of small feldspar porphyry dykes. Generally chloritized, sericitized and silicified, the units are sheared and fractured. Several mineralized veins with trace to 5% pyrite were logged, with lengths varying between 0.5 cm and 300 cm. The veins are made of quartz-carbonate-tourmaline-chlorite and/or sericite-feldspar assemblages; epidote and calcite are rare. Integra Gold used appropriate QA/QC protocols, employing duplicates, blanks and standards. A total of 1,597 core samples and 217 QA/QC control samples were collected; the total sampled length is 1,545.94 m (32% of total drill hole core length). Table 10.4 below presents the most significant intersections (complete results are available at the office of Integra Gold).

Table 10.4 – Table of significant assay results for the drilling campaign in the No. 3 Mine area

Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t M3-13-01 62.75 63.50 0.75 P243623 2.75

M3-13-01 84.00 85.00 1.00 P243637 2.74

M3-13-01 144.00 145.00 1.00 P243661 10.60

M3-13-01 146.00 146.75 0.75 P243663 3.58

M3-13-01 173.25 174.00 0.75 P243673 2.25

M3-13-01A 87.00 88.00 1.00 P243758 1.695

M3-13-01A 88.00 89.00 1.00 P243759 17.55

M3-13-02 145.00 146.00 1.00 P243833 3.44

M3-13-02 312.00 313.00 1.00 P243905 1.555

M3-13-02 322.00 323.00 1.00 P243908 10.65

M3-13-02 381.00 382.00 1.00 P243946 2.57

M3-13-02 382.00 383.00 1.00 P243947 6.58

M3-13-04 160.00 161.00 1.00 P133016 87.4

M3-13-04 161.00 162.00 1.00 P133017 9.19

M3-13-05 104.00 105.00 1.00 P133151 1.07

M3-13-05 305.00 305.70 0.70 P133227 2.04

M3-13-06 149.00 150.00 1.00 P133327 9.50

M3-13-06 154.00 155.00 1.00 P133333 1.34

M3-13-06 228.00 229.00 1.00 P133363 26.2

M3-13-06 236.00 237.00 1.00 P133366 1.205

M3-13-06 254.00 255.00 1.00 P133379 2.57

M3-13-06 267.50 268.50 1.00 P133385 1.06

www.innovexplo.com


Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t M3-13-06 286.50 287.50 1.00 P133394 5.69

M3-13-06 323.00 324.00 1.00 P133403 1.41

M3-13-06 324.00 325.00 1.00 P133404 1.01

M3-13-07 114.00 115.00 1.00 P133479 4.48

M3-13-07 120.00 121.00 1.00 P133483 4.13

M3-13-07 123.00 124.00 1.00 P133486 1.545

M3-13-07 126.00 127.00 1.00 P133489 1.125

M3-13-07 339.00 340.00 1.00 P133554 21.00

M3-13-08 137.00 138.00 1.00 P133631 1.09

M3-13-08 181.00 182.00 1.00 P133656 1.120

M3-13-08 316.00 317.00 1.00 P133706 1.810

M3-13-08 318.00 319.00 1.00 P133708 18.650

M3-13-08 319.00 320.00 1.00 P133709 1.680

M3-13-08 323.00 324.00 1.00 P133714 1.405

M3-13-08 326.00 327.00 1.00 P133717 1.135

M3-13-08 327.00 328.00 1.00 P133718 2.470

M3-13-08 397.00 398.00 1.00 P133748 3.450

M3-13-08 406.00 407.00 1.00 P133758 1.04

M3-13-08 407.00 408.00 1.00 P133759 4.41

M3-13-08 411.00 412.00 1.00 P133764 1.65

M3-13-09 191.00 192.00 1.00 P133843 1.2

M3-13-09 192.00 193.00 1.00 P133844 23.9

M3-13-09 193.00 194.00 1.00 P133845 15.55

M3-13-10 171.00 172.00 1.00 P124036 1.885

M3-13-10 335.50 336.00 0.50 P124095 97.3

M3-13-10 348.00 349.00 1.00 P124103 1.225

M3-13-10 390.00 390.50 0.50 P124117 2.37

M3-13-10 425.75 426.50 0.75 P124145 5.73

M3-13-10 430.50 431.00 0.50 P124152 3.06

M3-13-10 433.50 434.00 0.50 P124156 53.4

M3-13-10 435.00 435.50 0.50 P124158 1.865

M3-13-10 437.50 438.00 0.50 P124162 2.06

M3-13-10 438.00 438.50 0.50 P124163 6.76

M3-13-10 544.00 544.50 0.50 P124217 2.14

www.innovexplo.com


10.3. Parallel Zone

During the autumn of 2013, forty (40) diamond drill holes were completed in the Parallel Zone with a total of 12,589.14 m drilled, with the objectives of exploring the depth extension of the zone and completing the definition drilling. Table 10.5 presents the technical parameters of each drill hole and Figure 10.3 displays the drill hole locations.

Table 10.5 – 2013 technical parameters on the Parallel Zone Drill Hole

No. East UTM North UTM Elevation Azimuth Dip Length

(m) PV-13-01 294801.48 5329993.25 328.75 331.72 -72.70 613.69

PV-13-02 294901.33 5330096.22 327.36 327.41 -72.31 645.11

PV-13-03 294901.18 5330096.33 327.31 328.22 -65.14 588.00

PV-13-04 294952.48 5329884.93 327.49 325.93 -58.97 667.29

PV-13-05 294792.80 5330162.60 173.00 345.60 -73.10 400.18

PV-13-06 294830.40 5330105.10 123.00 329.20 -68.70 363.37

PV-13-07 295123.00 5330074.80 77.90 342.10 -72.20 462.81

PV-13-08 295012.00 5330062.40 91.80 345.30 -73.20 524.12

PV-13-09 294917.91 5330032.95 327.60 348.95 -66.41 515.62

PV-13-10 294959.47 5330079.20 327.16 345.64 -68.50 530.88

PV-13-11 294985.71 5330100.45 327.43 346.60 -63.31 231.60

PV-13-12 294777.82 5329790.96 328.81 349.98 -65.44 636.39

PV-13-13 294825.78 5329999.44 328.28 350.85 -55.56 256.22

PV-13-14 294898.47 5329994.81 328.18 341.58 -66.64 288.69

PV-13-15 294996.23 5330116.38 327.74 340.66 -53.23 151.62

PV-13-16 295000.00 5330046.01 327.23 348.27 -52.06 253.45

PV-13-17 295000.18 5330020.55 327.43 344.19 -48.50 216.78

PV-13-18 294880.77 5330022.19 327.79 347.44 -64.23 225.21

PV-13-19 294850.04 5330029.13 328.45 347.55 -59.81 210.53

PV-13-20 295002.12 5329964.03 327.28 342.61 -50.87 264.41

PV-13-21 295023.90 5329969.81 327.13 349.89 -52.90 246.71

PV-13-22 295049.43 5329980.03 327.06 345.51 -47.32 318.25

PV-13-23 295020.48 5330120.61 327.63 337.91 -50.11 174.42

PV-13-24 294824.56 5330114.53 332.52 357.27 -50.38 168.27

PV-13-25 294844.64 5330104.49 331.60 346.04 -63.29 198.29

PV-13-26 294872.93 5330097.59 328.31 347.64 -58.14 210.49

PV-13-27 294922.34 5330061.93 327.53 347.29 -53.82 282.54

PV-13-28 295048.53 5330074.62 327.69 337.93 -59.30 231.79

PV-13-29 294935.20 5330037.49 327.39 349.42 -54.36 249.52

PV-13-30 294973.30 5330104.93 327.34 347.13 -51.71 192.58

PV-13-31 295078.07 5330076.42 327.30 346.07 -56.66 234.49

PV-13-32 295075.00 5330065.00 326.80 344.40 -57.80 247.53

www.innovexplo.com


Drill Hole No.

East UTM North UTM Elevation Azimuth Dip Length (m)

PV-13-33 294926.03 5330002.87 327.92 342.28 -65.48 276.17

PV-13-34 295079.43 5329997.18 326.75 345.13 -59.77 52.90

PV-13-35 295097.29 5330009.37 326.83 338.00 -54.50 272.78

PV-13-36 294920.81 5329979.01 327.83 339.08 -65.06 276.11

PV-13-37 295099.99 5329992.68 326.92 353.45 -59.97 269.40

PV-13-38 295125.00 5329965.00 326.40 343.60 -60.00 38.70

PV-13-38A 295124.20 5329972.23 326.40 347.59 -59.93 284.00

PV-13-39 294973.68 5329960.22 327.71 344.03 -52.92 318.23

www.innovexplo.com


Figure 10.3 – 2013 diamond drilling program on the Parallel Zone

www.innovexplo.com

2014NI43‐101TechnicalReportontheLamaquepropertyApril25,2014IntegraGoldCorp 74

The drill holes were completed to verify the lateral and depth extensions of some of the best results obtained in 2010-2011. Drill holes intersected gabbro-diorite intrusive units with alternating metre-scale bands of felsic to intermediate volcanics of rhyolitic to dacitic composition. Locally silicification and sericitization are present. Shearing is observed within the gabbro-diorite accompanied with intense chloritization, local fracturing, brecciation and pyritization (pyrite content from 2 to 3%; locally to 10-15%). Locally, pyrrhotite is present with similar proportions as the pyrite content. The gabbro-diorite host is weakly to moderately magnetic. Several quartz-carbonate veins were sampled, of the thickest being 15 to 20 cm. Integra Gold used appropriate QA/QC protocols employing duplicates, blanks and standards. A total of 5,438 core samples and 749 QA/QC control samples were collected; the total sampled length is 5,364.52 m (43% of total drill hole core length). The Table 10.6 below presents the most significant intersections (complete results are available at the office of Integra Gold).

Table 10.6 – Table of significant assay results for the 2013 drilling campaign on the Parallel Zone

Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t PV-13-01 425.00 426.00 1.00 P124407 3.24

PV-13-01 510.00 511.00 1.00 185012 10.43

PV-13-01 511.00 512.00 1.00 185013 4.47

PV-13-01 513.00 514.00 1.00 185015 2.39

PV-13-01 572.00 573.00 1.00 185253 1.95

PV-13-01 588.00 589.00 1.00 185053 9.41

PV-13-02 80.00 81.00 1.00 185088 4.17

PV-13-02 83.00 84.00 1.00 185092 1.74

PV-13-02 84.00 85.00 1.00 185093 1.72

PV-13-02 85.00 86.00 1.00 185094 4.36

PV-13-02 86.00 87.00 1.00 185095 3.04

PV-13-02 88.00 89.00 1.00 185097 1.51

PV-13-02 145.00 146.00 1.00 185112 1.59

PV-13-02 147.00 148.00 1.00 185114 13.62

PV-13-02 430.00 431.00 1.00 185212 4.36

PV-13-02 474.50 475.50 1.00 185223 3.77

PV-13-02 484.50 485.50 1.00 185229 11.14

PV-13-02 487.00 488.00 1.00 185232 1.45

PV-13-02 534.00 535.00 1.00 185244 2.8

PV-13-02 551.00 552.00 1.00 185252 2.12

PV-13-03 85.00 86.00 1.00 185335 2.12

PV-13-03 94.00 95.00 1.00 185342 1.11

www.innovexplo.com


Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t PV-13-03 106.00 107.00 1.00 185348 1.53

PV-13-03 148.00 149.00 1.00 185369 1.8

PV-13-03 149.00 150.00 1.00 185371 24.31

PV-13-03 153.00 154.00 1.00 185376 3.6

PV-13-03 523.00 524.00 1.00 185507 5.9

PV-13-03 524.00 525.00 1.00 185508 2.41

PV-13-04 207.00 208.00 1.00 184618 1.58

PV-13-04 495.00 496.00 1.00 184702 1.06

PV-13-04 519.00 520.00 1.00 184708 1.51

PV-13-04 573.00 574.00 1.00 184721 1.21

PV-13-05 166.00 167.00 1.00 184367 1.00

PV-13-05 187.50 188.00 0.50 184389 379.75

PV-13-05 188.50 189.50 1.00 184392 1.79

PV-13-05 296.00 297.00 1.00 184422 1.42

PV-13-05 297.00 297.70 0.70 184423 2.11

PV-13-05 522.00 523.00 1.00 183774 5.87

PV-13-06 246.00 247.00 1.00 183814 2.15

PV-13-06 472.00 472.50 0.50 183947 5.3

PV-13-06 512.50 513.00 0.50 183964 4.63

PV-13-06 521.50 522.00 0.50 183978 6.66

PV-13-06 532.70 533.20 0.50 183985 4.57

PV-13-06 552.00 552.50 0.50 183998 2.22

PV-13-06 563.00 564.00 1.00 185512 1.55

PV-13-06 571.00 571.80 0.80 185522 2.58

PV-13-07 292.00 293.00 1.00 185552 2.74

PV-13-07 293.00 294.00 1.00 185553 1.18

PV-13-07 298.50 299.00 0.50 185555 8.51

PV-13-07 299.50 300.00 0.50 185557 1.44

PV-13-07 304.70 305.30 0.60 185561 9.27

PV-13-07 305.30 305.80 0.50 185562 80.90

PV-13-07 334.50 335.00 0.50 185577 5.68

PV-13-07 336.00 336.50 0.50 185579 4.45

PV-13-07 343.00 343.50 0.50 185585 4.44

PV-13-07 345.00 345.50 0.50 185588 9.34

PV-13-07 348.50 349.00 0.50 185593 2.90

www.innovexplo.com


Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t PV-13-07 352.00 352.50 0.50 185597 4.32

PV-13-07 373.00 373.50 0.50 185601 20.00

PV-13-07 374.50 375.00 0.50 185603 1.54

PV-13-07 418.00 419.00 1.00 185623 1.25

PV-13-07 424.40 425.00 0.60 185631 8.73

PV-13-07 441.00 442.00 1.00 185645 1.08

PV-13-07 459.00 460.00 1.00 185655 1.04

PV-13-08 448.00 449.00 1.00 184818 2.84

PV-13-08 449.00 450.00 1.00 184819 5.96

PV-13-08 749.00 750.00 1.00 184933 1.63

PV-13-09 24.00 25.00 1.00 185846 5.59

PV-13-09 40.00 41.00 1.00 185859 1.06

PV-13-09 145.00 146.00 1.00 185913 71.47

PV-13-09 158.00 159.00 1.00 185923 9.61

PV-13-09 189.00 190.00 1.00 185932 3.63

PV-13-09 485.00 486.00 1.00 186568 1.24

PV-13-10 151.00 152.00 1.00 186029 1.37

PV-13-10 171.00 172.00 1.00 186041 2.97

PV-13-10 181.00 182.00 1.00 186047 8.85

PV-13-10 431.00 432.00 1.00 186172 1.4

PV-13-11 61.00 62.00 1.00 186237 1.07

PV-13-11 79.00 80.00 1.00 186247 3.07

PV-13-11 157.00 158.00 1.00 186289 7.88

PV-13-11 158.00 159.00 1.00 186291 21.93

PV-13-11 159.00 160.00 1.00 186292 1.73

PV-13-11 160.00 161.00 1.00 186293 1.38

PV-13-11 161.00 162.00 1.00 186294 45.16

PV-13-11 162.00 163.00 1.00 186295 4.83

PV-13-11 163.00 164.00 1.00 186296 8.38

PV-13-11 164.00 165.00 1.00 186297 1.78

PV-13-11 165.00 166.00 1.00 186298 8.07

www.innovexplo.com


Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t PV-13-12 33.00 34.00 1.00 186597 1.11

PV-13-13 129.00 130.00 1.00 186915 1.11

PV-13-13 155.00 156.00 1.00 186935 3.17

PV-13-14 90.00 91.00 1.00 187066 1.02

PV-13-14 246.00 247.00 1.00 187182 1.64

PV-13-14 267.00 268.00 1.00 187196 2.34

PV-13-15 23.00 24.00 1.00 186321 2.3

PV-13-15 24.00 25.00 1.00 186322 1.83

PV-13-15 29.00 30.00 1.00 186326 1.13

PV-13-15 39.00 40.00 1.00 186331 1.03

PV-13-15 85.00 86.00 1.00 186362 1.24

PV-13-15 86.00 87.00 1.00 186363 3.41

PV-13-15 97.00 98.00 1.00 186365 1.53

PV-13-15 114.00 115.00 1.00 186377 7.85

PV-13-15 115.00 116.00 1.00 186378 4.85

PV-13-15 119.00 120.00 1.00 186379 1.07

PV-13-15 129.00 130.00 1.00 186384 3.65

PV-13-15 134.00 135.00 1.00 186385 1.76

PV-13-15 135.00 136.00 1.00 186386 3.03

PV-13-15 138.00 139.00 1.00 186387 6.9

PV-13-16 182.00 183.00 1.00 188003 2.7

PV-13-16 183.00 184.00 1.00 188004 5.63

PV-13-16 184.00 185.00 1.00 188005 14.57

PV-13-16 185.00 186.00 1.00 188006 2.82

PV-13-16 186.00 187.00 1.00 188007 10.97

PV-13-16 187.00 188.00 1.00 188008 9.3

PV-13-17 58.00 59.00 1.00 188055 1.61

PV-13-17 62.00 63.00 1.00 188057 1.88

PV-13-17 70.00 71.00 1.00 188065 1

PV-13-17 119.00 120.00 1.00 188076 2.48

PV-13-17 193.00 194.00 1.00 188117 1.33

PV-13-18 165.00 166.00 1.00 187278 15.26

PV-13-18 192.20 193.00 0.80 187301 1.37

www.innovexplo.com


Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t PV-13-18 202.00 203.00 1.00 187304 2.59

PV-13-19 110.00 111.00 1.00 187361 2.77

PV-13-19 111.00 112.00 1.00 187362 1.14

PV-13-19 112.00 113.00 1.00 187363 1.58

PV-13-19 192.00 193.00 1.00 187407 15.72

PV-13-19 203.00 204.00 1.00 187419 3.66

PV-13-20 166.00 167.00 1.00 188356 17.41

PV-13-20 219.00 220.00 1.00 188385 3.85

PV-13-20 228.00 229.00 1.00 188388 13.16

PV-13-20 239.00 240.00 1.00 188394 77.69

PV-13-20 240.00 241.00 1.00 188395 1.35

PV-13-20 241.00 242.00 1.00 188396 4.15

PV-13-20 242.00 243.00 1.00 188397 11.46

PV-13-21 119.00 120.00 1.00 188197 4.31

PV-13-21 151.00 152.00 1.00 188212 1.45

PV-13-21 186.00 187.00 1.00 188237 1.41

PV-13-21 207.00 208.00 1.00 188251 3.33

PV-13-21 208.00 209.00 1.00 188252 4.97

PV-13-21 219.00 220.00 1.00 188256 84.51

PV-13-21 220.00 221.00 1.00 188257 3.37

PV-13-21 221.00 222.00 1.00 188258 1.18

PV-13-21 232.00 233.00 1.00 188263 4.34

PV-13-21 233.00 234.00 1.00 188264 1.57

PV-13-21 234.00 235.00 1.00 188265 5.85

PV-13-21 235.00 236.00 1.00 188266 1.61

PV-13-22 103.00 104.00 1.00 188453 1.13

PV-13-22 117.50 118.50 1.00 188462 3.19

PV-13-22 151.00 152.00 1.00 188476 23.94

PV-13-22 224.00 225.00 1.00 188503 2.22

PV-13-22 225.00 226.00 1.00 188504 51.26

PV-13-22 226.00 227.00 1.00 188505 3.62

PV-13-22 227.00 228.00 1.00 188506 3.54

PV-13-22 231.00 232.00 1.00 188511 25.79

PV-13-22 232.00 233.00 1.00 188512 8.17

PV-13-22 237.00 238.00 1.00 188513 1.01

www.innovexplo.com


Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t

PV-13-23 96.00 97.00 1.00 188578 2.6

PV-13-23 97.00 98.00 1.00 188579 1.86

PV-13-23 108.00 109.00 1.00 188592 1.53

PV-13-25 126.00 127.00 1.00 189038 20.95

PV-13-25 127.00 128.00 1.00 189039 49.45

PV-13-25 162.00 163.00 1.00 189046 4.18

PV-13-25 165.00 166.00 1.00 189049 1.14

PV-13-25 167.00 168.00 1.00 189052 5.84

PV-13-25 185.00 186.00 1.00 189065 1.25

PV-13-25 188.00 189.00 1.00 189068 10.36

PV-13-26 31.00 32.00 1.00 189086 2.41

PV-13-26 67.00 68.00 1.00 189115 1.28

PV-13-26 68.00 69.00 1.00 189116 2.24

PV-13-26 94.00 95.00 1.00 189137 1.01

PV-13-26 117.00 118.00 1.00 189145 2.82

PV-13-26 133.00 134.00 1.00 189151 19.72

PV-13-26 148.00 149.00 1.00 189161 4.8

PV-13-26 149.00 150.00 1.00 189162 5.92

PV-13-26 159.00 160.00 1.00 189173 73.75

PV-13-26 182.00 183.00 1.00 189186 12.46

PV-13-27 72.00 73.00 1.00 189523 18.59

PV-13-27 73.00 74.00 1.00 189524 2.66

PV-13-27 108.00 109.00 1.00 189542 2.08

PV-13-28 26.00 27.00 1.00 188631 1.76

PV-13-28 79.00 80.00 1.00 188667 5.16

PV-13-28 80.00 81.00 1.00 188668 1.4

PV-13-28 83.00 84.00 1.00 188669 91.03

PV-13-28 87.00 88.00 1.00 188674 2.56

PV-13-29 62.00 63.00 1.00 190061 8.61

PV-13-29 63.00 64.00 1.00 190062 7.96

PV-13-29 143.00 144.00 1.00 190107 3.17

PV-13-29 145.00 146.00 1.00 190109 1.27

PV-13-29 146.00 147.00 1.00 190111 1.3

www.innovexplo.com


Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t PV-13-29 147.00 148.00 1.00 190112 16.6

PV-13-29 148.00 149.00 1.00 190113 3.84

PV-13-29 171.00 172.00 1.00 190135 4.92

PV-13-29 172.00 173.00 1.00 190136 8.62

PV-13-29 173.00 174.00 1.00 190137 7.61

PV-13-29 183.00 184.00 1.00 190144 2.07

PV-13-29 184.00 185.00 1.00 190145 11.92

PV-13-29 245.00 246.00 1.00 190193 4.19

PV-13-30 30.00 31.00 1.00 189215 1.67

PV-13-30 88.00 89.00 1.00 189239 1.27

PV-13-30 89.00 90.00 1.00 189241 1.15

PV-13-30 134.00 135.00 1.00 189265 1.18

PV-13-30 135.00 136.00 1.00 189266 2.07

PV-13-30 143.00 144.00 1.00 189273 1.11

PV-13-30 146.00 147.00 1.00 189277 3.13

PV-13-30 147.00 148.00 1.00 189278 2.27

PV-13-30 154.00 155.00 1.00 189282 1.11

PV-13-30 155.00 156.00 1.00 189283 3.24

PV-13-30 156.00 157.00 1.00 189284 20.13

PV-13-31 71.00 72.00 1.00 188926 2.68

PV-13-31 88.00 89.00 1.00 188939 2.4

PV-13-32 73.00 74.00 1.00 188792 4.95

PV-13-32 74.00 75.00 1.00 188793 4.34

PV-13-32 96.00 97.00 1.00 188808 1.5

PV-13-33 39.00 40.00 1.00 189626 14.93

PV-13-33 43.00 44.00 1.00 189631 1.78

PV-13-33 44.00 45.00 1.00 189632 2.02

PV-13-33 91.00 92.00 1.00 189666 3.63

PV-13-33 99.00 100.00 1.00 189671 4.09

PV-13-33 117.00 118.00 1.00 189678 1.37

PV-13-33 118.00 119.00 1.00 189679 1.01

PV-13-33 119.00 120.00 1.00 189681 1.01

PV-13-33 147.50 148.50 1.00 189692 1.23

www.innovexplo.com


Drill Hole #

From (m)

To (m)

Length (m)

Sample #

Au

g/t PV-13-33 157.00 158.00 1.00 189695 3.2

PV-13-33 222.50 223.00 0.50 189732 1.7

PV-13-33 224.00 225.00 1.00 189734 2.48

PV-13-33 229.00 230.00 1.00 189739 1.23

PV-13-33 232.00 233.00 1.00 189743 2.23

PV-13-35 87.50 88.50 1.00 189785 1.27

PV-13-35 105.00 106.00 1.00 189797 1.82

PV-13-35 112.00 113.00 1.00 189801 3.19

PV-13-35 116.00 117.00 1.00 189805 3.6

PV-13-35 157.00 158.00 1.00 189834 6.2

PV-13-35 158.00 159.00 1.00 189835 9.47

PV-13-35 184.80 185.80 1.00 189857 2.74

PV-13-35 185.80 187.00 1.20 189858 2.3

PV-13-35 187.00 188.00 1.00 189859 5.85

PV-13-35 208.00 209.00 1.00 189864 7.47

PV-13-35 209.00 210.00 1.00 189865 1.4

PV-13-36 106.00 107.00 1.00 189921 1.48

PV-13-36 151.00 152.00 1.00 189936 1.43

PV-13-37 45.00 46.00 1.00 190606 2.84

PV-13-37 89.00 90.00 1.00 190635 1.52

PV-13-37 222.00 222.80 0.80 190704 2.56

PV-13-37 224.00 225.00 1.00 190706 2.64

PV-13-38A 64.00 65.00 1.00 189329 2.73

PV-13-38A 79.00 80.00 1.00 189341 2.95

PV-13-38A 228.00 229.00 1.00 190248 4.99

PV-13-38A 231.00 232.00 1.00 190252 1.97

PV-13-38A 255.00 256.00 1.00 190269 3.82

PV-13-39 243.00 244.00 1.00 190572 1.47

PV-13-39 245.00 246.00 1.00 190574 2.28

PV-13-39 246.00 247.00 1.00 190576 14.67

www.innovexplo.com


10.4. No. 6 Vein And Sixteen Zone

In 2010, 2011 and 2012, Integra Gold completed fourteen (14) diamond drill holes on the Sixteen Zone totalling 2,816 m, and sixteen (16) diamond drill holes on the No. 6 Vein for a total length of 3,979.51 m. GéoPointCom used the drilling programs in their resource calculations (November and December 2013; see Items 14.1 and 14.2). Tables 10.7 and 10.8 respectively list the technical parameters of the drill holes for those zones and Figures 10.4 and 10.5 display the drill hole locations. All details for these zones are described in the respective NI 43-101 compliant technical reports dated June 23, 2011 (amended September 21, 2012) and November 1, 2013), both of which are available on SEDAR.

Table 10.7 – 2010 technical parameters on the Sixteen Zone Drill Hole # East UTM North UTM Elevation Azimuth Dip Length(m)

SX-04-18-EXT 292 366.30 5 329 609.79 324.11 112.3 -85 51.00

SX-04-19-EXT 292 393.38 5 329 608.97 323.70 93.8 -84 80.36

SX-04-31-EXT 292 364.62 5 329 625.91 324.12 73.9 -84 51.00

SX-10-01 292 312.00 5 329 564.50 323.00 329 -70 276.00

SX-10-02 292 280.00 5 329 610.00 320.00 78 -65 303.00

SX-10-02 292 280.00 5 329 610.00 320.00 85 -65 303.00

SX-10-03 292 363.00 5 329 595.00 320.00 95 -55 150.00

SX-10-04 292 398.00 5 329 556.00 320.00 341 -70 282.00

SX-10-04 292 398.00 5 329 556.00 323.00 341 -68 282.00

SX-10-05 292 451.00 5 329 600.00 320.00 43 -82 102.00

SX-10-06 292 377.00 5 329 717.00 320.00 163 -47 201.00

SX-10-07 292 540.00 5 329 745.00 320.00 159.8 -55 252.00

SX-10-08 292 631.00 5 329 790.00 320.00 160 -55 252.00

SX-10-09 292 647.00 5 329 585.00 320.00 334 -55 252.00

SX-10-10 292 340.00 5 329 570.00 320.00 344 -70 282.00

SX-10-11 292 385.00 5 329 555.00 320.00 341 -70 282.00

Table 10.8 – 2012 technical parameters on the No. 6 Vein Drill Hole # East UTM North UTM Elevation Azimuth Dip Length (m)

V6-12-01 293465.00 5330061.00 320.00 359.7 -52 300.00

V6-12-02 293465.00 5330061.00 320.00 358.8 -60 318.00

V6-12-03 293265.00 5330061.00 320.00 357.1 -52 471.00

V6-12-04 293265.00 5330061.00 320.00 359.5 -64 305.92

V6-12-05 293515.00 5330061.00 320.00 4.5 -52 250.00

V6-12-06 293515.00 5330061.00 320.00 359.4 -59 306.06

V6-12-07 293320.00 5330004.00 320.00 358.6 -54 330.00

V6-12-08 293320.00 5330004.00 320.00 352.7 -63 32.22

V6-12-08A 293320.00 5330004.00 320.00 351.4 -67 30.00

www.innovexplo.com


Drill Hole # East UTM North UTM Elevation Azimuth Dip Length (m)

V6-12-08B 293320.00 5330004.00 320.00 3.9 -64 381.00

V6-12-09 293420.00 5330010.00 320.00 8.2 -56 24.05

V6-12-09A 293420.00 5330010.00 320.00 355.2 -50 369.00

V6-12-10 293420.00 5330010.00 320.00 6.5 -64 258.93

V6-12-11 293320.00 5330004.00 320.00 0.9 -73 432.49

V6-12-12 293320.00 5329985.00 320.00 8.0 -74 21.00

V6-12-12A 293320.00 5329985.00 320.00 2.9 -71 149.84

www.innovexplo.com


Figure 10.4 – 2010 diamond drilling program on the Sixteen Zone

www.innovexplo.com


Figure 10.5 – Diamond Drilling Program 2012 on No. 6 Vein

www.innovexplo.com


The host rocks of the No.6 Vein mainly consist of chloritized intermediate to mafic tuffs and diorite. The diorite is silicified and locally sericitized and/or carbonatized. Shearing affects all rock types with, very locally, fractures and/or brecciation. Several mineralized veins range in thickness from 0.5 cm to 90 cm, with trace to 7% pyrite (locally up to 10%), trace to 1% pyrrhotite, and locally chalcopyrite. A few of the veins contain trace amounts of visible gold. The veins consist of quartz-carbonate-tourmaline-chlorite and/or sericite-feldspar (see all details for these drill holes presented in the NI 43-101 Technical Report filed on SEDAR in November 2013).

10.5. Additional Resource Potential

As of February 28, 2014 (effective date of the PEA), Integra Gold Corp had drilled an additional 39,235 m in 104 drill holes on the Lamaque Project which has not yet been included in any resource estimate calculation used in the PEA. Additional drilling that is not included in the estimate consists of:

Mine No. 3 4,785 m in 12 holes No. 5 Plug 4,761 m in 11 holes Fortune 0 m South Triangle: 6,966 m in 13 holes Parallel: 12,833 m in 42 holes Triangle: 9,880 m in 26 holes

The 2013 South Triangle drill program aimed at extending the Triangle Zone to the south and at depth. Drilling intersected high grade mineralization to a vertical depth of over 1,000 m as well as approximately 175 m downdip of the southern extent of the Triangle Zone resource limit. Drilling at the Parallel and Triangle Zones that is not included in the resources estimate consists of both infill and extension drilling. The results for the Parallel Zone were disclosed in early 2014 and confirm the continuity of the mineralized zones, which the Issuer hopes will assist in converting further ounces from the Inferred to Indicated category. The infill and extension drilling at the Triangle Zone started in January 2014.

www.innovexplo.com


11. SAMPLE PREPARATION, ANALYSES AND SECURITY

No information is available concerning sampling protocols for Teck’s programs prior to the joint ventures with Golden Pond and Tundra in 1985 and 1986, and the joint ventures with Tundra and Golden Pond between 1985 and 1989. From the drill logs for the E-series holes which were reviewed, samples varied in length from 1 to 3 ft (30 to 91 cm). Sample numbers are reported in the drill logs and assay reports listing; the sample numbers and assay results are appended to the logs. The header on the assay reports is Lamaque Mining Company Limited, indicating that the issuer was the company assay laboratory at the mine site. The assay reports indicate that many of the samples were assayed in duplicate. Geologica does not know if any of the core has survived. No records of surface sampling are known to Geologica. Geologica is not aware of any written description of sampling protocol for either of these two joint venture programs operated by Teck, but they were likely based on Teck’s protocols long established at the Mine. The drill logs reviewed by Geologica indicate that selected core samples varied in length from 1 ft to 4 ft (30 to 122 cm) and many were 3 ft (91 cm) long suggesting maximum sample lengths might have been 5 ft (152 cm). The logs and our observation of surviving core indicate that both veins and altered intervals peripheral to veins were sampled. For the drilling program carried out from 2003 to 2008 by Kalahari (now Integra Gold), core intervals for sampling were selected during the core logging procedure by the logging personnel. Terrence Coyle, P.Geo., logged almost all of the Kalahari 2003-2008 drill holes. Don Cross logged the first few holes. Sample selection was based on evidence of mineralization such as quartz-tourmaline veining, the presence of sulphides or alteration. Sample lengths varied from a minimum of 1 ft to a maximum length of 3 ft long (30 to 91 cm). Sample demarcation decisions were made on the basis of geology and for most part vein material was segregated from altered wall rock, which was segregated from apparently non-mineralized host rock. Core recovery is considered to have been close to 100%. The logging geologist marked the intervals to be sampled on the core using a lumber crayon and recorded the sampled intervals in the drill core logs. The intervals selected for sampling were sawn in half by a sampler using a diamond core saw. After cutting, one half of the sample was returned to the core tray in its original order to serve as an archive, while the other half was placed into a plastic sample bag with a numbered sample ticket. The same sample numbers were inscribed on the bag with a permanent marker. Three part sample tickets were used for sampling. One part of the ticket was placed in the sample bag, one part was placed in the core tray, secured under the first piece of archived core marking the beginning of the sampled interval and the third part was left in the sample book. The sample bags were secured with nylon pull locks. The samples were transported by Kalahari personnel to the Bourlamaque Assay Laboratory in Val-d’Or, Québec. The relationship between sample lengths and true thickness is highly variable and difficult to quantify. It is dependent on the orientation of the drill holes and the orientation of the vein structures, both strike and dip of which are highly variable. Mineralization in the Lamaque area as described under the Mineralization section of the report comprises a variety of structures of diverse orientation and it can often be difficult determining which type of vein structure and orientation is being intersected.

www.innovexplo.com


During logging, vein orientations are measured with respect to the core and recorded in the core log but even with such measurement of core angles, vein orientation and the relationships between sample length and vein structures can be problematic. The reader should also be aware that it is often vein systems rather than single veins that constitute mineralized zones and sample lengths for individual samples, even given known vein orientations often have little to do with the true widths of the mineralized zones. The recent 2009-2013 drilling program was planned, logged, validated and supervised by Geologica at the Company’s facilities in Val-d’Or, Québec (a core shack with logging stations, sampling, splitting with core saws and shipping to the laboratories). The archived core from Kalahari’s 2003-2008 and Kalahari/Integra Gold’s 2009-2013 programs has been cross piled and stored at the Company’s core shack facilities in Val-d’Or, Québec. Assay reports, appended to the drill logs, indicate that the pre-1985 samples were assayed at the mine assay laboratory still standing at the Main Mine-Mill site. Geologica is not aware of any record of crushing and grinding parameters or sub-sample size. It is believed likely that the assays were gravimetric fire assays using 1 assay ton (29.179 g) sub-samples. Records show that samples were often assayed in duplicate. Similar to that for the pre-joint venture Teck programs, Geologica is not aware of any description of laboratory procedures used for assaying the Tundra and Golden Pond JV program samples. Assay reports appended to the drill logs indicate the principal assay lab was the company’s facility at the mine site but also some check assaying was done at the independently owned, Bourlamaque Assay Laboratories Ltd. lab in Val-d’Or, Québec. The drill logs record the samples that returned assays above trace were re-assayed several times. It is not recorded whether these re-assays were from original pulps or a new pulp cut from the sample reject. It is unknown what kind of, or extent of, QA/QC programs were run in conjunction with the regular sample as saying. Geologica believes the Bourlamaque Assay Laboratories Ltd. lab, being a commercial laboratory, would have used standards to assure quality. The mine laboratory may also have used international standards but certainly would have used a set of in-house standards and re-assays at the Bourlamaque lab to help maintain high quality assays. One problem area concerning mine laboratories is with respect to mixing in the sample stream of mine samples that have a high potential to contain ore-grade values with exploration samples that rarely contain ore-grade values. Cross contamination problems can be aggravated by the lab failing to use clean sand or blank samples to clean the crushers prior to the introduction of differing sample batches and the accumulation of holdover gold, which can contaminate succeeding samples. As far as Geologica is aware, no other samples were being assayed at the lab at the mine during the period the joint venture programs were being operated. Geologica does not know, however, what the policy at the lab was with respect to crusher washing and high grade samples.

www.innovexplo.com


All samples from Kalahari’s 2003 to 2008 programs were assayed at the Bourlamaque Assay Laboratories Ltd. assay lab in Val-d’Or, Québec. All the analytical work completed at the Bourlamaque Laboratory was supervised by L.D. Melnbardis, B.Sc., licensed chemist, Order of Chemists of Québec. The Bourlamaque assay lab is a non-accredited facility but it participates in reference material certification programs, extensive round robin studies and the Proficiency Testing Program for Mineral Analysis Laboratories through Natural Resources Canada, CANMET Mineral Technology Branch. The standard protocol for Kalahari samples included the lead fire assay of a one assay ton subsample and a gravimetric finish. Sample preparation, which follows sample drying, includes primary (jaw) and secondary (cone) crushing of the entire sample to better than 60% -10 mesh (1.70 mm). A 250 g sub-sample is then split off from the crushed material using a 3/8" aperture Jones Riffler. In the final sample preparation step the 250 g sub-sample is ring pulverized to 95% -150 meshes (106 μm) and homogenized. The protocol requires that one or two samples of barren rock be used to clean the crushers between each sample batch, and the pulverisers be cleaned between each sample with a brush and jet of compressed air. The protocol requires that duplicate samples be prepared at intervals of ten run-of-the-stream samples and be assayed. Inspection of Bourlamaque assay certificates shows that for every 10th sample, a duplicate assay was made on the original pulp and another assay was made of a second pulp cut from the sample reject. Standard practice at Bourlamaque also included the re-assay of samples that returned an initial assay greater than a threshold of 0.20 oz Au/ton. These re-assays also entailed a second assay on a one assay ton charge from the original pulp and an assay on a new pulp cut from the archived sample reject. No standards were used as part of the assay protocol. Kalahari did not have a formal QA/QC program but carried out specific check sampling programs between 2003 and 2008 to help assure and assess assay quality. One step included the re-assay at Bourlamaque of a set of 45 samples by metallic screen procedures. These samples had been previously assayed at Bourlamaque using its standard fire assay on one assay ton subsample procedure. The procedure included the screening at 150 mesh (106 μm) of a nominal 250 g pulverized sample with duplicate one assay ton gravimetric fire assays on the undersize and fire assay of the entire oversize fraction. Kalahari undertook this re-assay by metallic screen procedure to compare the assays returned by the standard procedure with respect to the metallic screen procedure. For drilling programs carried out by Kalahari (2009-2010) and Integra Gold (2011-2013), the sampling completed by Geologica was assayed by Bourlamaque Assay Laboratories Ltd. and ALS Chemex Laboratory in Val-d’Or, Québec. Procedures for routine fire assaying are to initially crush the entire sample to –10 mesh, then a 300 g sub-sample is split and pulverized to 95% - 150 mesh, and a 30 g sub-sample is fire assayed using standard industry procedures, with the gold content determined by atomic absorption spectrometry. For the Bourlamaque Assay Laboratories Ltd., each sample was assayed by Fire Assay and AA Finish, when

www.innovexplo.com


values reporting ≥ 10 ppm Au are repeated by Fire Assay with a gravimetric Finish upper reporting limit of 100 g/t Au. For security and quality control, diamond drill core samples were catalogued on sample shipment memos, which were completed at the time samples were being packed for shipment. Duplicates and blanks were taken and the partial core was photographed by geologists. The splitting of samples and sample preparation for shipping were completed by C-Lab and Integra Gold’s technicians under the supervision of Geologica. The material used for standards comprised certified reference material purchased from commercial facilities specializing in their manufacture (RockLabs via Anachemia in Ontario and CND “Analytical Solution Limited”). All material used for blank samples consisted of barren limestone. Laboratories also added their own quality control standards. In case of any doubt regarding the validity of a sample, the entire batch was re-assayed. The authors believe that the sample preparation, security and analytical procedures were correctly applied. Results obtained by the laboratory are representative of the mineralization in comparison with results obtained in the past for all mineralized zones on the Lamaque Property.

11.1. Results of Quality Control

11.1.1. Blanks

The field blank used in the 2013 drilling programs (South Triangle, No. 3 Mine and Parallel Zone) is from a gold-barren sample of crushed white marble. One field blank is inserted for every 20 samples, alternating with standards. Geologica recommends a quality control protocol stipulating that if any blank yields a gold value above 20 ppb Au, the batch of sample containing the blank should be re-assayed. For the 2013 drilling program, no batch was re-assayed.

11.1.2. Certified Reference Material (Standards)

One certified reference material (CRM or “standard”) was inserted for every 20 samples, alternating with blanks, during the drilling programs. Four standards were used, with gold grades ranging from 1.348 g/t Au to 8.595 g/t Au as follows:

SH65 with a theoretical value of 1.348 ± 0.028 g/t Au SK62 with a theoretical value of 4.075 ± 0.473 g/t Au SL61 with a theoretical value of 5.931 ± 0.177 g/t Au SN60 with a theoretical value of 8.595 ± 0.223 g/t Au

Geologica’s quality control protocol stipulates that if any analyzed standard yields a gold value above or below three times the standard deviation (±3SD) of the certified grade for that standard, then ten samples before and after the standard in the batch should be re-analyzed. After reviewing the results, Geologica personnel re-assayed 4 samples in accordance with this criterion, which represented less 0.01% of the database samples for 2013 for Triangle and South Triangle Zones. The re-assaying was done using pulps and rejects, or quarter-split core when no pulp or reject was available. All

www.innovexplo.com


samples accompanying the standard (bearing the unacceptable gold result) during the fusion process were also re-assayed. The first gold values (the “wrong” values) were replaced in the database by the new gold values obtained by the re-assay.

11.1.3. Duplicates

The quality control protocol requires a coarse duplicate be prepared for one sample selected among every 30 samples. The duplicate is prepared by taking half of the crushed material derived from the original sample. By measuring the precision of the coarse duplicates, the incremental loss of precision can be determined for the coarse crush stage of the process, thus indicating whether two sub-samples taken after primary crushing is adequate for the crushed particle size to ensure a representative sub-split. Duplicates are used to check the representativeness of results obtained for a given population. To determine reproducibility, precision (as a percentage) is calculated according to the following formula:

Precision (%) =

(Duplicate Sample Gold Grade – Original Sample Gold Grade)

X 100 Average Between Duplicate Sample Gold Grade and Original Sample Gold Grade

Precision ranges from 0 to 200%, with the best being 0%, meaning that both the original and the duplicate sample returned the same grade. A total of 143 original–coarse duplicate pairs were identified in the database. The correlation coefficient (%) is given by square root of R² and represents the degree scatter of data around the linear regression slope. The results obtained indicate an excellent reproducibility of gold values for 2013 with 80% and 100%.

www.innovexplo.com


12. DATA VERIFICATION

The authors reviewed and verified the existing data of all available past and recent reports. According to elements reported in the statutory documents, sampling works and the analysis thereof seem to have been made according to standards in force at that time and still valid today, although a procedure and method are not described. Duplicates, blanks and standards samples were taken every 20 samples, on average. The authors believe that the sample preparation, security and analytical procedures were correctly applied and correspond with the present standards applied in the mining industry.

www.innovexplo.com


13. MINERAL PROCESSING AND METALLURGICAL TESTING

Three series of laboratory testwork were carried out with samples of the Lamaque deposit in 2012 and 2013 by ALS Metallurgy Kamloops (Roulston D., Johnston H., Shouldice T., Metallurgical Testwork on the Lamaque Deposit, April 3, 2013, May 29, 2013 and January 16, 2014). A summary of the results obtained is presented in this section. The samples used for this testwork were prepared by Integra Gold and WSP cannot determine if those samples were representative of the deposit.

13.1. Initial Testwork

Between winter 2012 and spring 2013, ALS Metallurgy Kamloops carried out two sets of analysis on samples from the Lamaque Project. The deposit consists of four independent zones named: No. 4 Plug, Triangle, Parallel and Fortune. Both reports, KM3569 of April 3, 2013 and KM3876 of May 29, 2013, present results of this initial work, which included:

Ball mill work index ( Bond method); o Performed on the Master Composite and two (2) Cluster Composites.

Determination of mineralogical - chemical composition (QEMSCAN method). o Performed on each of the six (6) composite samples.

Some exploratory work was performed to assess the metallurgical response of Lamaque samples to:

Gravity concentration ( Knelson and hand panning ); Cyanide leach (standard bottle roll procedure); Rougher flotation and regrind.

13.1.1. Samples Description

The samples arrived at the ALS Metallurgy facilities on November 5, 2012. The samples weighed approximately 247 kilograms and were crushed at < 6 mesh. They were split into 6 composites predetermined by the client, based on sample origin and gold grade. They were identified as follows in Table 13.1:

www.innovexplo.com


Table 13.1 – Sample Origin and Weight Location Sample Names Sample Origin Weight (kg)

South Cluster 1 – High Grade

86% No. 4 Plug 14% Triangle

41.4

Cluster 1 – Averag 70% No. 4 Plug 30% Triangle

44.4

Cluster 1 – Cut‐ofGrade

90% No. 4 Plug 10% Triangle

47.6

North Cluster 2 – High Grade

63% Parallel 37% Fortune

37.1

Cluster 2 – Averag 53% Parallel 45% Fortune

42.1

Cluster 2 – Cut‐ofGrade

65% Parallel 35% Fortune

33.7

13.1.2. Sample Preparation

The sample composites were homogenized and rotary split into two (2) kilogram charges. The laboratory prepared a Master Composite consisting of eight (8) kilograms of each of the six (6) composite samples i.e., a Cluster 1 Composite and a Cluster 2 Composite, each with an equal weight (4 kg) of the cut-off, average and high grade samples. All the composites were stored under nitrogen until needed in the test program.

13.1.3. Sample Characterization

13.1.3.1. Comminution Testwork

A Bond ball mill work index (BWI) test, with a closing screen size of 106 μm, was conducted on the Master Composite, the Cluster 1 Composite and the Cluster 2 Composite. The BWI can be used with the Bond’s Third Theory of comminution to calculate the net power requirement for grinding. The ball mill work index ranged from 13.8 to 14.9 kwh/tonne. This is considered as average in hardness, see Table 13.2.

Table 13.2 – Ball Mill Work Index Report Sample description kWh/t

KM3569 Master Composite 13.8

KM3876 Cluster 1 Composite 14.8

KM3876 Cluster 2 Composite 14.9

Both Cluster Composites generated very similar BWI indices. However, further tests will have to be performed with material of the hardest zones of the mineral deposit in order to clearly establish the overall grinding performance.

www.innovexplo.com


13.1.3.2. Mineralogy and Chemical Content

Duplicate Bulk Mineral Analyses via the QEMSCAN method were conducted on the six (6) composite samples, on the Master Composite and on the Cluster 1 Composite. The gold assay of the Master Composite sample gave about 8 g/t. The individual composite samples ranged from 3 g/t to 15 g/t Au, see Table 13.3.

Table 13.3 – Head Assays Summary

Sample Description Au Ag Fe S C As TOC*

g/t g/t % % % % %

Master Composite 8.15 4 5.1 1.47 ‐ 0.007 ‐

Cluster 1 – High 15.3 4 6.4 2.08 1.56 0.003 0.02

Cluster 1 – Average 6.28 1 6.4 1.72 1.76 0.002 0.02

Cluster 1 – Cut‐off 3.16 1 6.3 1.49 1.59 0.003 0.02

Cluster 2 – High 14.6 3 4.1 1.57 1.04 <0.002 0.05

Cluster 2 – Average 6.14 2 4.5 1.20 1.02 <0.002 0.02

Cluster 2 – Cut‐off 3.21 <1 4.0 0.99 0.79 <0.002 0.02

Cluster 1 Composite 8.66 ‐ 6.3 1.97 ‐ ‐ ‐ *TOC = Total Organic Carbon

The gangue minerals were primarily quartz and feldspars. The quartz content varied between 21.8% and 36% and the feldspars between 15% and 24.1%. The sulphur in the samples was present primarily as pyrite and traces of chalcopyrite. The content ranged from 1 to 2%. Iron ranged from 4 to 6%. The Cluster 1 samples contained more pyrite and amphibole which explains the greater presence of iron (6%). The assays showed a 1 to 2% carbon content but however, only a small part of the total carbon was present in the organic form. Organic carbon consumes cyanide and is then detrimental to cyanidation. A "Trace Mineral Search" was conducted using QEMSCAN on the high grade samples to determine the gold occurrence, including gold bearing particles analysis, liberation levels, size and association of the gold, see Table 13.4. The study was conducted at 84 μm K80 for Cluster 1 and 80 μm K80 for Cluster 2. It showed that 69 and 77% of the gold particles in Cluster 1 and Cluster 2 High grade samples, respectively, were liberated or in binary particles with sulphide minerals.

www.innovexplo.com


Table 13.4 – High Grade Trace Mineral Search

Sample Description Primary Grind

Size Liberated Gold

Au Binary – Sulphide

Minerals

μm % %

Cluster 1 – High 84 47 22

Cluster 2 – High 80 42 35

The portion of the gold particles present in inclusion was greater in Cluster 1 than in Cluster 2. Gold in inclusion is more difficult to leach as the exposed surface is low or nonexistent. In this case, finer grinding can minimize this impact.

13.1.4. Metallurgical Test Program

The purpose of the test program was to establish a preliminary flowsheet. The program consisted in:

Gravity separation and panning followed by cyanide leach of the gravity tailings;

Rougher flotation test.

13.1.4.1. Gravity and Cyanide Leach

Four gravity separation tests were performed on the Master Composite at different grind sizes (130, 105, 79 and 56 μm K80). The tests consisted of processing a 4 kg sample, pre-grinded into a laboratory Knelson unit. The concentrate thus obtained was then hand panned. Both gravity and pan tailings were subjected to leaching with a 1,000 ppm sodium cyanide concentration at pH 11 for a 48 hour period. The highest overall gold recovery obtained was 89% with the 79 μm grind size. About 23% of this recovery was via the gravity concentrate. The same procedure was applied to the six (6) zones composite samples with a K80 = 75 μm. These samples were submitted to three optimization tests as follows:

Increase of the retention time from 48 to 96 hours; Reduction of the grind size to the Knelson to 50 μm and increase of the

retention time to 96 hours; Increase of the cyanide concentration to 5,000 ppm only for the Cluster 1

samples.

A summary of the tests conditions and the recovery results is displayed in Table 13.5.

www.innovexplo.com


Table 13.5 – Gravity and Cyanide Leach – Conditions and Recoveries

Tests Conditions

K80 (μm) 75 NaCN(ppm) = 1,000 Retention (hrs) = 48 pH=11




Gold Recovery(%) Gravity Total Gravity Total Gravity Total Gravity Total

Master Composite 23 89

Cluster 1

Low Grade 12.3 81.0 18.5 86.8 18.4 88.7 20.3 91.6

Average 22.8 84.3 24.4 90.0 25.1 88.9 26.2 92.2

High Grade 14.8 79.3 17.3 86.4 16.4 87.5 20.9 91.1

Cluster 2

Low Grade 27.7 92.6 37.8 97.4 38.5 97.8

Average 31.6 94.3 29.9 96.9 44.3 98.2

High Grade 36.1 93.0 30.2 97.1 44.3 98.3

There was a notable difference in the performance of Clusters 1 and 2 samples. Gravity and leaching recoveries were better for Cluster 2. According to the tests conditions, the total gold recovery of Cluster 1 ranged from 79 to 92% whereas the one of Cluster 2 varied from 93 to 98%. In both cases, recovery increased with both the fineness of grind and retention time. Increase of the leaching time to ninety six (96) hours with the same cyanide concentration, increased the overall recovery. Increase of the sodium cyanide concentration to 5,000 ppm, with a forty height (48) hours leaching time, increased the overall recovery. Analysis of tailings showed that part of the gold present in the Cluster 1 samples was not in a leachable form (gold was locked in the sulphide matrix and gangue minerals).

13.1.4.2. Rougher Flotation

Rougher flotation tests were done on the Master Composite and the three (3) Cluster 1 samples, see Table 13.6. The objective was to evaluate the viability of incorporating flotation in the overall process. Two (2) kg samples were ground at 130 μm K80. Then, flotation was conducted in 4.4 liter lab cells with Potassium Amyl Xanthate (PAX) as the collector and Methyl Isobutyl Carbonyl (MIBC) as the frother. The Master Composite produced concentrate containing seventy one grams per tonne (71 g/t) of gold with a recovery of 89%. The concentrate produced from the Master Composite sample represented approximately eleven percent (11%) of the feed mass and 96% of the feed sulphur.

www.innovexplo.com


Table 13.6 – Summary of Flotation Gold Recoveries Obtained Sample % Recovery

Master Composite 89

Cluster 1 ‐ High 88

Cluster 1 ‐ Average 90

Cluster 1 – Cut‐off 82

13.1.4.3. Flotation, Regrind and Cyanidation

Testing consisted of a rougher sulfide flotation using Potassium Amyl Xanthate (PAX) as the collector and Methyl Isobutyl Carbonyl (MIBC) as the frother, at natural pH. Then, the flotation concentrate was finely ground to a K80 of 7µm. The concentrate was then intensively leached at 5,000 ppm of sodium cyanide for 48 hrs, and the tailings leached at 1,000 ppm of sodium cyanide for also 48 hrs. Cluster 1 Composite was submitted to rougher flotation at three pre-grinding sizes of 107,133 and 206 μm. The three composite samples (high, average, cut-off) were submitted to rougher flotation at 200 μm only, see Table 13.7.

Table 13.7 – Flotation, Regrind and Cyanide Leach – Recoveries Obtained

Description Gold Recovery (%)

107 μm 133 μm 206 μm 200 μm

Cluster 1 Composite

96.0 95.5 93.6

Cluster 1 – High 94.7

Cluster 1 – Average 95.1

Cluster 1 – Cut‐off 89.7 Decreasing the primary grind size from 206 to 107 μm, increased the overall recovery. More than 12 % of the gold content of the feed ended up in the tailings. Therefore, leaching these tailings increased the overall gold recovery. Without that, gold recovery would average 73, 87 and 84% respectively for the « cut-off grade », « average » and « high grade » samples.

13.2. Potential Processing Facility – Secondary Testwork

A third series of metallurgical testwork was undertaken in October 2013. The objective was to compare the potential metallurgical results of four (4) different flowsheets. These flowsheets were based on six (6) existing milling facilities that could potentially process the Lamaque mineralized material on a custom milling basis. These flowsheets were as follows:

www.innovexplo.com


Flowsheet 1: Gravity and carbon-in-leach (CIL); Flowsheet 2: Whole ore cyanidation; Flowsheet 3: Whole ore carbon-in-leach (CIL); Flowsheet 4: Flotation followed by cyanidation of concentrate and flotation

tailings.

13.2.1. Sample Description and Preparation

Two drums containing 183 samples were received at the ALS Metallurgy Kamloops laboratory, in October 2013. The material, crushed already at minus 6 mesh, was identified according to the mineralized zone it was sampled from: No. 4 Plug, Triangle, Parallel and Fortune, and grouped according to the gold grade (cut-off, average and high grade). The laboratory prepared a composite sample for each zone of the mineral deposit. For this part of the testwork, the grades were blended. The composites were homogenized, split into 2 kilogram lots and stored under nitrogen in a freezer, until used for testing. Chemical assays were performed on duplicate head cuts from each of the four composites and then the selected flowsheets were tested. All of them were tested with a primary grind size of 75 μm K80.

13.2.2. Properties of the Four (4) Zone Composites

Table 13.8 shows a summary of the assays. The gold grade varies from 4.5 g/t to 9.13 g/t, depending on the zone assayed.

Table 13.8 – Summary of Chemical Assays

Composite

Assays

Cu Zn Ag S S(s) Au*

% % g/t % % g/t

Fortune 0.024 0.03 4 1.11 1.08 6.25

Parallel 0.029 0.02 3 1.49 1.46 9.13

Triangle 0.009 0.01 5 1.58 1.54 8.79

No. 4 Plug 0.011 0.01 2 1.78 1.74 4.46 Note: S indicates total sulphur; S(s) indicates sulphur within sulphides * Au assays were completed using a screened metallic assay method.

13.2.3. Metallurgical Performance

13.2.3.1. Flowsheet 1: Gravity and Carbon-in-Leach (CIL)

10 kilograms of each composite were fed in a batch Knelson gravity concentrator with a hand panning procedure of the gravity concentrate. Tailings from this gravity concentration were submitted to a carbon- in-leach test with a concentration of 30 g/ litre of carbon, 1,000 ppm of cyanide, a pH of 11 and a retention time of 96 hrs, see Table 13.9.

www.innovexplo.com


Table 13.9 – Flowsheet 1 – Recoveries Obtained

13.2.3.2. Flowsheets 2 and 3: Whole Sample Cyanidation and Carbon-in-Leach (CIL)

For the second and third flowsheets, the same conditions were applied to each flowsheets with a 1,000 ppm cyanide concentration and a pH of 11, for a retention time of 96 hrs. However, for the third flowsheet, 30 g/l of carbon were added to the leach slurry. The following Tables 13.10-13.11 present the results obtained by flowsheet:

Table 13.10 – Flowsheet 2 – Recoveries Obtained Gold Recovery (%)

Composite Total

No. 4 Plug 83.2 Triangle 92.9

Parallel 97.1

Fortune 95.6


Composite Total

No. 4 Plug 85.1 Triangle 93.4

Parallel 96.6

Fortune 97.1

Gold Recovery (%)

Composite Gravity Total

No. 4 Plug 13.7 87.6 Triangle 17.6 93.0

Parallel 47.6 97.8

Fortune 26.8 96.6

www.innovexplo.com


13.2.3.3. Flowsheet 4: Flotation with Cyanidation of Concentrate

Flotation (rougher and cleaner) was carried out at a natural pH with Potassium Amyl Xanthate (PAX) as the collector and with Methyl Isobutyl Carbonyl (MIBC) as the frother. The concentrate recovered, which was about 3 to 4% of the feed mass, was leached at a concentration of 2,000 ppm of sodium cyanide with a pH of 11 for a period of 96 hours. The flotation tailings were also leached with a 1,000 ppm cyanide concentration, the same pH and retention time, see Table 13.12.


Composite Concentrates

Leach Tails Leach

Total

No. 4 Plug 58.4 24 82.4 Triangle 71.1 20.8 91.9

Parallel 85.8 8.9 94.7

Fortune 82.0 13.1 95.1 Overall gold recoveries from the first three flowsheets were comparable. With these flowsheets, gold recoveries from the Parallel, Triangle and Fortune composites ranged from 93 to 98% and recovery for the No. 4 Plug composite varied from 83 to 88%. The Trace Mineral Search performed during the first series of tests (report of April 3, 2013) shows that the Cluster 1 sample, the majority of which is composed of the No. 4 Plug Zone, presents gold particles in inclusion and therefore, more difficult to leach. But tests done on the Cluster 1 sample with a finer grind, demonstrated a better gold recovery. Gravity concentrates graded between 257 and 1,218 g/t Au. The sodium cyanide consumption doubles for a flowsheet using the CIL process. On average, it goes from 0.95 to 1.98 grams of reagent/tonne of feed. Flowsheet 4 using flotation, showed lower recoveries compared to the other three. Leaching of the concentrate recovered from 58 to 86% of the feed gold. Leaching of the flotation tailings recovered 9 to 24% of the feed gold.

13.3. Metallurgical tests summary

For the purpose of the preliminary economic assessment (PEA) of the project, WSP recommends to use recovery numbers that were obtained from retention times of 48 hrs because the plants that have been looked at for the various milling options, might offer retention times of about that order of magnitude or likely less. Based on the results obtained from tests KM4025-05 to 08 (ALS Metallurgy Kamloops, Roulston D., Johnston H., Metallurgical Testwork on the Lamaque

www.innovexplo.com


Deposit, January 16, 2014) and after adjusting leaching recoveries for a 48 hrs retention time, using the leach kinetic curves, the following recoveries could be used if the only process used is cyanidation:




No. 4 Plug: 82.0% (represents 98.5% of recovery obtained with a retention time of 96 hrs)

The percentage of the 96 hrs recovery for the Triangle Zone was lower than the one obtained with the others after 48 hrs (96.5% vs. >98%). To validate that, the ALS Kamloops laboratory redid a cyanidation test with the same Triangle Zone composite and similar test conditions (KM4232). The results obtained, indicate a gold recovery of 89.5 % after a leaching time of 48 hrs, which represents 98.5% of the gold recovered after 96 hrs. Always based on the same report, if the process involves gravity separation with cyanidation of the gravity tails, the results obtained in KM4015-01 to 04 (gravity testwork) and KM4025-25 to 28 (cyanidation of gravity tails), were used after calculating leaching recovery for a 48 hrs retention time, using the factors above, derived from the leach kinetic curves of tests 05 to 08, to get the following recoveries that could be used for the PEA:


*The average of both tests (Reports KM4025 and 4232) was used to calculate the recovery factor for the Triangle Zone (89.6% recovery after 48 hrs which is 97.5% of the 91.9% recovery after 96 hrs). However the results obtained to-date, from the series of laboratory tests performed by ALS Metallurgy Kamloops, are preliminary and further testwork will be needed to better study gravity concentration, determine the fineness of grind and reagents consumption that will enhance recoveries, and define the best flowsheet that could be used to optimize metallurgical performance as well as economics. A bulk sample run in the plant selected could be extremely useful to adjust the process developed in the lab to a plant scale process.

www.innovexplo.com


14. MINERAL RESOURCE ESTIMATES

The 2013 Mineral Resource Estimates for the Fortune, Parallel, No. 4 Plug and Triangle zones of the Lamaque Project, presented herein, were completed by the author, Christian D’Amours, P.Geo. (OGQ # 226) of GéoPointCom, using all available results as per the effective date of each zone. GéoPointCom was contracted by Hervé Thiboutot, P.Eng., Senior Vice-President of Integra Gold. The main objective was to publish revised mineral resource estimates for the above mentioned zones. The mineral resources presented herein are not mineral reserves as they have no demonstrable economic viability. The result is a Mineral Resource Estimate with Indicated and Inferred resources for each of the four mineralized zones, modelled for underground mining. The Qualified and Independent Person responsible for the Mineral Resource Estimate, as defined by NI 43-101, is Christian D’Amours, P.Geo (OGQ #226). The estimate is based on an interpretation performed by Alain-Jean Beauregard, P.Geo. (OGQ #227) and Daniel Gaudreault, P.Eng. (OIQ #39834) of Geologica. The resource estimates for the abovementioned zones are part of a NI 43-101 Technical Report prepared and supervised by Geologica Inc. and filed on SEDAR in November 2013. The effective dates of the Mineral Resource Estimates vary from zone to zone but the common publication date by news release is September 25, 2013. All details for these resource calculations are presented in Tables 14.1, 14.2 and 14.3. The November 2013 mineral resources report for the four zones were included in the current PEA. The 2014 Mineral Resource Estimates for the No. 6 Vein and Sixteen Zone, presented herein, were completed by Christian D’Amours, P.Geo. (OGQ #226) of GéoPointCom, using all available results as per the effective date of each zone. The publication date by press release is January 28, 2014. GéoPointCom was contracted by Hervé Thiboutot, P.Eng, Senior Vice-President of Integra Gold, to perform the mineral resource estimation. The result is Mineral Resource Estimates with Indicated and Inferred Resources (see Tables 14.1, 14.2 and 14.3). The mineral resources are not mineral reserves as they have no demonstrable economic viability. These mineral resources were not included in the current PEA.

Table 14.1 – Total Indicated Resource Estimate by zone using a 3.00 g/t Au cut-off


No. 4 Plug 1,325,100 5.6 237,450

Fortune Zone 125,500 5.8 23,600

Parallel Zone 793,900 8.2 209,570

Triangle Zone 599,700 9.9 190,670

No. 6 Vein 389,400 6.4 79,550

Sixteen Zone 91,700 5.2 15,440

Total Indicated 3,325,300 7.1 756,280

www.innovexplo.com


Table 14.2 – Total Inferred Resources Estimate by zone using a 3.00 g/t Au cut-off


No. 4 Plug 0 0.0 0

Fortune Zone 252,300 5.6 45,220

Parallel Zone 153,400 17.5 86,050

Triangle Zone 332,300 12.9 137,600

No. 6 Vein 111,600 6.9 24,590


Total Inferred 851,400 10.8 293,710

Table 14.3 – Key parameters for the 2013 and 2014 Mineral Resource Estimates by zone

14.1. Resource Estimate – Fortune Zone

The estimation was performed in November and December 2012. The validated database was completed November 8, 2012 and this was considered as the effective date for the resource estimation of the Fortune Zone (formerly the Forestel Zone). Diamond drill holes FOR-11-06 and T-06-04-09 were the last drill holes to be validated and verified for inclusion in the resource calculation.


Min. search radius

Max. search radius

Min. no. samples

Max. no. samples

Capping grade

Cut Off grade

Min. true thickness

Compo-site

length

Effective date


10X10X15m 25m 90m 3 10 None 3 2m 0.5m 2012 11/08


5X5X5m 50m 50m 3 15 100


2012 05/24

No. 4 Plug

Ordinary Kriging

10X10X10m 35X50X

16m 60X60X

16m 3 10 300 3 ** 1.0m

2013 03/19


5X5X5m 25m 50m 2 10 80


2013 04/24

No. 6 Vein

Ordinary Kriging

10X10X10m 50X50X

50 100X100

X100 4 8

40 [progressive]

3 2m 1.0m 2012 08/17


10X10X10m 15X15X

15 60X60X

60 5 10 35 3 ** 0.7m

2013 11/18



www.innovexplo.com


14.1.1. Methodology

The Mineral Resource Estimate and geostatistical study detailed in this report was performed using the Isatis (V.2012.4) software package. The method involves a 3D block model estimated with an ordinary kriging (“OK”) interpolator.

14.1.2. Drill hole Sample Database

The GeoticLog/MS Access diamond drill hole database was reviewed and validated by Geologica. It contains 43 surface diamond drill holes with conventional analytical gold assay results, as well as coded lithologies from drill core logs. The 43 drill holes account for 12,207.68 m of core and yielded 4,532 cut and assayed samples for a total of 4,058.32 m. The database also contains 422 QA/QC samples.

14.1.3. Interpretation of Mineralized Zones

The interpretation of the host geological units, their mineralized envelope, veins, structures combined and conjugate zones was completed by Benjamin Blaise, P.Geo., M.Sc., of Geologica under the supervision of Alain-Jean Beauregard, P.Geo. and Daniel Gaudreault, Eng. (OIQ # 39834). A wireframe solid representing eight (8) individual mineralized auriferous structures oriented N270/-45 was constructed by the author. The wireframe solid was created by digitizing an interpretation onto nine (9) sections spaced 50 m apart, and then using tie-lines to complete the wireframes for each individual vein (Figure 14.1). For each vein intersection, the minimum true thickness was set to 2 m. When true thickness was inferior to the minimum thickness, internal dilution was included to achieve the minimum thickness. Thus three selection sets were created within the database:

One selection representing all samples within the “vein” regardless of minimum thickness (“Vein Selection”);

Another selection representing samples added to achieve the minimum thickness (“Internal Dilution Selection”); and

The third selection is the midpoint of each Vein Selection where the true thickness of the Vein Selection is estimated (“Intersection Selection”).

www.innovexplo.com


Figure 14.1 – Wireframe solids of the Fortune Zone (formerly the Forestel Zone).

14.1.4. High Grade Capping

It is common in the industry to remove some of the highest (aberrant) values from assay distribution prior to compositing the samples. The main objective of this process is to make sure that a peak erroneous value could not affect grade estimation. With the development of statistical methods for estimating grade, this process became less important. In fact, Kriging and especially simulation techniques are less sensitive to some occasional very high values. On the other hand, the presence of some very high values may make the variogram difficult to establish. Geostatistics provide some efficient tools to solve this problem. The more common methods used to assess the necessity of using a capping value are listed below:

The first indication of the necessity to cap high values is the coefficient of variation “CV”. Ideally, this should be located close to 1. A CV above 2 is generally considered an indication that high values should be capped. In the case of very high CV, uncapped grades may make it difficult to produce a clean variogram.

In the case of a simple normal or log-normal population, the probability curve should form a relatively straight line. A positive break in the upper end is often interpreted as an indication that high values should be capped. This criterion is probably the strongest indication, especially when the interpolation method is based on a normal or log–normal distribution.

The metal factor method consists of comparing the cumulated metal percentage with the cumulated data percentage. This technique takes for granted that all samples represent an equivalent number of tonnes. Specialists generally agree to keep this factor below 10. In other words, there

www.innovexplo.com


should never be more than 10% of ounces in less than 1% of tonnes. This method really concentrates on the upper 1 or 2% of the distribution. It is more a security factor than a distribution analysis.

Figure 14.2 shows a comparative histogram of the log-transformed gold values from each vein and Figure 14.3 displays the probability plot from all samples within the Vein Selection population. From these graphs it is clear that the Vein Selection samples do not form a simple normal or log-normal population. It seems to be a bi-modal distribution where the upper part (50% of the population) may follow a log-normal distribution without any positive break at higher grades. The distributions of the metal content within the percentile classes do not show any abnormal concentration. So only the CV of 2.46 would support a capping grade. Thus, the author does not recommend using a capping grade.

Figure 14.2 – Comparative histogram.

www.innovexplo.com


Figure 14.3 – Probability plot

14.1.5. Compositing

Compositing was compiled using the grade of the adjacent material when assayed, or a value of zero when not assayed. In order to minimize any bias introduced by the variable sample lengths, uncapped gold assays (as determined in the previous subsection) were composited to 0.5 metre equal lengths (“0.5m composites”) within all intervals that define each of the mineralized zones (Vein Selection) and internal dilution (Internal Dilution Selection). Tails were not created; they were instead distributed over the previous composites.

www.innovexplo.com


14.1.6. Variography

The author modelled the composite variography of the three selected items (Vein, Dilution and Intersection) using Isatis software. The objectives of this step consist of:

Characterizing the anisotropy and setting the dimensions and direction of a search ellipsoid to use during the interpolation of values of the block model. The dimensions of the three axes of the ellipsoid are equal to the range measured on the variograms. This way, it is statistically shown that wherever the center of this ellipsoid is located, all samples included within it will have a variance inferior to that of the entire population. It is reasonable to use this sub selection to estimate the value of a central point. The anisotropy may also be used to select and weigh samples during the “de-clustering” steps.

Defining the Kriging equations. The equations deriving from the variograms are required during the Kriging interpolation. These equations take into account the nugget effect (C0), the model of dispersion (spherical, exponential…), the range and the variance in each of the three axes (σ1, σ2 and σ3). It is frequent to use more than one model in order to better represent the dispersion (short and long) of each of the axes before reaching the level (total variance).

The author first calculated an omnidirectional variogram to estimate the nugget effect (C0). Second, he calculated 18 directional variograms located in the horizontal plane. This allows for defining the direction of longer continuity (longitudinal direction). The next step includes 18 directional variograms set in the vertical plane parallel to the longitudinal direction and another 18 directional variograms set in a vertical plane perpendicular to the longitudinal direction. And finally 18 directional variograms on an inclined plane representing the best plan fit for all mid points (longitudinal plan). This step allows having an overview of the variance of the idealized sphere in order to verify the presence of a directional anisotropy. The variograms so obtained are considered robust and reliable. This means that they resist well enough to parameter changes such as tolerance or lag and they compare well to the results obtained using other types of variograms. The last step consists of fitting a mathematical model allowing representation of these experimental variograms. The author was unable to clearly demonstrate the presence of any directional anisometry laid on the best plan fit. Of course, the axes perpendicular to the vein are very short compared to all axes within the vein plan but this is not considered as an anisometry as we used hard boundaries from individual veins when interpolating. Figure 14.4 shows the omnidirectional model variography proposed for all three (3) composite selections considered.

www.innovexplo.com


. Figure 14.4 – Omnidirectional variography.

www.innovexplo.com


14.1.7. Bulk Density

A specific gravity of 2.80 g/cm3 was used to estimate the tonnage for the 2012 Mineral Resource Estimate.

14.1.8. Block Model Geometry

A block model was established to cover the entire drilled area. The origins of the block model are as follows (center of the front, bottom, left cell): Easting: 295980mE (29 cells 10 m each) Northing: 5330050mN (27 cells x 10 m each) Elevation: 90m (15 cells x 15 m each) The block model was not rotated.

14.1.9. Mineralized Zone Block Model

All blocks with at least 0.001% of their volume falling within a selected solid were assigned the corresponding solid vein code. A percent block model was generated reflecting the proportion of vein in each block.

14.1.10. Grade Block Model

For each cell, three different values (True Thickness, Ore Grade and Dilution Grade) were estimated. In all cases, vein identification and samples identification must strictly match in order to be used (hard boundary). All estimations were carried out using an OK estimator. For each value estimated, up to three consecutive passes were carried out using three different search ellipsoid sizes. For the first pass, the radius of the search ellipsoid was set to 25 m. If the minimum number of samples was not found, a try was made for the second pass. With the second pass the radius was extended to 45 m. If the minimum number of samples was still unavailable, then the radius was extended to 90 m for the final attempt. When interpolating true thickness, the minimum number of samples within the search ellipsoid was set to 2 and the maximum was 10. For Ore Grade and Dilution Grade, the minimum number of samples was set to 3 and the maximum to 10. No other restriction parameters (octants, parent hole, distance) were used. For each cell where the estimated true thickness was less than 2, the final grade was calculated using the proportional values of the Ore Grade and Dilution Grade with respect of the estimated true length. Of course, when the estimated true thickness was 2 or more, the final grade was set to the value of the Ore Grade without any internal dilution.

14.1.11. Resource Categories

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”.

www.innovexplo.com


Measured Mineral Resource: the 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: the 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: the 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. Resources from this category should not be used to support mine planning and evaluation of the economic viability of the deposit. The Indicated category was assigned to each cell where true thickness was estimated using the first pass ellipsoid, and at least two (2) drill holes were found within a 25 m radius. An Inferred category was assigned to all cells where the true thickness was estimated using the second or the third pass ellipsoid. No measured category was assigned to any blocks.

14.1.12. Minimum cut-off Value

By definition, the “Cut-Off" is the breakeven point considering total cost and revenue generated by the operation. The current estimation used the following economic parameters: Gold value = $1,450/oz Mining cost = $90/t (metric ton) Milling fees = $25/t (metric ton) Gold recovery = 92% These parameters give a first estimation of the cut-off at 2.27 g/t Au. It is clear that the parameters do not represent total cost. Environmental, developmental, administration costs, etc. were not taken into account. Another 25% may be added to the cost previously considered, which would push the estimation of the cut-off to 3.35 g/t. The author suggests using 3.0 g/t for the official estimation.

www.innovexplo.com


Other cut-off grade results were also compiled, but for illustrative purposes only. The cut-off grade must be re-evaluated in light of prevailing market conditions and other factors: gold price, exchange rate, mining method, related costs, etc. (Figure 14.5).

Figure 14.5 – Sensitivity of estimates to different cut-off values 14.1.13. Mineral Resource Estimate Results

Given all parameters listed above, GéoPointCom is of the opinion that the current Mineral Resource Estimate can be classified as Indicated and Inferred resources (Table 14.4). The estimate is compliant with CIM standards and guidelines for reporting mineral resources and reserves.

www.innovexplo.com


Table 14.4 – Fortune Zone Mineral Resources

14.1.14. Comparison to Previous Mineral Resource Estimates

Comparing mineral resource estimates may appear to be a simple exercise, but in reality, obtaining real conclusions from such comparisons is difficult due to differences in key assumptions, parameters and methods and the interaction between these features in the final results. In the current case, historical estimates were not NI 43-101 compliant and will not be commented on. In December 2010, GéoPointCom was first involved in a resource estimation for the Forestel Zone (now called Fortune). At that time, the zone was interpreted with five (5) subparallel veins oriented more or less N275/-28. The block model was estimated using an OK estimator and a search ellipsoid of 50 m radius laid on an idealized plan (perpendicular axis was 10 m). Because of the lower confidence level on the geological model, all the resources were classified as Inferred and no cut-off was

www.innovexplo.com


used. At that time, Inferred resources where estimated at 861,000 tonnes at 2.1 g/t Au (58,000 ounces). If we consider no cut-off and everything as Inferred, the current estimate will represent 1,270,000 tonnes at 2.4 g/t Au (98,700 ounces). Thus, new drilling and refinement of the geological model allowed for an impressive upgrade.

14.2. Resource Estimate – Parallel Zone

The estimation was performed in June 2013. The cut-off date for the database was May 24, 2012. The last drill hole included and considered is PV-11-31. The effective date for the resource estimate of the Parallel Zone is May 24, 2012. The current estimate represents an update of the last estimate prepared by Christian D’Amours of GéoPointCom (the author of this section) in March 2011. It is important to note that since the last estimation, a major correction to the database was to add a 19° counter-clockwise rotation (Az. 341° instead of 360°) to all drill holes prior to 2010. This had a major impact on the 3D modelling.

14.2.1. Methodology

The Mineral Resource Estimate and geostatistical study detailed in this report was performed using Isatis (V.2013.01) software. The method involves a 3D block model estimated with an OK interpolator.

14.2.2. Drill Hole Sample Database

The Geotic / MS Access diamond drill hole database is maintained and updated by Geologica employees (P.Geo. and P.Tech.) under the supervision of Alain-Jean Beauregard, P.Geo. and Daniel Gaudreault, Eng. It contains 158 surface diamond drill holes (compared to 124 for the March 2011 estimation) with conventional analytical gold assay results, as well as coded lithologies from the drill core logs. The 158 drill holes yielded 39,285 m of core and 13,305 samples (compared to 7,458 in March 2011) for a total of 11,555 m of saw-cut and assayed core. The database also contains 1,179 QA/QC samples.


The interpretation was completed by Benjamin Blaise and Laura Guillaume, junior geologists of Geologica, under the supervision of Alain-Jean Beauregard and Daniel Gaudreault of Geologica. In order to conduct accurate resource modelling of the Parallel Zone, the author constructed a wireframe solid representing 17 individual veins oriented more or less N077/-45. The difference between the actual interpretation and that made in March 2011 is due to the addition of 34 new drill holes and also to a major azimuth change for more than 50% to the drill hole because of a survey correction applied to the database. Figure 14.6 compares the previous interpretation (in black) with the current interpretation. The wireframe solid was created by digitizing an interpretation onto 25 sections spaced 25 m apart, and then using tie-lines to complete the wireframes for each individual vein (Figure 14.7). For each vein intersection, the minimum true thickness was set to 2 m.

www.innovexplo.com


Figure 14.6 – Changes to the interpretation from March 2011 to May 2012.

www.innovexplo.com


Figure 14.7 – Wireframe solids of the Parallel Zone.


It is common in the industry to remove some of the highest (aberrant) values from the assay distribution prior to compositing samples. The main objective of this process is to make sure that the erroneous value could not affect grade estimation. With the development of statistical methods for estimating grade, this process became less important. In fact Kriging, and especially simulation techniques, are less sensitive to some occasional very high values. On the other hand, the presence of some very high values may make the variogram very difficult to establish. Geostatistics provide some efficient tools to solve this problem. The more common methods used to assess the necessity of using a capping value are listed below:

The first indication of the necessity to cap high values is the coefficient of variation “CV”. Ideally, this one should be located close to 1. A CV value above 2 is generally considered as an indication that high values should be capped. In fact this is highly related to the difficulty of producing a clean variogram when the CV is very high.

www.innovexplo.com


In the case of a simple normal or log-normal population, the probability curve

should form a relatively straight line. A positive break in the upper end is often interpreted as an indication that high values should be capped. This criterion is probably the strongest indication, especially when the interpolation method is based on a normal or log–normal distribution.

The metal factor method consists of comparing the cumulated metal

percentage with the cumulated data percentage. This technique takes for granted that all samples represent an equivalent number of tonnes. Specialists generally agree to keep this factor below 10. In other words, there should never be more than 10% of ounces in less than 1% of tonnes. This method really concentrates on the upper 1 or 2% of the distribution. It is more a security factor than a distribution analysis.

Figure 14.8 shows the distribution (density, histogram, box plot and probability plot) of gold values for all samples within the Vein Selection (ore in red). From this graph it is clear that Vein Selection samples do not form a simple normal or log-normal population. It seems to be a bi-modal distribution where the number of samples having a grade equal to 0.5 g/t is abnormally high. The higher end of the probability plot shows a small bump around 200 g/t Au but returns to the normal (straight line) immediately after this sample. The metal content distributions within the percentile classes show that 25% of the metal content is contained within the two last percentiles. To satisfy the metal factor method explained above, the higher grade could be capped at 36 g/t but this approach would seem excessive. Finally, although the CV of 3.37 would argue in favor of a capping grade, even if we cap at 36 g/t, the CV will still be high at 2.09. At this point we may presume that the high CV will make variography relatively difficult to establish. The only way to really quantify the effect of high values on the estimate is to repeat the estimation at several capping grades and compare the final number of ounces in each case. The author used 16 different capping grades from 272 g/t (uncapped) to 10 g/t before compositing. The resource was then estimated using an OK estimator on a block model of 5X5X5. Figure 14.9 shows the percentage of ounces lost at different capping grades presented on a log scale. This graph shows that a cap at 80 g/t represents a little bit less than 1% (0.9%) of the data and will produce a loss of close to 10% (11%) of the estimated ounces. This graph shows an important slope change after 100 g/t. Slope change on this type of graph can be used to point out a subpopulation break. In the present case it means that a small grade interval over 100 g/t has a higher influence on total ounces than the same interval below 100 g/t. Thus the author is proposing to use a gradual capping method. It is suggested to reduce by 60% the grade interval over 100 g/t for all samples exceeding 100 g/t, according to the following formula: Au Capped = 100 + (0.4* (Au – 100))

www.innovexplo.com


Figure 14.8 – Comparative histogram

Figure 14.9 – Capping effect on total ounces.

www.innovexplo.com


14.2.5. Compositing

Compositing was compiled using the grade of the adjacent material when assayed, or a value of zero when not assayed. In order to minimize any bias introduced by the variable sample lengths, uncapped gold assays (as determined in the previous subsection) were composited to 0.6 m equal lengths (“0.6m composites”) within all intervals that define each of the mineralized zones (Vein Selection). Tails were not created; they were instead distributed over the previous composites.

14.2.6. Variography

The author modelled the composite variography of Vein 6B only using Isatis software. Mixing all the veins created a stationary problem and Vein 6B was the only one having sufficient samples for this type of study. The objectives of this step consist of:

Characterizing the anisotropy and setting the dimensions and directions of a search ellipsoid to use during the interpolation of values of the block model. The dimensions of the three axes of the ellipsoid are equal to the range measured on the variograms. This way, it is statistically shown that wherever the center of this ellipsoid is located, all samples included within it will have a variance inferior to that of the entire population. It is reasonable to use this sub selection to estimate the value of a central point. The anisotropy may also be used to select and weigh samples during the “de-clustering” steps.

Defining the Kriging equations. The equations deriving from the variograms

are required during the Kriging interpolation. These equations take into account the nugget effect (C0), the model of dispersion (spherical, exponential, etc.), the range, and the variance in each of the three axes (σ1, σ2 and σ3). It is frequent to use more than one model in order to better represent the dispersion (short and long) of each of the axes before reaching the level (total variance).

The author first calculated an omnidirectional variogram to estimate the nugget effect (C0). Second, he calculated 18 directional variograms located in the horizontal plane. This allows defining the direction of longer continuity (longitudinal direction). The next step includes 18 directional variograms set in the vertical plane parallel to the longitudinal direction and another 18 directional variograms set in a vertical plane perpendicular to the longitudinal direction. And finally 18 directional variograms on an inclined plane representing the best plane fit for all mid points (longitudinal plan). This step allows having an overview of the variance of the idealized sphere in order to verify the presence of a directional anisotropy. The model variogram thus obtained is considered robust and reliable. Surprisingly, the CV does not seem to cause any problems at this step. It has therefore been proposed to use a search ellipsoid where the two longer axes (60 m) are laid on the idealized plane of the vein. The third one is perpendicular and much shorter (3 m).

www.innovexplo.com


Figure 14.10 shows the experimental variogram obtained from the more abundant vein selection.

Figure 14.10 – Directional variography




A block model was established to cover the entire drilled area. The origins of the block model are as follows (center of the front, bottom, left cell): Easting: 294702.5mE (515 cells x 5 m each) Northing: 5329992.5mN (56 cells x 5 m each) Elevation: -292.5m (124 cells x 5 m each) The block model was not rotated and the error on volume calculated from the block model compared to the volume estimated from the wireframe is lower than 0.02%.

www.innovexplo.com





For each cell the grade was estimated for each vein individually using only relevant composites. All ore composites were used to estimate all the cells previously identified as included within a vein zone. All the estimations were done using the OK model as defined in Table 14.5. The search ellipsoid used for selecting samples has the same geometry as the Kriging model except for the length of the axes. The two longer ones (parallel to the vein) were shortened to 50 m and the short one (perpendicular to the vein) was extended to 15 m to accommodate vein undulations. When the ore grade is interpolated, the search ellipsoid as defined in section 14.2.10 must contain a minimum of 3 composites (the 3 closest), and in this case the 15 closest points were retained. If the minimum number of samples could not be reached, the cells remained unestimated. No other restriction parameters (octant, parent hole, distance) were used.

Table 14.5 – Model variogram

Structure sill Range Azimut/Dip

Nugget effect 33.8 ‐‐‐‐ ‐‐‐‐

Spherical(U) 54.82 60m N223/‐29

Spherical(V) 54.82 60m N118/‐24

Spherical(W) 54.82 3m N355/‐50

Axes Range Azimut/Dip

U 50m N223/‐29

V 50m N118/‐24

W 15m N355/‐50

Model Variogram

Search ellipsoid


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: the 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

www.innovexplo.com


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: the 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: the 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. Resources from this category should not be used to support mine planning and evaluation of the economic viability of the deposit. The Indicated category was assigned to each cell where the number of composites within a radius of 25 m is equal to or more than 3. All other cells were assigned the Inferred category. Because of the lack of information and thus the uncertainty of the 3D model, all cells from Vein 7 are classified as Inferred. No measured category was assigned to any blocks.


By definition, the cut-off is the breakeven point considering total cost and revenue generated by the operation. The current estimation used the following economic parameters.

Gold value = $1,450/oz Mining cost = $90/t (metric ton) Milling fees = $25/t (metric ton) Gold recovery = 92%

These parameters give a first estimation of the cut-off at 2.27 g/t. It is clear that the parameters do not represent total cost. Environmental, developmental and administration costs, etc. were not taken into account. Another 25% may be added to the costs previously considered. Doing so would push the estimation of the cut-off to 3.35 g/t. The author suggests using a cut-off of 3.0 g/t for official estimation. Other cut-off grade results were also compiled, but for illustrative purposes only (Table 14.6). The cut-off grade must be re-evaluated in light of prevailing market conditions and other factors including gold price, exchange rate, mining method, related costs, etc.

www.innovexplo.com


Table 14.6 – Sensitivity of Resource to different cut-off values

14.2.13. Mineral Resource Estimate Results

Given all parameters listed above, GéoPointCom is of the opinion that the current 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. If no cut-off is used, the current estimate will represent 1,550,800 tonnes at 4.8 g/t (239,830 ounces of gold) in the Indicated category and 275,900 tonnes at 10.3 g/t Au (91,090 ounces of gold) in the Inferred category. At a cut-off of 3.0 g/t Au, the official result is listed in Tables 14.7-14.8.

www.innovexplo.com


Table 14.7 – Mineral Resources in the Parallel Zone per vein

www.innovexplo.com


Table 14.8 – Mineral Resources in the Parallel Zone (all veins) per elevation

www.innovexplo.com



Comparing mineral resource estimates may appear to be a simple exercise, but in reality, obtaining real conclusions from such comparisons is difficult due to differences in key assumptions, parameters and methods and the interaction between these features in the final results. In the current case, historical estimates are not 43-101 compliant and will not be commented on. In March 2011, GéoPointCom was first involved in resource estimation for the Parallel Zone. Since that time, a major correction applied to the azimuth of all holes prior to 2010 has had a big influence on the 3D modelling. The previous resource estimation was reported at a 3.0 g/t Au cut-off. At that time, the Indicated resource, regardless of elevation, was estimated at 660,000 tonnes at 4.8 g/t Au (101,790 oz Au). Inferred resources were estimated at 211,500 tonnes at 1.7 g/t Au (11,680 oz Au). New drilling and azimuth corrections applied to all historical holes allow for more consistent and continuous vein modelling. Several previously considered intersections had been checked as isolated discontinuity values in the previous mineral resource estimation of the Parallel Zone. With a corrected azimuth of 19˚ on the diamond drill hole orientations, continuity of mineralization has now been demonstrated in the current resource estimation update. The March 2011 drill hole data only had 53 intersections above 3.0 g/t Au over a minimum width of 3 m, whereas the 2013 resource modelling used 102 intersections above 3.0 g/t Au over a minimum width of 2 m. For this reason, the recent resource update consists of a 20% increase in tonnage and a 70% increase in grade for an increase of 105% in the number of ounces.

14.3. Resource Estimate No. 4 Plug

The resource estimation of the No. 4 Plug was completed between January and March 2013. The data validation included diamond drill hole P4-12-21A, which was the last DDH to be considered in the calculation which was completed March 19, 2013.

14.3.1. Methodology

The Mineral Resource Estimate and Geostatistical Study were completed using the ISATIS (V. 2013.1) software package. The method takes into account the 3D Block Model which is estimated with an OK interpolator. Sample selection was not strictly constrained to lithology.

14.3.2. Drill Hole Database

The Geotic/MS Access diamond drill hole database was reviewed and validated by Geologica. It consists of 23,916 core samples taken from 48,152.6 m distributed over 113 surface diamond drill holes. A total of 19,865.8 m were saw cut and assayed for gold. The database also includes 1,426 QA/QC samples and associated coded lithological symbols from the core logging.

www.innovexplo.com



The interpretation of the host geological units, their mineralized envelope, veins, structures and combined zones was completed by Benjamin Blaise, P.Geo., of Geologica, under the supervision of Alain-Jean Beauregard, P.Geo., and Daniel Gaudreault, Eng.. Several approaches were used by the author in order to select the most adequate resource modelling of the mineralized gold zones in the No. 4 Plug. The approaches used either: (1) the main veins only; or (2) a conditional simulation based on the gold values being strictly constrained to the vein or to the vein cluster or to values hosted outside the veins within the host rock, considering assay results of ≥ 1g/t Au. The option that seems to be the most plausible is based on the construction of a solid wireframe representing an area containing a higher grade vein density or clusters (High Probability Ore) within the No. 4 Plug diorite intrusion with respect to lower vein density areas. The wireframe solid was created by digitizing an interpretation onto 14 sections spaced 25 m apart, and then using tie-lines to complete the High Probability Ore wireframes for this unique volume (Figure 14.11). Because the first 100 m was already mined during a previous operation, the upper limit was set to 150 m below surface. This wireframe was used for selecting all drill hole intersections. Areas selected but not sampled were considered as having a minimum grade of 0.001 g/t Au.

www.innovexplo.com


Figure 14.11 – Wireframe solids from No. 4 Plug .

www.innovexplo.com



It is common in the industry to remove some of the highest (aberrant) values from the assay distribution prior to compositing the samples. The main objective of this process is to make sure that the peak value does not affect grade estimation. With the development of statistical methods for estimating grade, this process became less important. In fact Kriging and especially simulation techniques are less sensitive to some occasional very high values. On the other hand, the presence of some very high values may make the variogram difficult to establish. Geostatistics provide some efficient tools to solve this problem. The more common methods used to assess the necessity of using a capping value are listed below:

The first indication of the necessity to cap high values is the coefficient of variation “CV”. Ideally, this should be located close to 1. A CV above 2 is generally considered an indication that high values should be capped. In the case of very high CV, uncapped grades may make it difficult to produce a clean variogram.


should form a relatively straight line. A positive break in the upper end is often interpreted as an indication that high values should be capped. This criterion is probably the strongest indication, especially when the interpolation method is based on a normal or log–normal distribution.

The metal factor method consists of comparing the cumulated metal


Figure 14.12 shows the distribution (density, histogram, box plot and probability plot) of gold values for all samples within the High Probability Ore. From this graph it is clear that High Probability Ore samples do not form a simple normal or log-normal population. It seems to be a bi-modal distribution in which the upper part (25% of the population) may follow a log-normal distribution. On this graph, a positive break at the higher grades (over 300 g/t) is visible. One can propose an assemblage of two log-normal populations to explain this distribution. The waste population may represent close to 75% of the samples with a median close to 0.04 and grades ranging from 0.002 to 1 g/t. In that case the ore population may have a median around 2.5 g/t and grade ranging from 0.1 to 300 g/t. The distributions of the metal content within the percentile classes show more than 35% of the total metal appearing within less than 1% of the samples. To satisfy the metal factor methodology (no more than 10% of the gold ounces within 1% of the sample), the grade must be capped at 13 g/t. This severe capping will affect 2.8% of the samples and will drive the CV from 8.1 to 2.3. This is considered overly conservative and unrealistic for the observed bi-modal distribution. The author recommends limiting the capping of higher gold values at 300 g/t. This capping limit will affect four (4) samples, drive the CV to 4.9, and reduce the total amount of gold by 5.9%. At this

www.innovexplo.com


point we may presume that the high CV will make variography very difficult to establish.

Figure 14.12 – Grade distribution within the High Probability Ore.

14.3.5. Compositing

From Figure 14.13, it is clear that an important bias exists between sample lengths and gold values (higher grade samples are shorter). The current sampling procedure tends to minimize this bias but it is definitely present in historical data. Compositing was compiled using the capped grade (300 g/t) of material when assayed, or a value of 0.001 when not assayed. In order to minimize the bias introduced by the variable sample lengths, capped gold assays (as determined in the previous subsection) were composited to 1.0 m equal lengths (“1.0m composites”) within all intervals that define High Probability Ore. Tails were not created; they were instead distributed over the previous composites.

www.innovexplo.com


Figure 14.13 – Grade versus length.

14.3.6. Variography

The author of this section 14 modelled the composite variography of the High Probability Ore using Isatis software. The objectives of this step consist of:

Characterizing the anisotropy and setting the dimensions and direction of a search ellipsoid to use during the interpolation of values of the block model. The dimensions of the three axes of the ellipsoid are equal to the range measured on the variograms. This way, it is statistically shown that wherever the center of this ellipsoid is located, all samples included within it will have a variance inferior to that of the entire population. It is reasonable to use this subselection to estimate the value of a central point. The anisotropy may also be used to select and weigh samples during the “de-clustering” steps.

www.innovexplo.com


Defining the Kriging equations. The equations deriving from the variograms are required during the Kriging interpolation. These equations take into account the nugget effect (C0), the model of dispersion (i.e. spherical, exponential), the range and the variance in each three axes (σ1, σ2 and σ3). It is common to use more than one model in order to better represent the dispersion (short and long) of each of the axes before reaching the level (total variance).

An omnidirectional variogram was first calculated to estimate the nugget effect (C0). Secondly, 18 directional variograms located in the horizontal plane were calculated and permits define the direction of longer continuity (longitudinal direction). The next step includes 18 directional variograms set in the vertical plane parallel to the longitudinal direction and another 18 directional variograms set in a vertical plane perpendicular to the longitudinal direction. And finally, 18 directional variograms on an inclined plane representing the best plane fit for all mid points (longitudinal plane). This step allows having an overview of the variance of the idealized sphere in order to verify the presence of a directional anisotropy. Variograms on raw data were impossible to establish. Thus variography was constructed and modelled using the Gauss transformation instead of raw distribution. Then the sill of the Gauss model was adjusted to respect the variance of the raw data using a cross validation procedure. Thus, it has been proposed to use a slightly anisotropic model distribution where the longer axis is oriented N190/-50 (dip direction); the second axis is oriented N100/00 (strike direction); and the third axis perpendicular to the idealized plane of the majority of the veins is oriented N010/-40. The retained model involves a nugget effect and two spherical structures. When adjusted to raw data, the sill of the nugget effect is 17.7. The first spherical structure has a sill of 19.2 and the second has a sill of 6.1. The ranges for the first and second spherical structures within the dip direction are 8 and 50 m. In the strike direction these two ranges were set to 5 and 35, respectively. Finally, the perpendicular axis was modelled with a range of 6 and 16. Figure 14.14 shows the experimental and model variogram obtained from the Gauss transformation.

www.innovexplo.com


Figure 14.14 – Directional variography of the Gauss transformed data.




A block model was established to cover the entire drilled area. The origins of the block model are as follows (center of the front, bottom, left cell): Easting: 296160 mE (24 cells x 10 m each) Northing: 5329100 mN (19 cells x 10 m each) Elevation: -660 m (80 cells x 10 m each) The block model was not rotated, and the error on volume calculated from the block model compared to the volume estimated from the wireframe is lower than 0.02%.

www.innovexplo.com



All blocks with at least 0.001% of their volume falling within a selected solid (High Probability Ore) were assigned the “In Ore” code. A percent block model was generated, reflecting the proportion of zone in each block.


For each cell the grade was estimated using the OK method. The search ellipsoid used to select samples for interpolation has exactly the same geometry as the variography model defined in section 14.3.6. This interpolation process required a minimum of 3 composites (the 3 closest) and in this case the 10 closest were retained. Almost all cells were estimated using these parameters. When the search ellipsoid cannot find the minimum of three samples, it was then extended from 35X50X16 to 60X60X16 keeping the same minimum and maximum sample number requirement. No other restriction parameters (octants, parent hole, distance) were used.


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: the part of a Mineral Resource for which quantity, grade or quality, densities, shape, and 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: the 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: the 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. Resources from this category should not be used to support mine planning and evaluation of the economic viability of the deposit.

www.innovexplo.com


The Indicated category was assigned to each cell where the grade was estimated using the first pass ellipsoid. The Inferred category was assigned to all cells estimated with the extended search ellipsoid. No other adjustment was made to this classification scheme. No measured category was assigned to any blocks.


By definition, the cut-off is the breakeven point considering total cost and revenue generated by the operation. The current estimation used the following economic parameters. Gold value = $1,450/oz Mining cost = $30/t (metric ton) Milling fees = $25/t (metric ton) Gold recovery = 92% These parameters give an estimation of the cut-off at 3.0 g/t. Even if the author uses a cut-off of 3.0 g/t for official estimation, other cut-off grade results were also compiled for illustrative purposes (Table 14.9). The cut-off grade must be re-evaluated in light of prevailing market conditions and other factors, including gold price, exchange rate, mining method, related costs, etc. This cut-off value is based on a standard mining method using a relatively small stope. Due to the nature of the mineralized area (complex distribution of narrow veins within a barren host rock), the interpolation was done using an unconstrained selection of composites. This has a large impact on selectivity of the “ore” and “waste” smaller mining units (“SMU”). In such cases, the selectivity is highly affected by the size of the reference volume. The key for successful mining of this material will require a stope design large enough to reduce mining costs (long hole mining) and the ability to discriminate “ore” and “waste” pockets within that large stope (ore sorting). Figure 14.15 shows the effect of the SMU size on resource estimation.

www.innovexplo.com


Table 14.9 – Sensitivity of Indicated and Inferred categories to different cut-off values

On Figure 14.15, it is clear that applying the OK method to a 5X5X5 or to a 10X10X10 array of cells are almost identical and do not make any valuable difference in terms of selectivity. Selectivity on a 10X10X10 cell is much better than on a larger stope. However, when compared with the theoretical selectivity obtained from point data, it is clear that a sorting process prior to milling can be extremely valuable. It may represent up to 100,000 more ounces at 1.3 g/t cut-off.

www.innovexplo.com


Figure 14.15 – Selectivity capability according to SMU size 14.3.13. Mineral Resource Estimate Results

Given all the parameters listed above, GéoPointCom is of the opinion that the current 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. If we consider a 3.0 g/t cut-off, the current estimation represents 1,325,120 tonnes at 5.6 g/t Au (237,450 oz Au) in the Indicated category (Table 14.10).

www.innovexplo.com


Table 14.10 – Mineral Resources of No. 4 Plug

www.innovexplo.com



Comparing mineral resource estimates may appear to be a simple exercise, but in reality, obtaining real conclusions from such comparisons is difficult due to differences in key assumptions, parameters and methods and the interaction between these features in the final results. In the current case, the historical estimate made by D.R. Scammell in May 1989 is not 43-101 compliant. His estimation was designed for a very selective mining method, which defined only twelve (12) individual veins. The volume was estimated with polygon on section and grade is derived from the nearest neighbour composited at the midpoint of each intersection. Most of the key assumptions are not described in the supporting report. We can only assume that the grade was capped at 1 oz/t (34.286 g/t). At that time, the resource was estimated at 1,042,800 tonnes at 6.13 g/t Au for a total of 205,760 ounces of gold. This could not be compared with the current estimate because the likely mining method and selectivity are too different.

14.4. Resource Estimate of the Triangle Zone

Estimation was performed from July to September 2013. The database used was completed April 24, 2013. The last hole included in the resource calculation is TM-13-15. The effective date for the resource estimation of the Triangle zone is April 24, 2013.

14.4.1. Methodology

The Mineral Resource Estimate and geostatistical study detailed in this report was performed using the Isatis (V.2013.1) software package. The method involves a 3D Block Model estimated with an Inverse Power distance interpolator.

14.4.2. Drill hole Sample Database

The Geotic/MS Access diamond drill hole database is maintained and updated by Geologica employees under the supervision of Alain-Jean Beauregard, P.Geo. and Daniel Gaudreault, Eng. It contains 84 surface diamond drill holes with conventional analytical gold assay results, as well as coded lithologies from the drill core logs. The 84 drill holes yielded 25,970 m of core. A total of 7,861 saw-cut and assayed samples for a total of 6,956 m. The database also contains 829 QA/QC samples.


The interpretation was completed by Benjamin Blaise of Geologica under the supervision of Alain-Jean Beauregard and Daniel Gaudreault. In order to conduct accurate resource modelling of the Triangle zone, the author constructed a wireframe solid representing eight (8) individual veins oriented more or less N116/-38. The wireframe solid was created by digitizing an interpretation onto 18 sections spaced 25 m apart, and then using tie-lines to complete the wireframes for each individual vein (Figure 14.16). For each vein intersection, the minimum true thickness was set to 2 m. When true thickness was less than the minimum thickness, internal dilution was included to achieve the minimum thickness.

www.innovexplo.com


–

Figure 14.16 – Wireframe solids from the Triangle Zone.

www.innovexplo.com



It is common in the industry to remove some of the highest (aberrant) values from the assays distribution prior to making composites. The main objective of this process is to make sure that erroneous values cannot affect grade estimation. With the development of statistical methods for estimating grades, this process became less important. In fact, Kriging and especially simulation techniques are less sensitive to some occasional very high values. On the other hand, the presence of some very high values may make the variogram difficult to establish. Geostatistics provide some efficient tools to solve this problem. The more common methods used to assess the necessity of using a capping value are listed below:

The first indication of the necessity to cap high values is the coefficient of variation “CV”. Ideally, this should be located close to 1. A CV above 2 is generally considered as an indication that high values should be capped. In the case of very high CV, uncapped grades may make it difficult to produce a clean variogram.


should form a relatively straight line. A positive break in the upper end is often interpreted as an indication that high values should be capped. This criterion is probably the strongest indication, especially when the interpolation method requires normal or log–normal distribution.

The metal factor method consists of comparing the cumulative metal


Figure 14.17 shows comparative distribution (density, histogram, box plot and probability plot) of gold values for all samples within the Vein Selection. From this graph it is clear that Vein Selection samples do not form a simple normal or log-normal population. It seems to be a bi-modal distribution in which the upper part (50% of the population) may follow a log-normal distribution without any positive break at the higher grades. The lower part (lower grades) shows a small bump corresponding to the higher density of the “waste” distribution. It may be caused by the inclusion of samples that really belong to the “waste” population within the “ore” population instead. The distribution of the metal contents within the percentile classes suggests the use of a cap at 75 g/t. The relatively low value for the CV (2.65) may not argue in favor of a capping grade. Figure 14.18 shows the difference in gold content in the Indicated category when calculated at a 3 g/t cut-off using different capping grades (prior to compositing). This graph demonstrates that 15% of the ounces are controlled by 1% of the samples. This is often the case for this type of mineralization. It is also evident that the slope of the reduced ounces changes abruptly when the capping is over 80 g/t prior to compositing. When the grade is over 80 g/t, the author suggests using the following formula:

Capped value = 80 + 0.5(original grade – 80)

www.innovexplo.com


Figure 14.17 – Probability plot.

Figure 14.18 – Effect of capping in resource estimation.

www.innovexplo.com


14.4.5. Compositing

Compositing was compiled using the grade of the adjacent material when assayed, or a value of zero when not assayed. In order to minimize any bias introduced by the variable sample lengths, uncapped gold assays (as determined in the previous subsection) were composited to 1.0 m equal lengths (“1.0m composites”) within all intervals that define each of the mineralized zones (Vein Selection). Tails were not created; they were instead distributed over the previous composites.

14.4.6. Variography

The author modelled the composite variography considering all veins as one uniform group using Isatis software package. The objectives of this step consist of:



are required during the Kriging interpolation. These equations take into account the nugget effect (C0), the model of dispersion (i.e. spherical, exponential, etc.), the range and the variance in each of the three axes (σ1, σ2 and σ3). It is common to use more than one model in order to better represent the dispersion (short and long) of each of the axes before reaching the level (total variance).

An omnidirectional variogram was first calculated to estimate the nugget effect (C0). Eighteen (18) directional variograms located in the horizontal plane were then calculated which allows the direction of longer continuity (longitudinal direction) to be defined. The next step includes 18 directional variograms set in the vertical plane parallel to the longitudinal direction and another 18 directional variograms set in a vertical plane perpendicular to the longitudinal direction. The final step was 18 directional variograms on an inclined plane representing the best fit for all mid points (longitudinal plane). This step provides an overview of the variance of the idealized sphere in order to verify the presence of a directional anisotropy. The variograms have not demonstrated a clear anisotropy. Thus it has been proposed to use an omnidirectional variogram as shown in Figure 14.19.

www.innovexplo.com


Figure 14.19 – Omnidirectional variography.




A block model was established to cover the entire drilled area. The origins of the block model are as follows (center of the front, bottom, left cell): Easting: 296212mE (92 cells x 5 m each) Northing: 5328323mN (97 cells x 5 m each) Elevation: -207m (104 cells x 5 m each) The block model was not rotated and the error on volume calculated from the block model compared to the volume estimated from the wireframe is lower than 0.23%.




For each cell identified as within one of the eight (8) veins, the grade was estimated using only the composites from the specific vein (“Hard Boundary”). This is one of the major changes made since the previous estimation. All estimations were done using Inverse Distance Squared (IPD2).

www.innovexplo.com


When interpolating Ore Grade, the search ellipsoid (50 m radius), as defined in section 14.4.6, must contain a minimum of 2 composites (the 2 closest) and in this case the 10 closest points were retained. No other restriction parameters (octants, parent hole, or distance) were used.


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: the 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: the 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: the 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. Resources from this category should not be used to support mine planning and evaluation of the economic viability of the deposit. The Indicated category was assigned to each cell where a minimum of 2 composites can be found within a radius of 25 m. An Inferred category was assigned to all other cells. No measured category was assigned to any blocks.

www.innovexplo.com



By definition, the cut-off is the breakeven point considering total cost and revenue generated by the operation. The current estimation used the following economic parameters. Gold value = $1,450/oz Mining cost = $90/t (metric ton) Milling fees = $25/t (metric ton) Gold recovery = 92% These parameters give a first estimation of the cut-off at 2.27 g/t. It is clear that the parameters do not represent total costs. Environmental, developmental, administration and associated costs were not taken into account. Another 25% may be added to the cost previously considered. Doing so would push the estimation of the cut-off to 3.35 g/t. The author chose a cut-off of 3.0 g/t for the official estimation. Other cut-off grades were also used, but for illustrative purposes only. The cut-off grade must be re-evaluated in light of prevailing market conditions and other factors: gold price, exchange rate, mining method, related costs, etc. (Table 14.11).

Table 14.11 – Sensitivity of the resource estimation to different cut-off values


Given all parameters listed above, GéoPointCom is of the opinion that the current 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. If we consider a 3.0g/t cut-off, the current estimation will represent 599,700 tonnes at 9.9 g/t (190,700 oz Au) in the Indicated category and 332,300 tonnes at 12.9 g/t Au (137,600 oz Au) in the Inferred category (Table 14.12).

www.innovexplo.com


Table 14.12 – Mineral Resources of the Triangle Zone

www.innovexplo.com


14.4.14. Comparison with Previous Mineral Resource Estimates

Comparing mineral resource estimates may appear to be a simple exercise, but in reality, obtaining real conclusions from such comparisons is difficult due to differences in key assumptions, parameters and methods and the interaction between these features in the final results. In the current case, historical estimates were not 43-101 compliant and will not be commented on. In October 2011, GéoPointCom was first involved in resource estimation for the Triangle Zone. At that time interpretation of the Triangle Zone was uncertain. The zones were interpreted as six (6) subvertical parallel veins oriented more or less N100/-65; tension veins were also recognized but were very difficult to follow. At that time the model selected was qualified as uncertain and six (6) additional drill holes were recommended to verify the model and check if a horizontal model might be more appropriate. The block model was estimated using an OK estimator and a search ellipsoid of 50 m radius laid on idealized planes (perpendicular axis of 25 m). The previous resource estimation used a 3.0 g/t Au cut-off. At that time, Indicated resources, regardless of elevation, were estimated at 195,900 tonnes at 10.9 g/t Au (68,540 oz Au). Inferred resources were estimated at 246,200 tonnes containing 17.2 g/t Au (136,080 oz Au). In December 2012, additional drilling led to a more comprehensive geological interpretation, a refinement of the geological model, and new parameters for interpolating grades. The additional drilling resulted in a global decrease of the grade in respect of the number of ounces, but the total tonnage of this zone was slightly higher. The number of tonnes in the Indicated category reported at a 3 g/t cut-off was increased by 71% compared to the 2011 estimation. The grade was lowered (-11%) resulting in an enhancement of the total number of ounces by a factor of 54%. When considering the Inferred category, the tonnes, grades and ounces were lowered by -14%, -31% and -41% respectively. The interpretation of the mineralized structures and veins from vertical to horizontal veins, between 2011 and 2013, may raise some questions regarding the robustness of the geological interpretation. According to Geologica, the horizontal interpretation was clearly confirmed by the addition of strategic drill holes located to verify this recent interpretation and the new model. In April 2013, additional drilling allowed the author and Geologica to gain confidence in the horizontal model rather than the vertical one. The hard boundary used for sample selection and the decrease in the number of samples used for interpolating are all factors influencing the significant gain observed in the Indicated category (3 times more ounces).

14.5. Resource Estimate – Vein No. 6

The estimation was performed from September to November 2013. The database used was completed on August 17, 2012. The last drill hole included and considered was V6-12-12A.

www.innovexplo.com


14.5.1. Methodology

The Mineral Resource Estimate and geostatistical study detailed in this report was performed using Isatis (V.2013.1) software. The method involves a 3D Block Model estimated with Kriging.

14.5.2. Drill hole sample database

The Geotic/MS Access diamond drill hole database is maintained and updated by Geologica employees under the supervision of Alain-Jean Beauregard, P.Geo. and Daniel Gaudreault, P.Eng. It contains 57 surface diamond drill holes with conventional analytical gold assay results, as well as coded lithologies from the drill core logs. The 57 drill holes yielded 12,436 m of core, with 2,534 samples for a total of 2,378 m of saw-cut and assayed core. The database also contains 268 QA/QC samples.


The interpretation was completed by Benjamin Blaise of Geologica under the supervision of Alain-Jean Beauregard and Daniel Gaudreault. In order to conduct accurate resource modelling of Vein No. 6, the author constructed a wireframe solid representing three individual veins oriented N054/-30 (Vein A), N073/-30 (Vein B) and N082/-40 (Vein C). The wireframe solid was created by digitizing an interpretation onto twelve (12) sections spaced 25 m apart, and then using tie-lines to complete the wireframes for each individual vein (Figure 14.20). For each vein intersection, the minimum true thickness was set to 2 m. When true thickness was less than the minimum thickness, internal dilution was included to achieve the minimum thickness.

www.innovexplo.com


Figure 14.20 – Wireframe solids from Vein No. 6

14.5.4. High grade capping

It is common in the industry to remove some of the highest (aberrant) values from the assays distribution prior to compositing the samples. The main objective of this process is to make sure that the erroneous values could not affect grade estimation. With the development of statistical methods for estimating grade, this process became less important. In fact, Kriging, and especially simulation techniques, are less sensitive to some occasional very high values. On the other hand, the presence of some very high values may make the variogram difficult to establish. Geostatistics provide some efficient tools to solve this problem. The more common methods used to assess the necessity of using a capping value are listed below:

The first indication of the necessity to capping high values is the coefficient of variation “CV”. Ideally, this should be located close to 1. A CV above 2 is generally considered an indication that high values should be capped. In the case of very high CV, uncapped grades may make it difficult to produce a clean variogram.

In the case of a simple normal or log-normal population, the probability curve should form a relatively straight line. A positive break in the upper end is often interpreted as an indication that high values should be capped. This

www.innovexplo.com


criterion is probably the strongest indication, especially when the interpolation method is based on normal or log–normal distribution.

The metal factor method consists of comparing the cumulated metal percentage with the cumulated data percentage. This technique takes for granted that all samples represent an equivalent number of tonnes. Specialists generally agree to keep this factor below 10. In other words, there should never be more than 10% of ounces in less than 1% of tonnes. This method really concentrates on the upper 1% or 2% of the distribution. It is more a security factor than a distribution analysis.

Figure 14.21 shows comparative distribution (density, histogram, box plot and probability plot) of gold values for all samples within each vein. From this, one can note that sample distribution from each vein may be assimilated to a log-normal distribution. The distribution of metal contents within the percentile classes would suggest the use of a capping at 25 g/t. The relatively low value for the CV (2.68) may not argue in favor of a capping grade. Figure 14.22 shows the effect of different capping grades (prior to compositing; total ounces estimated at 3 g/t cut-off). This graphic suggest the presence of two different populations. It is also evident that at capping grades over 40 g/t, the slope of percentage of ounces lost changes abruptly. To make it smoother, it is recommended to gradually limit the higher gold values (over 40 g/t) prior to compositing. When the grade is over 40 g/t, the author suggests using the following formula: Capped value = 40 + 0.53(original grade – 40) This capping grade will affect only three (3) samples (1.8% of the selected samples) and will reduce the total amount of ounces by 3.3%.

www.innovexplo.com


Figure 14.21 – Probability plot (Vein No. 6)

www.innovexplo.com


Figure 14.22 – Effect of capping on resource estimation (Vein No. 6)

www.innovexplo.com


14.5.5. Compositing

Compositing was compiled using the grade of the adjacent material when assayed, or a value of zero when not assayed. In order to minimize any bias introduced by the variable sample lengths, capped gold assays (as determined in previous subsection) were composited to 1.0 metre equal lengths (“1.0m composites”) within all intervals that define each of the mineralized zones. Tails were not created; they were instead distributed over the previous composites.

14.5.6. Variography

The author modelled the composite variography considering all the veins as one uniform group using Isatis software. The objectives of this step consist of:



are required during the Kriging interpolation. These equations take into account the nugget effect (C0), the model of dispersion (spherical, exponential), the range and the variance in each three axes (σ1, σ2 and σ3). It is common to use more than one model in order to better represent the dispersion (short and long) of each of the axes before reaching the level (total variance).

An omnidirectional variogram was first calculated to estimate the nugget effect (C0). Second, 18 directional variograms located in the horizontal plane were then calculated. This allows defining the direction of longer continuity (longitudinal direction). The next step includes 18 directional variograms set in the vertical plane parallel to the longitudinal direction and another 18 directional variograms set in a vertical plan perpendicular to the longitudinal direction. The final step involved 18 directional variograms on an inclined plan representing the best plan fit for all mid points (longitudinal plan). This step allows having an overview of the variance of the idealized sphere in order to verify the presence of a directional anisotropy. The variography was very difficult to establish. In fact all structures were hidden by only one sample (sample number P174683 with a grade of 94.3 g/t). When this sample is removed, the variograms became more clear, not perfect but usable. This study did not demonstrate a clear anisotropy. Thus, it was proposed to use an omnidirectional variogram as shown in Figure 14.23.

www.innovexplo.com


44

131

147

415 435

284 323

480521

473

0

0

25

25

50

50

75

75

100

100

125

125

Distance (m)

Distance (m)

0 0

10 10

20 20 Variogram : Au

Variogram : Au

Figure 14.23 – Omnidirectional variography (Vein No. 6)


A specific gravity of 2.80 g/cm3 was used to estimate the tonnage for the previously 2012-2013 Mineral Resource Estimate.


A block model was established to cover the entire drilled area. The origins of the block model are as follows (center of the front, bottom, left cell): Easting: 293250mE (32 cells x 10 m each) Northing: 5330000mN (45 cells x 10 m each) Elevation: -100m (41 cells x 10 m each) The block model was not rotated and the error on volume calculated from the block model compared to the volume estimated from the wireframe is lower than 0.04%.



www.innovexplo.com


14.5.10. Grade block model

For each cell identified as within one of the three (3) veins, the grade was estimated using only composites from the specific vein (hard boundary). All estimations were completed using a Kriging function based on the variogram presented in Figure 14.21. When 4 composites can be found within a radius of 50 meters, the 8 closest composites (within a search radius of 50 m) were used for estimating the grade of the cell. If the minimum number of composites cannot be found within the first search ellipsoid, then the search was extended to 100 m. Again, if a minimum of 4 composites can be found within the second search ellipsoid (100 m) then the 8 closest points are used for block estimate. All cells, where the minimum number of composites cannot be found even within the second search ellipsoid, will remain un-estimated. No other restriction parameters (octants, parent hole, distance) were used.

14.5.11. Resource categories

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. Resources from this category should not be used to support mine planning and evaluation of the economic viability of the deposit.

www.innovexplo.com


The Indicated category was assigned to each cell where the first search ellipsoid was used. For each cell where the second search ellipsoid was used, the Inferred category was assigned. No measured category was assigned to any blocks.

14.5.12. Minimum cut-off value

By definition, the cut-off is the breakeven point considering total cost and revenue generated by the operation. The current estimation used the following economical parameters.

Gold value = 1 450$/oz Mining cost = 90$/t (metric ton) Milling fees = 25$/t (metric ton) Gold recovery = 92%

These parameters give a first estimation of the cut-off at 2.27 g/t. However, it is clear that the parameters do not represent the total cost. Environmental, development, administration costs, etc. were not taken into account. Another 25% may be added to the costs previously considered. Doing so will push the estimation of the cut-off to 3.35 g/t. A cut-off of 3.0 g/t for the official estimation is proposed. Other cut-off grades were also compiled, but for illustrative purposes only. The cut-off grade must be re-evaluated in light of prevailing market conditions and other factors, including gold price, exchange rate, mining method, related costs, etc. (Tables 14.13 to 14.14).

www.innovexplo.com


Table 14.13 – Cut-off sensitivity on resource estimation (Vein No. 6)

Ton Grade Onces Ton Grade Onces Ton Grade Onces Ton Grade Onces

0.0 402 500 3.6 46 070 324 900 2.3 23 610 105 200 7.7 26 010 832 600 3.6 95 680

0.5 384 100 3.7 45 820 217 700 3.3 22 960 105 200 7.7 26 010 707 000 4.2 94 780

1.0 334 500 4.2 44 630 160 000 4.2 21 560 105 200 7.7 26 010 599 700 4.8 92 200

1.5 292 700 4.6 42 920 126 900 5 20 280 105 200 7.7 26 010 524 800 5.3 89 200

2.0 260 600 4.9 41 140 111 700 5.4 19 460 105 200 7.7 26 010 477 400 5.6 86 600

2.5 221 900 5.4 38 310 98 600 5.8 18 520 105 200 7.7 26 010 425 700 6.1 82 830

3.0 198 600 5.7 36 210 85 600 6.3 17 340 105 200 7.7 26 010 389 400 6.4 79 550

3.5 173 000 6 33 550 73 900 6.8 16 110 105 200 7.7 26 010 352 200 6.7 75 670

4.0 138 600 6.6 29 400 59 000 7.6 14 330 105 200 7.7 26 010 302 800 7.2 69 740

4.5 117 200 7 26 480 51 200 8.1 13 260 105 200 7.7 26 010 273 600 7.5 65 750

5.0 93 600 7.6 22 900 46 400 8.4 12 530 105 200 7.7 26 010 245 200 7.8 61 440

Ton Grade Onces Ton Grade Onces Ton Grade Onces Ton Grade Onces

0.0 60 400 3.4 6 500 60 300 1.4 2 620 70 100 7.7 17 250 190 800 4.3 26 370

0.5 56 200 3.6 6 450 22 200 3.5 2 470 70 100 7.7 17 250 148 500 5.5 26 160

1.0 45 100 4.3 6 180 14 000 5.1 2 310 70 100 7.7 17 250 129 200 6.2 25 730

1.5 41 000 4.6 6 020 13 300 5.3 2 280 70 100 7.7 17 250 124 500 6.4 25 550

2.0 38 200 4.8 5 860 13 300 5.3 2 280 70 100 7.7 17 250 121 600 6.5 25 390

2.5 32 500 5.2 5 450 13 300 5.3 2 280 70 100 7.7 17 250 115 900 6.7 24 980

3.0 28 400 5.6 5 090 13 100 5.4 2 260 70 100 7.7 17 250 111 600 6.9 24 590

3.5 25 600 5.8 4 800 12 400 5.5 2 180 70 100 7.7 17 250 108 100 7 24 220

4.0 22 400 6.1 4 420 11 700 5.6 2 100 70 100 7.7 17 250 104 200 7.1 23 760

4.5 20 000 6.4 4 090 7 900 6.2 1 590 70 100 7.7 17 250 98 100 7.3 22 930

5.0 17 100 6.6 3 650 6 200 6.6 1 330 70 100 7.7 17 250 93 400 7.4 22 220

Vein A Vein B Vein C Total (all 3 veins)Cut Off grade

Cut Off sensitivity on resource estimation

Indicated (Capped gradually over 40 g/t)

Inferred (Capped gradually over 40 g/t)

Cut Off gradeVein A Vein B Vein C Total (all 3 veins)

www.innovexplo.com


Table 14.14 – Resources Estimates for Vein No. 6 calculated using progressive capping above 40 g/t (cut-off grade of 3 g/t)

Ton Grade Onces Ton Grade Onces

285 <100 4.7 <10 0 0 0

255 29 100 4.1 3 870 300 3.1 30

225 79 700 5 12 760 5 300 6.6 1 140

195 118 300 7 26 690 12 500 5.5 2 210

165 45 400 5.7 8 240 5 500 5.7 1 010

135 11 800 5.3 2 010 17 600 5.2 2 920

105 0 0 0 300 4.4 50

75 0 0 0 0 0 0

45 0 0 0 100 7.6 20

15 13 300 7.6 3 230 18 100 7.6 4 410

‐15 44 800 7.6 10 940 0 0 0

‐45 47 100 7.8 11 840 4 700 7.8 1 190

‐75 0 0 0 46 000 7.7 11 320

‐105 0 0 0 1 300 7.6 310

Total 389 400 6.4 79 550 111 600 6.9 24 590

Bench toe

(m)

Indicated Inferred


Given all parameters listed above, GéoPointCom is of the opinion that the current 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. If we consider a 3.0 g/t cut-off, the current estimation for Vein No. 6 will represent 389,400 tonnes at 6.4 g/t (79,550 oz Au) in the Indicated category. These mineral resources were excluded from the current PEA.

14.5.14. Comparison to previous mineral resource estimates

The historical estimate made by Teck in 1986 was based on only 5 holes. (Blecha, june 1986) This report does not contain any details on technical parameters used except the minimum width fixed to 6 feet (1.83m). At that time Teck estimate the resources of the vein 6 to 106 100 metric ton at 6.9 g/t. To GeoPointCom knowledge no other estimate was done on this zone. Thus this estimate cannot be compared to the actual estimate. The 1986 Teck estimate dos not qualify as a mineral Resource according to the standards of NI43-101.

www.innovexplo.com


14.6. Resource Estimate of the Sixteen Zone

The estimation was performed from November to December 2013. The database used was the same as that used for the last resource estimation in April 2012, except that 254 samples were added to fill the gaps in most of the drill holes completed since 2004. The April 2012 resource estimation includes the last drill holes and takes into account SX-11-03. The effective date for the resource estimation of the Sixteen Zone is November 18, 2013.

14.6.1. Methodology

The Mineral Resource Estimate and geostatistical study detailed in this report was performed using Isatis (V. 2013.1) software. The method involves a 3D Block Model estimated with an ordinary kriging (OK) interpolator on selected samples which are strictly constrained to an area of high vein density within a felsic intrusive plug.

14.6.2. Drill hole sample database

The Geotic/MS Access diamond drill hole database is maintained and updated by Geologica employees (P.Geo and P.Tech.) under the supervision of Alain-Jean Beauregard, P.Geo. and Daniel Gaudreault, P.Eng. It contains 63 surface diamond drill holes with conventional analytical gold assay results, as well as coded lithologies from the drill core logs. The 63 drill holes yielded 14,073.6 m of core, and 3,413 samples for a total of 2,801.5 m of saw-cut and assayed core. The database also contains 180 QA/QC samples.

14.6.3. Interpretation of mineralized zones

The interpretation was completed by Geologica, under the supervision of Alain-Jean Beauregard and Daniel Gaudreault. In order to conduct accurate resource modelling of the Sixteen Zone, the retained option consists of constructing a solid wireframe representing an area within the dioritic intrusive where vein density containing higher grade is more abundant than the surrounding intrusive host rock (High Probability Ore). The wireframe solid was created by digitizing an interpretation onto fifteen (15) sections spaced 25 m apart, and then using tie-lines to complete the wireframes composed of four (4) closed areas (Figure 14.24). This wireframe (High Probability Ore) was used for selecting all drill hole intersections. Unsampled areas were considered to carry a grade of 0.000 g/t Au.

www.innovexplo.com


Figure 14.24 – Wireframe solids of the Sixteen Zone.

www.innovexplo.com



It is common in the industry to remove some of the highest (aberrant) values from the assay distribution prior to compositing the samples. The main objective of this process is to make sure that the erroneous values could not affect grade estimation. With the development of statistical methods for estimating grade, this process became less important. In fact, Kriging and especially simulation techniques are less sensitive to some occasional very high values. On the other hand, the presence of some very high values may make the variogram difficult to establish. Geostatistics provide some efficient tools to overcome this problem. The more common methods used to assess the necessity of using a capping value are listed below:

The first indication of the necessity to cap high values is the coefficient of variation “CV”. Ideally, this value should be close to 1. A CV above 2 is generally considered an indication that high values should be capped. In the case of very high CV, uncapped grades may make it difficult to produce a clean variogram.

In the case of a simple normal or log-normal population, the probability curve should form a relatively straight line. A positive break in the upper end is often interpreted as an indication that high values should be capped. This criterion is probably the strongest indication, especially when the interpolation method requires a normal or log–normal distribution.

The metal factor method consists of comparing the cumulated metal percentage with the cumulated data percentage. This technique takes for granted that all samples represent an equivalent number of tonnes. Specialists generally agree to keep this factor below 10. In other words, there should never be more than 10% of ounces in less than 1% of tonnes. This method really concentrates on the upper 1% or 2% of the distribution. It is more a security factor than a distribution analysis.

Especially when the grades of all the intersections are not the most important parameters for marking the zone, it is always possible to proceed with a full resource estimation using different capping grades. Then a graph showing the percentage of total ounces lost relative to the capping grade used may point out a change in the gold distribution.

Figure 14.25 and 14.27 show the distribution (density, histogram, box plot, probability plot and the percentage of ounces lost relative to capped samples) of gold values for all samples within the High Probability Ore. From this graph it is clear that High Probability Ore samples do not form a simple normal or log-normal population. It seems to be a bi-modal distribution. One can propose an assemblage of two log-normal populations to explain this distribution. The waste population may represent close to 75% of the samples with a median close to 0.1 and grades ranging from 0.002 to 1 g/t. In that case the ore population may have a median around 5 g/t and grades ranging from 0.1 to 70 g/t. The distribution of metal contents within the percentile classes show more than 24% of the total metal appearing within 1% of the samples. To satisfy the metal factor methodology, the grade must be capped at 15 g/t. This is considered to overly conservative and unrealistic for the observed bi-modal distribution. Based on the graph showing the reduction of total ounces relative to the capping grade, the author recommends limiting higher gold values at 35 g/t. This capping limit will affect six (6) samples, drive the CV to 2.7,

www.innovexplo.com


and reduce the total amount of gold by 10%. At this point we may presume that the high CV will make variography very difficult to establish.

Figure 14.25 – Capping grade distribution versus % ounces lost (Sixteen Zone)

www.innovexplo.com


Figure 14.26 – Grade distribution within the High Probability Ore (Sixteen Zone)

14.6.5. Compositing

In order to minimize the bias introduced by the variable sample lengths, capped gold assays (as determined in a previous subsection) were composited to 0.7 metre equal lengths (“0.7m composites”) within all intervals that define the High Probability Ore. Tails were not created; they were instead distributed over the previous composites.

14.6.6. Variography

The composite variography of the High Probability Ore was modelled using Isatis software. The objectives of this step consist of:

Characterizing the anisotropy and setting the dimensions and direction of a search ellipsoid to use during the interpolation of values of the block model. The dimensions of the three axes of the ellipsoid are equal to the range measured on the variograms. This way, it is statistically shown that wherever the center of this ellipsoid is located, all samples included within it will have a variance inferior to that of the entire population. It is reasonable to use this subselection to estimate the value of a central point. The anisotropy may also be used to select and weigh samples during the “de-clustering” steps.

www.innovexplo.com


Defining the Kriging equations. The equations deriving from the variograms are required during the Kriging interpolation. These equations take into account the nugget effect (C0), the model of dispersion (spherical, exponential, cubic, etc.), the range and the variance in each three axes (σ1, σ2 and σ3). It is frequent to use more than one model in order to better represent the dispersion (short and long) of each of the axes before reaching the level (total variance).

An omnidirectional variogram was first calculated to estimate the nugget effect (C0). Second, 18 directional variograms located in the horizontal plane were calculated permitting to define the direction of longer continuity (longitudinal direction). The next step includes 18 directional variograms set in the vertical plane parallel to the longitudinal direction and another 18 directional variograms set in a vertical plan perpendicular to the longitudinal direction. The final step involved 18 directional variograms on an inclined plan representing the best plan fit for all mid points (longitudinal plan). This step allows having an overview of the variance of the idealized sphere in order to verify the presence of a directional anisotropy. The author was unable to model a clear geometrical anisotropy, thus an omnidirectional model was created in which the Nugget effect counts for 78% of the total variance; a cubic model with a range of 72 m was used. Figure 14.27 shows the experimental and model variogram obtained from the Sixteen Zone composites.

www.innovexplo.com


0

0

10

10

20

20

30

30

40

40

50

50

60

60

70

70

Distance (m)

Distance (m)

0 0

1 1

2 2

3 3

4 4

5 5

6 6

7 7

8 8

9 9

10 10 Variogram : Au25

Variogram : Au25

Figure 14.27 – Omnidirectional variography of the composites (Sixteen Zone)




A block model was established to cover the entire drilled area. The origins of the block model are as follows (center of the front, bottom, left cell): Easting: 292260 m E (20 cells x 10 m each) Northing: 5329560 m N (9 cells x 10 m each) Elevation: 40 m (20 cells x 10 m each) The block model was not rotated and error on the volume calculated from the block model was compared to the volume estimated from the wireframe revealing a variation lower than 0.01%.

www.innovexplo.com



All blocks with at least 0.001% of their volume falling within a selected solid (High Probability Ore) were assigned the “In Ore” code. A percent block model was generated reflecting the zone proportion in each block.

14.6.10. Grade block model

For each cell the grade was estimated using an ordinary kriging (OK) procedure. The search ellipsoid use to select samples for interpolation was set to a radius of 15 m. This interpolation process required a minimum of 5 composites (the 5 closest) and in this case the 10 closest were retained. Almost all (587 cells on a total of 614) cells were estimated using these parameters. For the cells where the minimum number of composites could not be found (27 cells of a total of 614), the radius of the search ellipsoids was extended to 60 m. No other restriction parameters (octants, parent hole, distance) were used.

14.6.11. Resource categories

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. Resources from this

www.innovexplo.com


category should not be used to support mine planning and evaluation of the economic viability of the deposit.

The Indicated category was assigned to each cell where the grade was estimated using the first pass ellipsoid. An Inferred category was assign to all cells estimated with the extended search ellipsoid. No other adjustment was made to this classification scheme. No measured category was assigned to any blocks.

14.6.12. Minimum cut-off value

By definition, the cut-off is the breakeven point considering total cost and revenue generated by the operation. The current estimation used the following economical parameters.

Gold value = $1,450/oz Mining cost = $90/t (metric ton) Milling fees = $25/t (metric ton) Gold recovery = 92%

These parameters give a first estimation of the cut-off at 2.27 g/t Au. However, it is clear that the parameters do not represent total cost. Environmental, development and administration costs, etc. were not taken into account. Another 25% may be added to the cost previously considered. Doing so will push the estimation of the cut-off to 3.35 g/t Au. The author suggests a cut-off at 3.0 g/t Au for the official estimation (see Table 14.15). Other cut-off grades were also compiled, but for illustrative and comparison purposes only. The cut-off grade must be re-evaluated in light of prevailing market conditions and other factors: gold price, exchange rate, mining method, related costs, etc.


Given all parameters listed above, GéoPointCom is of the opinion that the current 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. If we consider a 3.0 g/t Au cut-off, the current estimation represents 91,700 tonnes at 5.2 g/t Au (15,440 oz Au) in the Indicated category and 1,800 tonnes at 4.2 g/t Au (250 oz Au) in the Inferred category (Table 14.15). These mineral resources were excluded from the current PEA.

www.innovexplo.com


Table 14.15 – Cut-off sensitivity on the resource estimation (Sixteen Zone)

14.6.14. Comparison to previous mineral resource estimates

Comparing mineral resource estimates may appear to be a simple exercise, but in reality, obtaining real conclusions from such comparisons is difficult due to differences in key assumptions, parameters and methods and the interaction between these features in the final results. The historical estimate made by D.R. Scammell in May 1989 is not NI 43-101 compliant. His estimation was designed for a very selective mining method. The volume was estimated with polygon on section and grade was derived from the nearest neighbor composited at the midpoint of each intersection. Most of the key assumptions are not described in the supporting report. We can only assume that the grade was capped at 1 oz/t (34.286 g/t). This could not be compared with the current estimate because the mining methods and selectivity are too different. The last estimation undertaken by GéoPointCom in April 2012 was based on a similar method. At that time, the 2004 drilling campaign had many gaps (intervals not sampled within the ore selection). This situation has been corrected in the updated estimation herein. The present estimation is also much more selective in respect of High Probability Ore. It is also important to mention that in April 2012, capping was set at 25 g/t and the cut-off used was 2.25 g/t Au. If we use the current data applying a capping grade of 25 g/t Au and estimate the resource with a cut-off of 2.25 g/t Au, we will obtain 153,900 tonnes at 3.9 g/t Au of Indicated ore (19,120 oz). Compared to the 227,160 tonnes at 2.87 g/t Au (20,949 oz) previously reported, we can say that the current estimation is more selective. This reflects the fact that the current estimation was designed for a more selective mining method.

www.innovexplo.com


15. MINERAL RESERVE ESTIMATES

At the current stage of the project, there is no relevant information to be included in this item

.

www.innovexplo.com


16. MINING METHODS

16.1. Caution to the reader

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.

16.2. Introduction

Integra Gold’s Lamaque Project is separate from the adjacent Sigma and Lamaque mines, which have collectively produced over 9 Moz gold to a vertical depth of 2,000 m. Mineralization at the Lamaque Project would be accessed via two separate ramps, or declines, located in the Parallel Zone to the north (the “North Ramp”) and in the Triangle Zone to the south (the “South Ramp”), approximately 2 km apart. Material would then be transported to an off-site mill for toll processing, thereby eliminating the need for the construction and permitting of a new mill and tailings facilities. The mining plan for the Lamaque Project calls for a combination of conventional and mechanized mining. Two mining methods are proposed based on the vein geometry of the four deposits: long-hole and room and pillar. The approach in this study has been to force the application of long-hole mining where applicable. Waste material generated from drift development will be used to backfill part of the long-hole open stopes. An administration and mine service hub would be located on Highway 117, part of the Trans-Canada highway system. The service hub would be served by a 25 KV power line, natural gas and municipal services. There will be two production centers, each with a ramp to access resources (the “North Ramp” and “South Ramp”), and both will include basic surface infrastructure. The overall life of the Project is expected to be approximately 6.25 years, including a 2-year preproduction period, followed by 4.25 years of production. Once mining operations are completed, 1 year will be required to complete the mine closure work. At this early stage of the Project, several options can be considered for the processing plant. However, these assumptions would have to be re-examined and optimized during a pre-feasibility study. One of the main underlying assumptions of this PEA is that the mineralized material would be transported and processed off-site.

16.3. Mineral resources considered in the mining plan

InnovExplo designed the conceptual underground preliminary mine plan based on the Measured, Indicated and Inferred resources presented in an earlier report entitled “NI 43-101 Technical Report on the Lamaque Property”, published in November 2013 and prepared by Geologica Groupe-Conseil Inc. and GéoPointCom

www.innovexplo.com


Inc. (see section 14 for description). Details of the available resources for generating a preliminary mine plan are presented in Table 16.1.

Table 16.1 – Resources available to produce a preliminary mine plan with a 3.00 g/t Au cut-off

Gold Deposit Name Tonnes (metric tons)

Grade (grams per tonne)

Ounces

No. 4 Plug 1,325,100 5.6 237,450 Fortune Zone 377,800 5.7 68,820 Parallel Zone 947,300 9.7 295,620 Triangle Zone 932,000 11.0 328,270 Total 3,582,200 8.1 930,160

A portion of the tonnage included in the Parallel Zone comes from the No. 7 veins and consists of Inferred resources only. Two of the three No. 7 veins demonstrate much higher grades than the average grade of the Project. For this PEA, in order to minimize the influence of these satellite zones that have been estimated based on only one (1) diamond drill hole per vein, InnovExplo used only half the grade value for veins C and F. The long-hole tonnage was also reduced by half for all the No. 7 veins. See Table 16.2 for details.

Table 16.2 – Resources considered for the No. 7 veins to produce a preliminary mine plan

No. 7 Resources potentially available Resources considered for the mine plan

Tonnage Grade Ounces Tonnage Grade Ounces

Veins (t) (g/t) (oz) (t) (g/t) (oz)

C 41,178 17.94 23,751 24,369 8.97 7,028

D 43,476 4.69 6,556 26,778 4.69 4,038

F 40,274 27.72 35,893 23,917 13.86 10,658

TOTAL 124,928 16.48 66,199 75,064 9.00 21,723

For the Parallel, Fortune and Triangle zones, a cut-off grade of 4.0 g/t was applied to sectors amenable to the long-hole mining method and a cut-off grade of 4.5 g/t to sectors amenable to the room and pillar mining method. Small satellite veins were eliminated when located too far from principal points of access. A surface pillar of approximately 25 m from surface bedrock was considered. For No. 4 Plug, François Chabot of Integra Gold delineated stopes based on resource blocks calculated by GéoPointCom, in higher grade areas and provided the information to InnovExplo. Table 16.3 presents the resources potentially amenable to mining that were used to prepare the preliminary mining plan for the Lamaque Project.

www.innovexplo.com


Table 16.3 – Resources used to prepare a preliminary mine plan Deposit Name Tonnage

(metric tons) Grade (g/t)

Contained Gold (oz)

No. 4 Plug 401,400 6.4 82,213 Fortune Zone 194,484 6.9 43,012 Parallel Zone 723,557 9.3 215,540 Triangle Zone 694,360 13.0 289,849 Total 2,013,801 9.7 630,614 16.4. Preliminary geotechnical assessment

The lack of data on discontinuity sets and conditions for the Project makes it difficult to perform a proper rock mass classification. Basic assessments of the rock mass have been done using available data defined at the Lamaque-Sigma Complex, which is considered most representative with respect to the Lamaque Project’s mineralization and mining method. Assumptions about long-hole stope dimensions are based on standard designs in similar conditions. Integra Gold provided Golder Associates Ltd (“Golder”) with geotechnical data in the form of drilling records for 452 exploration drill holes. In early 2013, Golder provided training in order to improve the geotechnical data collected from exploration drill holes. As the Golder training was only conducted in early 2013, detailed geotechnical data was not collected for all the holes drilled. Rock Quality Designation (RQD) and fracture count data was collected from 100 holes located across the Property. Geotechnical data was only collected on eight (8) holes from the Triangle Zone that were logged following the training. The information provided below was extracted from Golder’s report.

16.4.1. Rock quality designation (RQD)

Rock quality designation (RQD), taken as the cumulative length of core longer than 10 cm divided by the length of the run and expressed as a percentage, was the most collected geotechnical information in the exploration drill holes. RQD statistics are summarized in Table 16.4 by rock type overall (see Figure 16.1). The RQD for both of the main rock types is quite high, with intrusives having a slightly higher RQD on average than the volcanics. Lower observed values are expected to correspond to fault zones in the rock mass.

Table 16.4 – Statistics for RQD data

www.innovexplo.com


Figure 16.1 – Histograms showing distribution and cumulative percentage RQD for a) Intrusive Rock, and b) Volcanic Rock

16.4.2. Total core recovery (TCR)

Total core recovery (TCR) is the measured length of core recovered divided by the length of the run, expressed as a percentage. TCR statistics are summarized in Table 16.5 by main rock unit and overall units. Drilling recovery was excellent, with an overall average of 99.2% and a standard deviation of 4.3%. There is no observable difference in TCR values between the main rock units. Note that all TCR data was collected from the Triangle Zone.

Table 16.5 – Statistics for TCR data

16.4.3. ISRM field hardness

ISRM field hardness is a scale that indicates the hardness of the core based on field tests. This field hardness then provides a coarse estimate of the UCS of the rock. Field hardness was only measured on eight (8) holes, all of which were located in the Triangle Zone. The statistics by main rock units are shown in Tables 16.6 and 16.7. These values can be used as an estimate for the rest of the rock mass at Lamaque.

www.innovexplo.com


Table 16.6 – Statistics for ISRM field hardness

Table 16.7 – Approximate UCS ranges by ISRM field hardness (modified from ISRM, 1978)

Based on the field hardness values measured from the core from the Triangle Zone, rock mass UCS is similar for the volcanic and intrusive rock types and can be estimated at approximately 75 MPa, with an overall range between 25 MPa and 250 MPa. While weaker sections of rock can be expected to be encountered locally (e.g., associated with faulting or shearing), an overall lower bound for the rock mass, corresponding to the lower limit of the R4 classification of 50 MPa, can be assumed.

16.4.4. Fracture frequency and fracture spacing

Fracture Frequency is a count of the number of fractures per metre of a length of core. The reciprocal of the fracture frequency provides a coarse estimation of the joint spacing within the rock mass. Fracture Frequency was measured in 99 boreholes across the site. The statistics are summarized in Table 16.8.

Table 16.8 – Statistics for fractures per metre of core

The fracture count indicates that an average of 4 fractures per metre of core can be expected for all rock types. However, the upper bound indicates that there are up to

www.innovexplo.com


7 fractures per metre in the volcanics, and 6 fractures per metre in the intrusives. This leads to an average fracture spacing of 0.25m in all rock types, with a lower bound fracture spacing of 0.14m in the volcanics, and 0.17m in the intrusives.

16.4.5. Crown pillar

Once the historical stopes have been located and confirmed, an investigation program should be undertaken, as necessary, to determine the presence, thickness and condition of crown pillars located above the historical mining. Of particular concern is the location of the stopes relative to the old mine tailings and existing surface infrastructure. The stability of these crowns, the associated risk of failure, and possible remediation measures should be addressed. The impact of dewatering the underground workings on the stability of the existing crown pillars, particularly above the existing Mine 3 workings, will also need to be further investigated. The historical workings in this area lie directly below the Lamaque tailings basin.

16.4.6. Typical ground support patterns

These preliminary ground support recommendations are based on standard industry practices. More detailed recommendations will require additional information regarding joint spacing and continuity. Based on Farmer and Shelton (1983) the following bolt lengths for the back (Table 16.9) are proposed based on the excavation span (Bolt Length = 0.3 Span).

Table 16.9 – Typical ground support bolt length. Bolt length* Maximum Span 5 ft (1.5 m) 16.5 ft (5 m) 7 ft (2.1 m) 23.0 ft (7 m) 8 ft (2.4 m) 26.2 ft (8 m) 10 ft (3.1 m) 32.8 ft (10 m) 12 ft (3.7 m) 39.4 ft (12 m) * Bolt length indicates the length installed within the rock and excludes any threads or bar outside the drill hole.

The standard support is illustrated in Figure 16.2 and consists of:

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

Wall: rows (number of rows based on excavation height) of rock bolts (length

= 1.2 – 1.5 m) on a 1.2 m x 1.2 m (4’ x 4’) pattern

www.innovexplo.com


Figure 16.2 – Standard requirement for ramp ground support

Screening of the back to 1.5-2 m above the floor is recommended for all excavations 3.5 m or higher (Figure 16.3). 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 are available it will be possible to optimize the ground support standards.

www.innovexplo.com


1,5

2

2,5

3

3,5

4

4,5

5

5,5

6

6,5

1,5 2 2,5 3 3,5 4 4,5 5 5,5 6 6,5 7

Excavation Height (m

)

Excavation Span (m)

Standard Support Pattern as a function of Excavation Dimensions

Add 1 bolt to back / additional

category

Add 1 row of wall bolts / additional

category

Add screen (back to 1.5 m above BOR)

Figure 16.3 – Standard support pattern as a function of excavation dimensions. Note that the transitions between bolting patterns are suggested guidelines and will vary slightly depending on the degree of arching of the excavation back.

16.4.7. Summary of geotechnical data

In order to be able to provide a complete classification of the rock mass, more data is required for the Lamaque Project site. In particular, information is needed with respect to the number of joint sets present in the rock masses, the condition of the discontinuities in the rock mass, the state of stress anticipated, and the water conditions in the areas under investigation. Understanding the orientation of the joint sets for each rock type can assist with mine design and ground support consideration. Understanding the intact rock strength at both zones is also required since current information is based on the Triangle Zone. The RQD and TCR results suggest a solid rock mass. The fracture frequency data suggests a blocky rock mass with an average joint spacing of 0.25 m. Based on experience, available descriptions in the literature, and available information for the Project, it is expected that the rock mass will classify as Fair to Good according to both the NGI Q and Bieniawski RMR76 classification systems. It is anticipated that some areas of poorer ground exist in each zone and are

www.innovexplo.com


associated with large-scale faults and shears. These assumptions will have to be verified during subsequent investigations and levels of study.

16.5. Mining method

As mentioned earlier, two mining methods are proposed to accommodate the geometry of the mineralization. The mining methods were selected according to vein geometry and common practices for comparable mining operations in the region, an area with an extensive history of underground mining. For mineralized zones dipping less than 45°, a room and pillar mining method is proposed, and sublevel long-hole retreat is proposed for zones dipping more than 45°.

16.5.1. Long-hole method

For mineralized zones dipping more than 45° in the North Ramp and Triangle zones, the long-hole mining method will be used with mechanized sublevel development completed at 16 m intervals along the vein. Typical stopes will have a thickness of 2 m and a length of 20 m as presented in Figure 16.4. The development sequence consists of 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 (Figure 16-5). The method consists of drilling and blasting 64-mm (2.5”) diameter holes in a pattern parallel to the walls. Mucking will be done longitudinally using remote controlled scoops. Rock-fill will be used in long-hole stopes for the North Ramp. Assuming un-cemented backfill and a 3m pillar between adjacent stopes, a mining recovery of 85% is considered.

www.innovexplo.com


Figure 16.4 – Typical isometric stope view of North Ramp and Triangle zones.

www.innovexplo.com


Figure 16.5 – Typical longitudinal stope view of North Ramp and Triangle zones

The mining method planned in the No. 4 Plug is only long-hole. However, since the stopes are larger than other areas of the mines, drilling will be achieved on a 76-mm (3”) diameter. Figure 16.6 presents a typical drilling pattern in the No. 4 Plug. In this study, 11 stopes are present in the No. 4 Plug, distributed over 400 m in elevation, and no backfill is planned to be used.

Figure 16.6 – Typical drilling patterns in the No. 4 Plug

www.innovexplo.com


16.5.2. Room and pillar

The proposed room and pillar stope configuration is based on typical industry practices for currently operating mines in deposits with similar vein geometry. The typical mining height will vary from 2.0 m to 3.0 m. The room and pillar mining method entails the excavation of a series of “rooms” following the vein, leaving “pillars” or columns of rock in place to help support the mine roof. In conventional room and pillar mining, drilling is achieved using hand-held drill equipment and holes are loaded with explosives. Bolts are then installed in the mine roof to ensure the roof is properly supported. The broken rock is scraped to either a raise or a drawpoint using electric slusher with scraper. From there, the broken material is taken with a scooptram to be hauled to surface with truck. A typical room and pillar design is proposed for the North and South ramps using 6 m wide rooms with 3 m x 6 m pillars and an access of 4 m between pillars. Figure 16.7 presents a schematic of a typical room and pillar design. Mechanized sublevel development from which the broken material is loaded by scooptram will be completed at 60 m intervals along the vein. A mining recovery of 85% is considered.

Figure 16.7 – Typical room and pillar design

16.6. Existing mine infrastructure

Existing mine infrastructure that might impact the mining proposed mining plan elaborated in this PEA is Mine No. 3, which was developed in 1961. The Mine No. 3 was developed from a three-compartment vertical shaft excavated to a depth of 146.3 m (480 ft) with three levels. It is located below the present-day tailings. Historical records indicate that 152,015 tonnes were mined from the Mine No. 3.

www.innovexplo.com


Drifts were extended toward the No. 4 Plug from the No. 3 Mine area to provide access at the 450-ft (137.16 m) and 700-ft (213.36 m) levels. The No. 4 Plug was mined and yielded 160,000 tons from workings above the 700-ft level. On the 300-ft level (91.44 m), the No. 3 Mine is also connected with the upper part of the No. 5 Plug. In the proposed mining plan, the 700-ft (213.36 m) level is located at a distance of 15 m from the safety raise planned for the Parallel Zone at the elevation 100-2. The 300-ft (91.46 m) level is located in-between the plan ramp and the safety access drift connecting Parallel-Fortune zones at a distance of 10 m from the ramp and 7 m from the drift. In the South Ramp, the 700-ft (213.36 m) level runs 30 m over stope 24 of the No. 4 Plug. These openings will not be rehabilitated, although the mine will be dewatered before these areas are mined (see Figure 16.8).

Figure 16.8 – Plan view of Lamaque Project existing infrastructure

www.innovexplo.com


16.7. Dewatering

In parallel with the development work, the existing Mine No. 3 mine infrastructure will have to be dewatered from surface to level 700-ft (213.36 m). The volume of the existing opening is estimated to be 100,000 m3.

16.8. Underground mine design

16.8.1. North Ramp development

The North Ramp will be located approximately 1 km from the service hub and will reach a vertical depth of 615 m to initially access the Parallel and Fortune zones. If Integra Gold outlines resources at its No. 5 Plug and No. 3 mine targets, it is anticipated these zones would also be accessed through the North Ramp. The compilation of mine development quantities considered in the preliminary mine plan is presented in Table 16.10.

Table 16.10 – Mine development quantities for the North Ramp Total

‐1 0 1 2 3 4

North Ramp

Ramp (4.5m x 4.5m or 4.5 m x 5m)) 1 152 2 629 2 834 2 695 1 303 ‐ 10 613

Drift (4m x 3m) 0

Drift (4m x 4m) 0

Sub level (3m x 3m, 3.5m x 3m or 3.5m x2.8m) 28 2 721 2 325 1 768 2 261 ‐ 9103

Alimak raise (4.3m x 4.3m) 172 172

Alimak raise (3m x 3m) 236 194 430

Conventional raise (4.3m x4.3m) 37 37

Conventional raise (2.5m x 2.5m or 3m x 3m) 30 75 54 159

Development Preproduction (year) Production (year)

The portal ramp will start from an outcrop and the ramp size will be 4.5 m x 4.5 m in order to accommodate the required equipment. In the upper part of the Parallel Zone, six (6) levels will be developed. In the lower part of the Parallel Zone (No. 7 veins), sublevels will be developed to access 3 mineralized materials zone. The excavation in mineralized materials zone is usually 3 m high, but the width varies in relation of the thickness of the veins in order to provide the best development scenario for drilling positions. Ore and waste will be hauled by LHDs from the production area to either a remuck bay or to a loading point close to the ramp and loaded into trucks to be hauled directly to surface. Figure 16.9 illustrates the general arrangement of development in the Parallel and Fortune zones.

www.innovexplo.com


Figure 16.9 – Longitudinal view of Parallel Zone (incl. No. 7 veins) and Fortune Zone, looking north.

Parallel is connected to Fortune by a 4.5 m x 4.5 m ramp for 45t trucking haulage and by a 3 m x 3 m track drift access for ventilation and a safety exit (Figure 16.10.)

www.innovexplo.com


Figure 16.10 – Plan view of the Parallel-Fortune connection.

www.innovexplo.com


16.8.2. North Ramp sequence

According to the mine design, development will start in the second quarter of year -1 with a jumbo development crew consisting of 3 men working 10-hour shifts day and night (see Table 16.11). At the end of Year 0, another jumbo crew will be added. In the middle of Year 0, a third development crew is scheduled to start developing accesses and drifts for the mineralized material using modified long tom rigs on wheels. Adding a third team will accelerate the start of production in the Parallel Zone once the safety exit from the 200-2 Level to surface is completed (Figure 16.11). Production using the room and pillar method will commence with two crews of 2 men per shift in the second quarter of Year 0. The number of team will gradually increase up to the maximum of 12 teams at Year 1. Long-hole production will start in the beginning of Year 1 and will finish at the end of Year 4. Production drilling will be carried out by 6 drillers on 2 drill rigs.

Table 16.11 – Operational work place for the North Ramp

‐1 0 1 2 3 4 5

North Ramp

Jumbo development (3 miners/crew, 3 crews/gear) 1 2 2 2 2

Coventional development (drift, 2 miners/crew, 3 crews/gear) 1 1

Room & pillar (2 miners/crew, 2 crews/stope) 3 6 6 6 4

Operational work place

Excavation type Preproduction (year) Operational period (year)

www.innovexplo.com


Figure 16.11 – Longitudinal view of the Parallel Zone without the No. 7 veins, looking north.

www.innovexplo.com


In the second quarter of Year 1, development will be completed on the Parallel Zone. The next priority will be to develop access to the No. 7 veins and the Fortune Zone. One jumbo crew will be assigned to developing the ramp to access the No. 7 veins, and a second team will focus on the Fortune Zone. The long tom equipment will then be modified to operate on rails, in order to develop the safety access between the Parallel and Fortune zones. The development of the No. 7 veins and the Fortune Zone will be completed at the end of Year 3 and production will be finished at the end of Year 4.

16.8.3. South Ramp development

The South Ramp will be located approximately 3 km from the service hub and will reach a vertical depth of 620 m. The South Ramp will access the Triangle Zone and No. 4 Plug. The South Ramp area would be connected to an existing gravel road a few hundred metres to the south, allowing for two entry points to the site. The portal ramp will start from an outcrop and the ramp size will be 5 m x 4.5 m. In the upper part of the Triangle Zone, 6 levels will be developed with sublevels. In the No. 4 Plug, a crosscuts will be developed from a main ramp to access the mineralized zones. The mineralize material and waste will be hauled by LHDs from the production area to either a remuck bay or to a loading point close to the ramp and loaded into 45t trucks to be hauled directly to surface. Figure 16.12 shows the longitudinal view, and Table 16.12 presents a compilation of mine development quantities.

Table 16.12 – Mine development quantities for the South Ramp Total

‐1 0 1 2 3 4

South Ramp

Ramp (4.5m x 4.5m or 4.5 m x 5m)) 803 2 376 2 257 2 338 1 405 207 9386

Drift (4m x 3m) 75 492 2 580 419 3566

Drift (4m x 4 m) 99 440 153 219 911

Sub level (3m x 3m, 3.5m x 3m or 3.5m x2.8m) 1 412 1 738 1 158 4308

Alimak raise (4.3m x 4.3m) 0

Alimak raise (3m x 3m) 233 408 205 188 1033

Conventional raise (4.3m x4.3m) 0

Conventional raise (2.5m x 2.5m or 3m x 3m) 15 42 57

Development Preproduction (year) Production (year)

www.innovexplo.com


Figure 16.12 – Longitudinal view of the Triangle Zone and No. 4 Plug, looking west. 16.8.4. South Ramp sequence

According to the mine design, development will start in the third quarter of the -1 year with jumbo development crews consisting of 3 men working 10 hour shifts day and night. At the end of Year -1, another jumbo crew will be added. In the middle of Year 0, a third development crew is scheduled start developing access and drifts in mineralized zones using a modified long tom on wheels (see Table 16.13). The Triangle Zone will be mined by room and pillar method only. The first 2 crews will start in the last quarter of Year 0 and the number will increase to a maximum of 11 room and pillar crews in Year 1. Long-hole production in the Triangle Zone will also start in the last quarter of Year 0 and will finish in the first quarter of Year 5 in the No. 4 Plug. Development of the No. 4 Plug will start in the first quarter of Year 1 and will finish at the beginning of Year 4. Long-hole production of the No. 4 Plug will start at the end of Year 3 to finish at the beginning of Year 5.

www.innovexplo.com


Table 16.13 – Operational work place for the South Ramp

‐1 0 1 2 3 4 5

South Ramp

Jumbo development (3 miners/crew, 3 crews/gear) 1 2 2 2 2 2

Coventional development (drift, 2 miners/crew, 3 crews/gear) 1 1

Room & pillar (2 miners/crew, 2 crews/stope) 4 5 6 5

Excavation type

Operational work place

Preproduction (year) Operational period (year)

16.9. Mining dilution and recoveries

A mine recovery factor of 85% has been considered in this study for room and pillar stopes because high-grade portions of the pillars will be recovered. Typical room and pillar stopes will have a minimum height of 2m and an external dilution of 5% (at 0.0 g/t Au). For long-hole stopes, an average dilution factor of 20% (at 0.0 g/t Au) has been applied when stopes thickness is greater than 3 m, and a 35% dilution factor was applied to stopes less than 3 m. However, in the No. 4 Plug, the stopes are thicker and more vertical; in this sector, a 15% dilution factor (at 0.0 g/t Au) was applied with a planned mining recovery factor of 90%.

16.10. Mining rate

The expected production rate will start at an average of 690 tpd from the North sector in the third quarter of preproduction Year 0, and will slowly ramp up to an average of 750 tpd in Year 1. The South sector production will start in last quarter of preproduction Year 0, to bring the production level to 790 tpd. The preliminary conceptual mine plan extends over a period of 4.25 years, including a 2-year preproduction period. The average production level for North and South Ramp is 1,480 tpd (312 days/year). Limited production would occur during the preproduction stage, accounting for approximately 28,000 ounces over the course of the two-year preproduction period. Average annual production after the preproduction stage is 460,500 tonnes at a diluted grade (or head grade) of 8.24 g/t Au, for 112,400 Au ounces recovered (average recovery of 92.1%).

16.11. Mine plan schedule criteria

Contractors will be used for Alimak raise excavation, portal entrance and specific surface work. Mine workers will be used for all other type of mine development, mine production and mineralize material haulage activities. A staff workforce will be hired to provide technical and administrative support and direction to mine activities. The design criteria used to develop the mine plan are as follows:

Jumbo multi-faces (>3) drift development: 200 m/month Jumbo with maximum two-face drift development: 175 m/month

www.innovexplo.com


Jumbo with single-face drift development only: 150 m/month For the jumbo development, a progressive efficiency factor of 40%-60%-80%

is assumed for the first three months to account for the learning curve; Track drift development: 100 m/month Sublevel development: 60-100 m/month; Alimak raises: 75 m/month; Conventional raises: 60 m/month; Room and pillar stope: 18 t/man per shift, 8-10 hours per shift (day-night); Average mining production rate of 1,480 tpd, after preproduction stage.

The manpower resource on each working shift used to prepare the following mine schedule includes:

2 men per crew, usually with cross-shift for room and pillar; a maximum of 23 crews will be required for the Project;

3 long-hole drillers; 4 truck drivers; 3 LHD operators; 12 development crew per jumbo, for a total of 4 jumbos: Each crew consists of 1 jumbo operator and 2 workers for ground support

and services; the truck operator is not included; Three crews of 2 workers for sublevel development in the North Ramp; Two crews of 2 workers for sublevel development in the South Ramp.

16.12. Equipment

This study is based on new equipment that would be acquired by Integra Gold through financing agreements. The list of equipment considered for the Project is presented in Table 16.14.

Table 16.14 – Mining equipment for the Lamaque Project

‐1 0 1 2 3 4 5

Truck (AD45) 1 2 3 4 4 4 2

Scoop R1300G (4.4 cyd) 2 5 5 4 4 3 2

Scoop ST1030 (6 cyd) 4 4 4 4 3 2 1

Scissor lift SL3 2 2 2 2 2 2 1

Scissor lift SL2.5 1 2 2 2 1 0 0

Jumbo boomer 282 4 4 4 4 3 0 0

Bolter Maclean MEM 928 2 2 2 2 2 0 0

Grader CAT 12M 0 1 1 1 1 1 1

Long Tom (rails & sub) 0 2 2 0 0 0 0

Long hole drill (C‐MAC PLH) 0 3 3 3 3 3 0

Tractor ‐ KUBOTA 4 6 10 10 10 8 5

Lube and service truck ‐ Maclean 0 1 1 1 1 1 1

Surface Loader 2 2 2 2 2 2 1

Water TANK (TRAILER) 1 1 1 1 1 1 1

Pickup Truck 2 4 4 4 4 4 2

School bus 2 2 2 2 2 2 1

Total 27 43 48 46 43 33 18

Mining equipmentPreproduction (year) Operational period (year)

www.innovexplo.com


16.13. Manpower requirements

The mine will operate seven (7) days a week, excluding night shifts on Fridays and Saturdays. This schedule is equivalent to 312 days per year of operation.

Technical services and administration will be on a schedule of 5 days work - 2 days rest.

Development and production crews will be on a schedule of 7 days work – 4 days rest – 5 days work (night shift) – 5 days rest, for ten (10) hours per shift.

Room & pillar crews will be on a schedule of 5 days work (8 hours/shift) - 2 days rest – 4 days work (10 hours night shift) - 3 days rest.

Table 16.15 and 16.16 presents the manpower requirements during the LOM, excluding process plant manpower.

Table 16.15 – Manpower requirements – Administration and Surface Services

‐1 0 1 2 3 4 5

Administration

Manager 1 1 1 1 1 1 1

Secretary 1 1 1 1 1 1 1

Senior Accountant 1 1 1 1 1 1 1

Int Accountant 1 1 1 1 1 1 1

jr. Accountant 1 2 2 2 2 2 2

Human ressources Coordinator 1 1 1 1 1 1 1

Communication Coordinator 1 1 1 1 1 1 1

Senior Purchaser 1 1 1 1 1 1 1

Purchaser 1 3 3 3 3 3 3

Clerk 1 2 2 2 2 2 2

Safety coordinator 1 1 1 1 1 1 1

Safety preventionnist 1 1 1 1 1 1 1

Nurse 1 1 1 1 1 1 1

Training coordinator 1 1 1 1 1 1 1

Surface services

Surface Labourer 4 4 4 4 4 4 2

Construction Superintendent 0,5 0 0 0 0 0 0

Civil technicien supervision 0,5 0 0 0 0 0 0

Dryman 1 2 2 2 2 2 2

ManpowerPreproduction (year) Production (year)

www.innovexplo.com


Table 16.16 – Manpower requirements – Technical Services, Maintenance, Supervision and Operations

‐1 0 1 2 3 4 5

Technical Services

Geology

Chief Geologist 1 1 1 1 1 1 1

Database technician 1 1 1 1 1 1 1

Senior geology technician 1 1 1 1 1 1 1

Senior geologist 2 2 2 2 2 2 1

Junior geologist 2 4 4 4 4 4 2

Junior geology technician 1 3 4 4 4 4 2

Engineering

Chief Engineer 1 1 1 1 1 1 1

Rock mechanic engineer 1 1 1 1 1 1 1

Mining technician project 0 1 1 1 1 1 1

Senior mining engineer 2 2 2 2 2 2 1

Junior mining engineer 2 3 3 3 3 3 2

Senior mining technician 2 2 2 2 2 2 1

Junior mining technician 2 3 4 4 4 4 2

Environmental department

Environmental coordinator 1 1 1 1 1 1 1

Environmental surface supervisor 1 1 1 1 1 1 1

Environment technician 1 1 1 1 1 1 1

Maintenance

Mechanical department

Chief mechanic 1 1 1 1 1 1 1

Preventive maintenance planner 1 1 1 1 1 1 1

Mechanic supervisor 2 2 2 2 2 2 1

Mechanics 6 10 12 12 12 12 6

Welder 1 1 2 2 2 2 1

Electrical department

Chief electrician 1 1 1 1 1 1 1

TI Technician 1 1 1 1 1 1 1

Electrician supervisor 1 1 1 1 1 1 1

Electrician 6 8 8 8 8 8 4

Supervision Operation

Underground superintendant 1 1 1 1 1 1 1

Mine Captain 2 2 2 2 2 2 1

Shift Boss 6 12 14 14 14 14 7

Underground supervisor service 0 0 2 2 2 2 1

Operation

Labourer 1 1 1 1 1 1 1

Service miner 6 10 12 12 12 12 6

Construction miner 0 4 4 4 4 4 2

Development lead miner 6 17 17 12 12 6 0

Development miner 12 29 29 24 24 12 0

Development muker 0 3 9 9 11 12 6

CP miner 0 26 44 48 44 16 0

Long hole driller 0 6 9 9 9 9 6

Blaster 0 2 4 4 4 6 4

Scoop operator 0 6 12 12 12 12 6

Truck operator 4 6 12 12 12 12 6

Trainer 2 6 6 6 6 6 3

Total Hourly & Staff 101 209 261 255 253 210 109

Staff 52 68 74 74 74 74 54

ManpowerPreproduction (year) Production (year)

www.innovexplo.com


16.14. Development and production schedule

InnovExplo developed a preliminary development and production schedule based on the mineral resources discussed in section 14. The underground mine design provides for a 4.25-year mine plan producing 2,081,410 tonnes of mineralized material assaying 8.19 g/t Au. With a resulting average mill recovery of 92.2%, a total of 505,611 oz of gold will be produced during the LOM, with 28,000 oz of gold from the preproduction period. During the LOM, development would generate approximately 14% of the mineralization tonnage, room and pillar mining 36%, and long-hole mining 50% (Table 16.17).

Table 16.17 – Mine plan tonnage distribution Mining North Ramp

(tonnes) South Ramp (tonnes)

Development 159,900 126,400 Room and pillar 290,000 468,200 Long-hole 539,400 497,500 Total 989,300 1,092,100

The mining plan includes all development required to access and mine the mineralized zones. Estimated development quantities are presented in Table 16.10 and 16.12 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 presented the type of mineralized material by sector and Table 16.18 summarizes the yearly tonnage distribution.

www.innovexplo.com


Table 16.18 – Conceptual mining plan, yearly tonnage distribution for the North and South ramps

Preproduction

Total

PRODUCTION

North Zone

PARALLELE

Development (t) 59,573 32,987 6,275 18,925 - 117,760

Grade (g/t) 6.70 5.99 13.68 7.08 - - 6.94

Long Hole (t) - 106,476 164,098 38,287 73,079 - 381,940

Grade (g/t) - 7.59 6.56 5.31 9.09 - 7.21

Room and pillar (t) 19,257 78,044 79,689 62,457 26,338 - 265,785

Grade (g/t) 6.25 7.63 9.54 9.71 9.75 - 8.80

Total (tonne milled) 78,830 217,507 250,061 119,669 99,417 765,485

Grade (g/t) 6.59 7.36 7.69 7.89 9.26 7.72

Inventory

Grade (g/t)

FORTUNE

Development (t) 16,254 25,880 0 42,134

Grade (g/t) 5.18 4.98 0.00 5.06

Long Hole (t) 15,600 76,771 65,071 157,442

Grade (g/t) 5.20 5.11 5.14 5.13

Room and pillar (t) 4,095 14,268 5,848 24,211

Grade (g/t) 5.71 6.20 7.15 6.35

Total (tonne milled) 35,949 116,919 70,919 223,787

Grade (g/t) 5.25 5.21 5.31 5.25

South Zone

TRIANGLE

Development (t) 28,431 34,711 25,353 - - 88,495

Grade (g/t) 7.95 8.05 10.17 - - 8.63

Long Hole (t) 4,000 28,197 26,785 53,779 - 112,761

Grade (g/t) 6.49 6.44 7.36 10.57 - 8.63

Room and pillar (t) 13,230 118,105 170,687 134,840 31,323 468,185

Grade (g/t) 11.83 9.44 12.96 14.53 15.57 12.67


Grade (g/t) 8.95 8.71 11.97 13.40 15.57 11.45

PLUG 4

Development (t) 3,529 25,101 9,287 37,916

Grade (g/t) 3.59 5.06 3.72 4.59

Long Hole (t) 13,415 290,550 80,815 384,781

Grade (g/t) 6.16 5.40 5.88 5.53

Room and pillar (t)

Grade (g/t)

Total (tonne milled) 3,529 38,516 299,837 80,815 422,697

Grade (g/t) 3.59 5.44 5.35 5.88 5.44

TOTAL North and South

Total (tonne milled) 124,491 398,520 512,364 463,724 501,496 80,815 2,081,410

Grade (g/t) 7.46 7.97 9.35 9.25 6.76 5.88 8.19

Parallele Recovery (%) 97% 97% 97% 97% 97% 97% 97%

Fortune Recovery (%) 95% 95% 95% 95% 95% 95% 95%

Triangle Recovery (%) 90% 90% 90% 90% 90% 90% 90%

Plug 4 Recovery (%) 86% 86% 86% 86% 86% 86% 86%

Gold Produced (oz) 28,001 95,493 143,257 127,046 98,666 13,148 505,611

Operation

Year 0 Year 1 Year 2 Year 3 Year 4 Year 5

www.innovexplo.com


16.15. Mining services

16.15.1. Ventilation for North and South ramps

The underground ventilation installation for the North and South ramps will be similar. The ventilation raise will serve as an emergency escape way in both cases (see Figures 16.13 and 16.14). For the current study, InnovExplo performed a preliminary simulation to establish a main ventilation network from which secondary fans will bring air to the working area. The maximum required ventilation was determined to be 240,000 cubic ft per minute (6,800 cmm) and will be provided by two 300 hp main air fans (see Table 16.19). Fresh air will be heated by two 12-MBtu/hr capacity propane burner systems and will exhaust via the ramp. The heating installation will be at the surface, but the main fans will be installed underground in a sublevel close to the surface. The ventilation OPEX cost calculation was done using 67% of full operating fan capacity or 160,000 cfm. InnovExplo believes it will be possible to optimize ventilation requirements. The system will be flexible enough to permit adequate ventilation even if some modifications are made to the equipment fleet.

Table 16.19 – Airflow requirements YEAR 0 1 2 3 4 5

MAXIMUM DESIGN 180,000 210,000 240,000 240,000 240,000 200,000

OPERATING 90,000 135,000 160,000 160,000 160,000 130,000

OVER ALL SYSTEM 240,000

AIRFLOW

REQUIREME

NTS (CFM)

Figure 16.13 – Ventilation network for the North Ramp

www.innovexplo.com


Figure 16.14 – Ventilation network for the South Ramp

16.15.2. Dewatering

16.15.2.1. Dewatering: North Ramp

The dewatering system consists of secondary sumps located on each level and 3 main pumping stations, two located in the Parallel Zone and one at the bottom of the Fortune Zone, as presented in Figure 16.15. Stations 2 and 3 pump to Station 1, where all collected water is pumped to surface at the dewatering water plant. Each sump will be connected by drain hole and a scoop tram will be used to keep the sumps clear of mud. The dewatering from the bottom of the Parallel Zone will be served by the development pumping system. The LOM of this sector (No. 7 veins), is too short to invest in permanent infrastructure. The development dewatering system uses a submersible 8.9 HP pump with metal tub mounted on a skid. These setups are installed in series every 20m in elevation, depending on the piping size.

www.innovexplo.com


Figure 16.15 – Longitudinal view of the dewatering system for the Parallel and Fortune zones, looking west.

16.15.2.2. Dewatering: South Ramp

In the south sector, there are 4 principal pumping stations: three located in the Triangle Zone and one approximately in the lower third of the No. 4 Plug (Figure 16.16). The bottom dewatering of the Triangle Zone and No. 4 Plug will use the development pumping system. The LOM of these sectors are too short to invest in permanent infrastructure.

www.innovexplo.com


Figure 16.16 – Longitudinal view of the Triangle Zone and No. 4 Plug dewatering system, looking west.

All dewatered water of the South Ramp will be pumped to pumping station 1 and then to the sump closest to surface. The water will be repumped to the North Ramp water treatment plant using two 15hp pumps via 6-inch pipe.

16.15.3. Compressed air

Four 41.8 m3/min (1,476 cfm) electric compressors (3 fixed speed and 1 variable speed) will be installed in a secondary garage on surface in the North and South areas. A preliminary network of pipe lines will be installed down each escapeway 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 provide emergency air supply to the refuge station.

16.15.4. Industrial water

Industrial water will be recycled from the water pond. During preproduction, dewatering of Mine 3 will provide the main source of industrial water. During production, industrial water will be provided via dewatering of Mine 3 and recycling treated water. A network complete with pressure reducing valves will supply water to the underground operations.

www.innovexplo.com


16.15.5. Underground power distribution

Two 25 kV/600 V substations will be installed at each ramp (North and South). The North and South substations will be used to supply all underground mobile substations. For the underground main fan installations, one substation of 2 MVA 4.16 KV/600 V will be installed at each ramp to supply the main fans. For the entire North Ramp project, four (4) 1 MVA 4.16 KV/600 V underground mobile substations will be installed at different levels to follow production requirements. These substations will be skid-mounted and will provide 600V power to the underground loads, such as pumps, fans, jumbos, diamond drill, lunchroom, etc. For the entire South Ramp project, six (6) substations 1 MVA 4.16 KV/600 will be installed at different levels to follow production requirements and provide the flexibility required by the mining operations.

www.innovexplo.com


17. RECOVERY METHODS

In view of potential mining activities, custom milling will be the preferred option. The metallurgical testwork achieved to-date, has demonstrated the amenability of the Lamaque mineralized material to the gravity, leaching and flotation processes, although further work is required to better determine the specific flowsheet that will optimize the metallurgical performance. A study carried out by InnovExplo (Poirier S. and Tremblay A.: Review of Custom Milling Options, Lamaque Property Project, February 18, 2013) reviewed the potential of a number of existing milling facilities in the Val d’Or – Rouyn-Noranda area for custom milling. The Lamaque Project is planned on a 5- to 10-year period at a production rate that could likely range from 500 to 1,500 tpd. Five (5) gold concentrators located within an 80 km radius were then identified as being able to potentially process the Lamaque material: Beacon Gold Mill, Kiena Mill, Sigma-Lamaque Complex, Camflo Mill and Doyon 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

Beacon Gold Mill Bankruptcy Leaching/Merrill‐Crowe 500 ‐ 750 tpd 18 km

Kiena Mill Wesdome Leaching/CIP 1,000 – 2 200 tpd 19 km

Sigma‐Lamaque Complex Bankruptcy Gravity Concentration &

Leaching/CIP 1 200 ‐ 2,400 tpd 3 km

Camflo Mill Richmond Mines Leaching/Merrill‐Crowe 800 ‐ 1,200 tpd 25 km

Doyon Mill IAMGOLD Gravity Concentration &

Leaching/CIP 1 300 tpd 80 km

The Kiena Mill would be an interesting option but has no current availability for custom milling. The Kiena mine has been shut down since July 2013. It is not known if the current owners would be interested in restarting the milling facility for custom milling. Also, the CIP process (Carbon-In-Pulp) could be problematic in a custom-milling situation if Integra Gold is not the unique mill feed. Management of loaded carbon inventory in the circuit could make gold recovery reconciliation complicated. The Sigma-Lamaque Complex could be an interesting choice with its mill capacity of 1,200 to 2,400 tpd and its proximity of less than 3 km. The CIP (Carbon-In-Pulp) gold recovery process could again be problematic in a custom-milling situation if Integra Gold is not the unique mill feed. It is not known if there is any interest for custom milling at the moment due to the bankruptcy situation. The Beacon Gold Mill was refurbished in the 1990s. The condition of the equipment and the interest for custom-milling are not known as the company is in bankruptcy.

www.innovexplo.com


Treatment capacity would need to be increased to reach the anticipated mining production. However, the proximity and gold recovery process of the mill make it a possible candidate to treat the Lamaque material. IAMGOLD has shown interest for custom milling at the Doyon concentrator. The plant already processes the ore from the Westwood and Mouska Mines and the current circuit could be modified to treat an additional 1,300 tpd of feed. The ore transportation cost would be high, taking into account its distance from the Lamaque Project. Here too, the CIP process (Carbon-In-Pulp) could be problematic in a custom-milling situation, with the Westwood ore being milled, unless circuits were independent. Based on the metallurgical results and the above considerations, the Camflo Mill appears to be a good compromise. A few years back, a bulk sample of the Croinor ore was processed there. It could be difficult in 2013-2105 to have the mill available for custom milling due to ore from the Monique Project being processed at this facility. However, the plant is likely to have availability in the near future for custom feed. Typical flowsheets for both Merrill-Crowe and CIP (Carbon in Pulp) leaching processes are depicted in Figures 17.1 and 17.2 respectively.

Figure 17.1 – Typical Merrill-Crowe flowsheet

www.innovexplo.com


Figure 17.2 – Typical CIP (Carbon-in-Pulp) flowsheet

www.innovexplo.com


18. PROJECT INFRASTRUCTURE

18.1. Plant and site layout

Figure 18.1 presents the general surface layout, including the proposed location of the required infrastructure for the Lamaque Project.

www.innovexplo.com


Figure 18.1 – Surface view of the Lamaque Project.

www.innovexplo.com


18.2. Lamaque Project 25 kV surface distribution

A 25 kV voltage level was selected to optimize voltage regulation and to use lower cost standard equipment. The North and South ramps will be linked to the Hydro-Québec grid via a private 25KV to be built. As previously stated, the underground main distribution feeders will be operating at 4.16 kV to ensure proper cable sizing and adequate voltage regulation. The site surface buildings will be supplied by one 25 KV/600 V line mounted on wooden poles.

18.3. Site access

The Lamaque Project, located within 3 km of Val-d’Or, a mining community of over 35,000 people, benefits from world-class infrastructure. In addition to the considerable physical infrastructure, there is a high level of underground mining expertise readily available in the Val-d’Or region. The administrative and hub site of the Lamaque Project is located along Trans-Canada Highway, with all services readily available at the site. Integra Gold’s current office is located between the highway and the Project, on a property owned by Integra Gold, and there is sufficient land to accommodate the proposed development needs of the Project, including the proposed service hub. The plan as outlined in the PEA will have minimal impact on the community as there are no homes, businesses or other infrastructure where the proposed mining will take place.

18.4. Camp

No permanent or temporary camp is included. Given the proximity to Val-d’Or, special transportation arrangements are not necessary for the workers.

18.5. Mine site entrance/guardhouse

All visitors, contractors, delivery personnel and mine personnel will go through a main entrance located on the north side of the site. A temporary guardhouse will be erected at this point and anyone entering the site will do so only after being authorized by security personnel. A fenced-in car park will be located next to the gate and will have electric outlets to plug in the vehicles during cold weather.

18.6. Office building and dry complex

A modular building will be installed to serve as offices for administration, engineering and geology. The dry installation will be adjacent to the offices. Phase 1 will cover the preproduction period, and Phase 2 will complete the installation for the production period. A supervision office will be installed between the dry and the office.

www.innovexplo.com


18.7. Service buildings

A main garage will be built in the main office area to be used for equipment maintenance (mechanic and electric) and a stock room will be included. Offices for the maintenance supervisor and buying department will be included as well as a large lunch room and meeting room. A secondary garage and a cold storage will be erected on a cemented pad at each ramp area (North and South).

18.8. Site Roads

A service road providing access to various parts of the Property will be built and services (electric line, water line) will be installed along the road.

18.9. Compressor building

All compressors will be installed on surface in secondary garages at the North and South surface areas.

18.10. Fuel storage

Diesel fuel for the mine equipment and vehicles will be stored in an above-ground tank. A 7,500 litre tank (2,000 US gal.) will be at the main garage, and a 28,000 litre tank (7,500 US gal.) will be at both secondary garages at the North and South installations. There is no gas station at the site; only surface pick-up trucks will be gas operated and they will be fuelled off-site.

18.11. Site fencing

Fencing will only be provided for the main substation, the propane tanks, and for a short distance on each side of the main entrance gate. The North and South ramp sites will be fenced with automatic gates.

18.12. Water systems

Fresh water for the administration and hub site will be supplied by Val-d’Or. The current core shack is already connected to the town’s services.

18.13. Communication system

The Lamaque site will be connected to the public telephone service using an IP telephone network. The communication between buildings will use monomode optic fiber. To allow employees to have wireless access, a network access point, WIFI Unifi AP Pro, will be installed in each of the buildings to permit cellular and computer connections. The surface radio system consists primarily of channels with local short-distance coverage or extended coverage. The following channels are planned for the site:

Security/emergency; Surface operations; General and maintenance (mechanical/electrical/housekeeping/etc.); Underground operations (underground link with surface).

www.innovexplo.com


18.14. Sewage

The administration and hub site will be connected to the city system. The core shack installation is already connected to Val-d’Or’s services. No used water treatment system is planned on the production site, only portable toilet will be used.

18.15. Water treatment plant and settling pond

On surface, all underground water will pumped to the treatment plant where environmentally safe polymers will added to the water before being clarified in geotubes. Over 99% of solids will be captured, and the clear filtrate will be collected and used for production water or discarded into the environment. The solids will remain in the geotube bag. When full, the geotube containers and their contents can be deposited at a landfill or the solids removed and used as a surface cover when appropriate. Ongoing dewatering during mine operations is estimated at 1,814 m3/day based on operating data. In the first 2 years of operation, 196 m3/day is expected to be pumped from existing openings. The maximum capacity of the water treatment plant is design to accommodate 3,600 m3/day. A settling pond of 12,000 m3 will be necessary to collect water treated from the underground operation. This water can be used again in the underground operation; any overflow can be returned to the environment.

18.16. Mineralized material stockpile

A mineralized material stockpile will be erected near the North Ramp, and another one near the South Ramp entrance. The proposed stockpile will have a capacity of about 15,000 tonnes. The planned capacity will allow managing low grade material to be resampled if required. The mineralize material will be transported on a daily basis to the selected mill for treatment. The mineralized material stockpile is a non-permanent infrastructure. At the end of mine life, it should be completely depleted so that the land may be returned to its initial appearance as part of the mine closure plan.

18.17. Waste stockpile

A waste stockpile will serve as a permanent storage infrastructure for the waste rock extracted from the underground mine. It will be erected near each entrance of the mine, as close as possible to the underground portal in order to benefit from the following advantages:

Shortened haulage distance, which will reduce operating costs; Decreased environmental impact by reducing on-site haul road

construction. For the North ramp, the waste pile will cover about 25,000 m2; at the South ramp it will cover 30,000 m2. Depending on the results of geochemical characterization, it should be possible to sell waste material on the local market to serve as construction material.

www.innovexplo.com


18.18. Overburden stockpile

No overburden dump has been designed. The ramp entrances and escape manways have been located to minimize overburden displacement.

18.19. Project implementation schedule

Following the permitting and financing process, the construction period will take 6 months and preproduction is anticipated to take 2 years. The following will be built during that period: access roads, water pond, electric line, water line, Phase 1 buildings and secondary garage on the production site will be built. The majority of the preproduction expenses will be for ramp construction and for sufficient development of mineralized zones, or working faces, to conduct mining at the proposed mining rate and mill throughput. Ramp construction will commence in the second quarter of preproduction in the Parallel Zone where a 15 m vertical rock face outcrop located at surface provides an ideal location to construct the portal. Ramp construction at the Triangle Zone would commence in the third quarter of preproduction, where overburden is estimated to be between 1 and 5 m thick.

www.innovexplo.com


19. MARKET STUDIES AND CONTRACTS

19.1. Market studies

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.

19.2. Contracts

No contracts have been assigned considering the early stage of the Project.

www.innovexplo.com


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

20.1. Regulations and Permitting

The first approval required for such a project may be the Environmental Impact Assessment (“EIA”). At the federal level, this is dictated by the 2012 Canadian Environmental Assessment Act (“CEEA”) and the accompanying Regulations Designating Physical Activities. The regulation specifies that any proposed gold mine with a production capacity of 600 tonnes per day (600 tpd) or more must carry out such as EIA. The recently introduced Regulations Amending the Regulations Designating Physical Activities maintain this designation. The Québec government has recently passed an amendment to the existing Mining Act to update it, thereby reflecting new realities. A key environmental consideration amongst these amendments is that all projects (except rare earth minerals) with a production capacity of 2,000 tpd or more will now be subject to an EIA and a public hearing under the Bureau d’audiences publiques sur l’environnement (“BAPE”) process. Proponents of smaller projects, while being exempt from the formal EIA and BAPE processes, will be responsible for their own stakeholder consultations. So considering the actual planned level of production of less than 2,000 tpd, the Lamaque Project will not have to conduct a provincial EIA. The CEAA (2012) allows for the federal EIA process to defer to a provincial process, if the provincial process is considered to be the equivalent of the federal process, and if the province so requests in relation to a given project. In such an instance the federal Minister of Environment will base his decision upon the provincial EIA report. The environmental baseline study of the Lamaque Project site recently completed by AMEC on behalf of Integra Gold Corp. will contribute substantially to the production of such a document, if required. The baseline information may also support various other applications for approvals, as well as provide important input to a closure plan. The federal Fisheries Act has been amended and came into force on November 25, 2013. A major change in the Act relates to the categories of waters to which it applies, these waters now being limited to those supporting commercial, recreational or Aboriginal fisheries. The old Policy for the Management of Fish Habitat has been replaced by a new Fisheries Protection Policy Statement. The waters on the Site do not directly support a commercial, recreational or Aboriginal fishery, nor do the fish species indicated during the baseline survey contribute to the ongoing productivity of such a fishery. Therefore, the fishery and habitat protection aspects of the federal Fisheries Act will not apply to this Project. However, the Metal Mining Effluent Regulations (“MMER”) will apply in some form. These regulations, written under the auspices of the federal Fisheries Act and administered by Environment Canada, are currently under review. All aspects are being considered, including the lowering of regulated effluent limits of specified substances, the addition of new substances and their limits, the Environmental Effects Monitoring (“EEM”) requirements, the acute lethality tests, and consideration of selenium. The regulations will also be extended to include coal and diamond mines (which will require a re-naming). The schedule for the amended regulations

www.innovexplo.com


coming into effect is uncertain; multi-stakeholder consultations are underway and are expected to extend to at least late spring of 2014, after which it will be up to the Minister of Environment to decide upon the publication of the amendments in the Canada Gazette. Key provincial permits will be required during the construction and operation of the mine. The Ministry of Sustainable Development, Environment, Wildlife and Parks (“MDDEFP”: Ministère du Développement durable, de l’Environnement, de la Faune et des Parcs) is the Québec entity responsible for environmental protection and the conservation of biodiversity to improve the environmental quality of life. This department is responsible for the control and enforcement of laws and regulations concerning environmental protection, including the analysis of requests for authorizations and for permits. The department also works at the level of laws and regulations for the prevention or reduction of the contamination of water, air, and soil, drinking water quality, combating climate change, as well as conserving and protecting wildlife and its habitats. It is also the lead agency for the environmental assessment of projects and environmental issues, as is the case for impact studies. The Environment Quality Act (EQA) of Québec is divided into two chapters. Chapter I sets out provisions of general application while Chapter II outlines the particular provisions of James Bay and Northern Québec covered by the James Bay and Northern Québec Agreement. The Lamaque Property is located south of the James Bay territory so that only Chapter I is of interest for the Project. The main articles of Chapter I associated with obtaining environmental certificates of authorization or authorizations are articles 22 (general case), article 31.1 (environmental impact assessment studies), article 32 (drinking water and domestic wastewater), and article 48 (atmospheric emissions). In theory, during operations, the company would not be subject to a de-pollution attestation under section 31.11 since no ore processing or enrichment are anticipated to be carried out on site. Under subsection 31.1 of the EQA: “No person may undertake any construction, work, activity or operation, or carry out work according to a plan or program, in the cases provided for by regulation of the Government without following the environmental impact assessment and review procedure and obtaining an authorization certificate from the Government.” Following review of the Québec mining act in December 2013, there are a number of other significant amendments of an environmental or socio-economic nature. A specific First Nations consultation policy is to be developed by the Québec government, and the bill prohibits the expropriation of aboriginal burial grounds. Before a mining license can be issued, a project must file a restoration and rehabilitation plan which must be approved, and which must be accompanied by a security deposit of 100% of the estimated cost of implementation (the previous deposit being 70%). The project must have a Certificate of Authorization (“CofA”) under the EQA; the formalities for obtaining such a CofA may be streamlined if the

www.innovexplo.com


Minister deems them unreasonable or detrimental to the realization of the mining project under consideration. The bill gives regional county municipalities the right to designate portions of their lands as "incompatible" with mining activity or subject to specified conditions, thereby empowering a host of 3rd parties to trump the responsible provincial agencies. The Project is subject to Section 22 of the Québec EQA for the operation of a mine. The application to the MDDEFP for this CofA will be accompanied by sufficiently comprehensive studies to address the requirements of Directive 019 for the Mining Industry. In addition, the proponent is planning to obtain other permits, authorizations, approvals and leases from both the Ministry of Natural Resources (“MRN”: Ministère des Ressources naturelles) and the MDDEFP for various components of the overall Project development work, as required. These applications will be submitted as part of the ongoing process of developing the site and should therefore not impact the Project critical path schedule. These may include:

Management of the overburden material to be reused at closure stage; Stripping works that could result in disrupting the soil, water or hydraulic

regime; Bulk sampling; Electric substation if more than 120kv; Treatment of mine waste water; Sinking of ramps and shafts; Dewatering and dry maintenance; Storage of hazardous materials; High risk petroleum products containment installation and/or oil separators; Ore and waste rock storage; Underground backfill; Borrow pits or quarry; Atmospheric emissions purification devices; and Site restoration.

The established fees to be submitted to the governmental agency with the applications for these various permits/approvals are normally between $1,000 and $5,000 for each of the requests.

20.2. Environmental Baseline Study

An environmental baseline study of the Site was undertaken on behalf of Integra Gold by AMEC Environment & Infrastructure (“AMEC”) during 2013. The study reviewed available information across a number of disciplines, including geology and soils, hydrogeology, hydrology, air quality and noise, flora and fauna, socio-economic setting, and archaeology. The Lamaque Project site (the “Site”) is located in the Province of Superior and sub-province of Abitibi, near the Cadillac Deformation zone. This region is known as s greenstone belt, due to the degree of rock deformation and metamorphism forming

www.innovexplo.com


greenschist facies. The area around Cadillac, Malartic and Val-d’Or hosted many operating mines in the past. The surficial deposits on-site are mostly of glacial origin, but the northern portion of the Site is occupied by a tailings impoundment, the result of earlier mining on an adjacent property. Surficial deposits in the other portions of the site include: glacial lake deposits (e.g., mostly stratified sand and silt), marsh deposits (e.g., organic), glacio-fluvial deposits (e.g., sand and gravel esker) and glacial deposits (e.g., till). The Abitibi region is known for the quality and quantity of groundwater contained in its eskers. The Lamaque Project is partially located within the urban limits of Val-d’Or, where the aquifer potential appears average. Val-d'Or uses the water contained in Harricana’s interlobate moraine for supply of drinking water for the population. This aquifer, which crosses Val-d’Or in a north-south axis, has branches that run adjacent to the Lamaque gold property (depending on the map of surface deposits consulted). This proximity should be the subject of further study, given the potential impact of a hydraulic connection between the groundwater contained in the moraine and the Site, particularly with respect to mine dewatering. Nevertheless, it should be noted that when adjacent Sigma and Lamaque mines were in operation, no concern was identified concerning water supply for the town. Sampling and analysis of groundwater from eight exploration wells on the Site indicated that several constituents exceeded both provincial and federal standards, suggesting that mine water would require treatment prior to discharge to the receiving environment. The Site is located within a large watershed of approximately 710 ha drained by a minor stream without name. The stream is typical of wetland regions with thick aquatic vegetation lining the shores, varying widths and depths, and a bed composed of a mixture of organic material and small stones. The Site experiences frequent precipitation events; on average, the Site receives 914 mm of precipitation, of which about 314 mm falls as snow. Under extreme conditions, the stream can be expected to collect a peak flow reaching 8 m3/s for a 1-in-100 year rain event. The Abitibi-Témiscamingue region of western Québec experiences a sub-Arctic, humid continental climate which is characterized by short warm summers and long, cold, snowy winters. No ambient air quality measurements were available for the Site or in close proximity to it. Regional National Air Pollution Surveillance program stations (55 km to the north-east, 94 km to the west) indicated that ozone and PM2.5 at times exceeded provincial and/or federal standards while SO2 was at acceptable levels. A dispersion model of the projected emissions from the Canadian Malartic Mine (Genivar, 2008) some 30 km from the Site indicated that only PM2.5 demonstrated slight exceedances of provincial standards. The only greenhouse gas (GHG) emitter registered around the Site was the Canadian Malartic Mine in 2011. These emissions constituted 0.1% of the total provincial emissions, and projected GHG emissions from the Integra Gold property are expected to be significantly lower. The majority of the Site is characterized by organic deposits of variable thickness generally concentrated in low-lying areas. The southern part of the study area is characterized by a mixture of terrestrial and wetland ecosystems. Many wetlands are present on the edge of the tailings impoundment area and they occupy about 50% of

www.innovexplo.com


the study area. Wetland ecosystems are protected under provincial legislation and authorization is required prior to their disturbance. Some logging is also present. No “at risk” plant species were observed on the Site, and no exceptional forest ecosystem was identified. Finally, no wildlife habitats, as described in the Regulation respecting wildlife habitats, are present on the Site. Woodland caribou, a vulnerable species in Québec, frequent a so-called “buffer zone” some five km from the Site. The presence of 12 mammal species was identified, including four bat species, of which two are considered “endangered”, but are provided no legal protection under provincial legislation. Of the 91 bird species identified on the Site, three “species at risk” have been identified: these are also not provided with legal protection. Two species of snakes and five species of amphibians were also observed. None of these species present an impediment to Site development, as long as certain mitigation measures are observed. One stream, with two tributaries arising from wetlands on the Site, provides habitat for six species of fish and nine taxa of benthic invertebrates. None of the fish species are exploited by commercial, recreational or Aboriginal fisheries. Surface water quality analyses indicated levels exceeding provincial and/or federal criteria for some substances, indicating possible contamination from the old tailings. Sediment samples also suggested possible contamination from previous gold mining efforts, and were categorized as “probable effect concentration” in terms of protection of aquatic life.

20.3. Operations

There will be no tailings impoundment area on the Site, as the plan is to transport the ore off-site to an existing processing facility where tailings will be managed. Waste rock will be disposed of on-site in a dedicated storage pile, with control and capture of run-off from the waste rock to treat as may be indicated prior to discharge to the environment. Mine water will require a wastewater treatment system prior to discharge to the receiving environment, as indicated by the results of the hydrogeological investigation described above. Finally, the overall Site drainage will require a management plan in order to ensure that there is no contamination of water flows resulting from contact with the on-Site transportation network including vehicle fueling stations, waste rock disposal piles, on-Site ore stockpiles, and other work areas. At the operational phase of the Project, the outfalls of wastewater treatment systems and any other water discharged to the receiving environment will need to be monitored in keeping with the requirements of the federal Metal Mining Effluent Regulations, and the provincial Directive 019, to ensure that concentrations of specified contaminants are in compliance with acceptable levels. In addition, Environmental Effects Monitoring (“EEM”) will need to be conducted, again in keeping with the requirements of the MMERs. These regulations are currently under review (see point (c) below). A targeted water quality monitoring program will be developed as an integral part of the mine closure plan and the site reclamation plan. It will be consistent with the requirements of the MMERs and Directive 019. This program will be intended to demonstrate the effectiveness of the implemented closure and reclamation measures in preventing any contamination from the mine and disposal area workings from moving into the surrounding environment. The program will concentrate on the receiving waters of the principle stream crossing the Site, and will encompass both

www.innovexplo.com


upstream (control) and downstream sampling stations of those areas which conceivably could contribute some form of contaminant, should the closure and reclamation measures be to some degree lacking in effectiveness. The program will be implemented over a minimum period of five years following reclamation.

20.4. Reclamation

Where selected exploration work has been conducted on land, as per Article 108 of the Regulation respecting mineral substances other than petroleum, natural gas and brine, a rehabilitation and restoration of this land is required under Article 232.1 of the Mining Act. Several of these designated works are likely to be carried out on the site, including:

Excavation for the purpose of mining exploration, including any of the following elements:

A movement of 1000m3 or more of overburden; Rock stripping or the removal of overburden covering an area of more than

10,000m2; The extraction or removal of mineral substances at a volume of 500 metric

tons or more for the purposes of geological and geochemical sampling. Work performed with respect to material deposited in storage areas, in

particular one or more of the following activities: Boreholes; The excavation, removal or sampling of accumulated or cover materials; Works related to underground mining, including the following activities: The sinking of access ramps, wells, or any other excavation; Dewatering of wells and maintaining the excavations in a dry state; The rehabilitation of worksites or other underground structures; The delivery of mineral substances to the surface. The construction of storage areas pertaining to the activities referred to

above. A plan for rehabilitation and restoration will be submitted to the MRN in order to obtain authorization for any of the exploration or development work mentioned above. This plan must be prepared according to the “Guidelines for preparing a mining site rehabilitation plan and general mining site rehabilitation requirements” (1997). The plan must be reviewed every 5 years, but significant changes to the Project might also trigger the need for update, at the request of MRN. Since 7 August 2013 (Decree 838-2013), the financial guarantee must cover the full costs (100%) of the site restoration plan and it must accompany the submission of that restoration plan. Recent regulatory changes also imply that the restoration plan can no longer be simply "submitted" but must be "approved" by the MRN before receiving authorization for the work referred to above. Rehabilitation and restoration will form a major part of the closure plan. Buildings and other infrastructure will be dismantled and demolished at the end of mining activities. The impacted area will then be managed in keeping with the requirements of the

www.innovexplo.com


EQA and backfilled with overburden and top soils to re-establish vegetation on the site. A natural water regime will be re-established on the site. Management of dismantled materials will be done in compliance with the Guide des bonnes pratiques sur la gestion des matériaux de démantèlement, published by the MDDEFP. Non-hazardous residual materials will be managed by disposal at the approved landfill site nearby. Overburden will be kept in a proper storage area to be used at the end of the mine’s life, for recapping and rehabilitation work. The organic portion of the material will be segregated from the lower soil horizons. The storage of these materials will require a governmental authorization. The EQA requires, from Appendix III of the Regulation for the Protection and Rehabilitation of Lands (“RPRL”), that an Environmental Site Assessment (phase I and II) is performed in the months following the cessation of activities. Should any identified contaminants exceed the limits stated in Appendix II of the RPRL, a Rehabilitation Plan would have to be submitted to the MDDEFP. Following its approval, the company would have to conduct rehabilitation works in compliance with the plan and in a manner compatible with future site utilization. Waste rock piles are generally considered safe for the long term. Integra Gold will investigate the possibility of reutilizing the material for backfilling of galleries, or the Company might choose to use it for its civil needs at surface, or to offer the material to local entrepreneurs for civil works and thus reduce the use of borrow-pit material for construction. A monitoring program will be conducted periodically for the evaluation of the integrity of structures. This will consist of visual inspections of any dikes, crown pillars and any anomalies that could jeopardize stability. Particular attention will be paid to water exit points to note any signs of erosion. Visits will be made once a year. An environmental monitoring program will be conducted during closure. There should be no effluent following the end of activities, and thus no post-closure effluent monitoring is planned. Agronomic monitoring will be undertaken after closure, in the areas of waste rock piles, buildings and other facilities footprints. This program will consist of annual visits and will focus on an assessment of the percentage of vegetation recovery in selected sample plots and, if required, recommendation on amendments (fertilization and reseeding) of specific areas. It is expected that the vegetation will become self-sustaining after two or three years. Costs related to the rehabilitation and restoration of the site is estimated to be approximately $2.4 million, in 2013 dollars (see Table 201.1). This estimation is according to the recent plan provided for the proposed facilities. Québec now requires that the costs of financial security correspond to 100% of the amount provided for in the restoration plan. And because the Project is still at a preliminary stage, Integra Gold could expect the MRN to require a single up-front payment.

www.innovexplo.com


Table 20.1 – Cost Estimate for Restoration and Financial Guarantee - Lamaque Property Project, Integra Gold

Work related to: Quantity Unit Unit Cost Amount Financial Guarantee

Buildings and equipmentsDismantling (including transportation and disposal costs) 500,000 $ 500,000 $ Scrap metal related to dismantling (1) 300 tons 175 $ 52,500 $-

Accumulation Areas (waste rock piles)Re-establishment of vegetation (2) 1.9 ha 18,000 $ 34,200 $ 34,200 $

Accumulation Areas (Ore piles)Site sloping & reprofiling (3) 0.5 ha 12,000 $ 6,000 $ 6,000 $ Re-establishment of vegetation 0.5 ha 8,000 $ 4,000 $ 4,000 $

Surface Infrastructure Environmental Site Investigation (ESA Phase I & II) 100,000 $ 100,000 $ Dismantling - Pumping stations, basins 2 unit 15,000 $ 30,000 $ 30,000 $ Sludge management & disposal / soil remediation 5000 tons 125 $ 625,000 $ 625,000 $ Site sloping & reprofiling (3) 23.6 ha 12,000 $ 283,200 $ 283,200 $ Re-establishment of vegetation 23.6 ha 8,000 $ 188,800 $ 188,800 $

Site Safety after closureRestrict access (fences, signs), including for roads 4 unit 2,000 $ 8,000 $ 8,000 $ Other safety measures (seal off openings, barricades) 4 unit 10,000 $ 40,000 $ 40,000 $

Monitoring Program for post mining - 5 yearsEnvironmental follow-up of the mine effluent (5) 5 year - $ - $ - $ Agronomical monitoring 2 visit 5,000 $ 10,000 $ 10,000 $ Groundwater follow-up (4) 5 year 6,000 $ 30,000 $ 30,000 $ Geotechnical Integrity of structure 9 visit 5,000 $ 45,000 $ 45,000 $

Monitoring Program for post restoration - 5 yearsEnvironmental follow-up of the mine effluent (5) 10 year - $ - $ - $ Agronomical monitoring 2 visit 5,000 $ 10,000 $ 10,000 $ Groundwater follow-up (4) 5 year 6,000 $ 30,000 $ 30,000 $ Geotechnical Integrity of structure 5 year 5,000 $ 25,000 $ 25,000 $

Sub-Total of the work subject to the financial guarantee 1,916,700 $ 1,969,200 $

Supervision 10% 191,670 $ 196,920 $ Contingency 10% 210,837 $ 216,612 $

Total : 2,319,207 $ 2,382,732 $

(4) A total of 20 monitoring wells is considered for the environmental follow-up. Wells will target the 'North Ramp' (Parallel Zone), 'South Ramp' (Triangle Zone)

and 'Industrial Area' (along HW 117) sectors.

(5) No effluent assumed at site closure.

Notes:

(3) For the concerned area, transportation & placement costs are estimated at 8$/m3. A layer of 50cm thick is currently considered.

(2) For hydroseeding, the average price is in the range of 18 000 to 35 000$/ha.

(1) According to the market price.

in 2013 CAD dollars

www.innovexplo.com


20.5. Socio-Economic Setting

An investigation of socio-economic information was carried out as part of the environmental baseline study referenced above. This investigation relied solely upon information available in the public domain and did not collect any primary information. The Lamaque Project is located in the Abitibi-Témiscaminque Administrative region (08) in the regional county municipality (RCM) of the “Vallée-de-l’Or. The Project area falls entirely within the territory of the municipality of Val-d’Or. Responsibility for land use planning is divided between the Ministry of Natural Resources, the MRC de La Vallée-de-l’Or, and the municipality of Val-d’Or. According to the Val-d’Or zoning plan, the Lamaque Project is located in an area zoned as natural resources, within which mining operations are a conforming use. The northwest portion of the Lamaque Project where no mining operation is planned, is located inside the Val-d’Or urban perimeter, a zone mainly designated and used for residential and commercial purposes where no extractive industries or high impact industry is permitted. Also, wood harvesting rights are present in the south-eastern portion of the Project. The Public Land Designation plan of the Abitibi-Témiscamingue region did not identify any protected designated zones that are located on or the adjacent land to the Lamaque Project. Also, no category I land as per the James Bay and Northern Québec Agreement Aboriginal land classifications, and no aboriginal trap lines are located on the Lamaque Property. The Project area lies in the Algonquin Anishinabeg Nation Tribal Council land assertion; the Algonquin-Anishnabe community of Lac-Simon will likely have interests in the Project. An Archaeological Potential Study was conducted and revealed no areas of interest in the proposed mining sector of the Project. Two sites of historical interest (the Bourlamaque Mining Village and the historical former Lamaque Mine) are located north of the Lamaque Project. The Forestel Hotel is located on the adjacent property northwest of the Project. Also, the Trans-Québec network of snowmobile trails, which is a regional recreational focus area, overlaps with the south portion of the Lamaque Property. Operations and truck movement as well as vibration from blasting could be a major issue for the Forestel Hotel, the historical sites, the residential areas adjacent to the Project and the snowmobile track activities. Integra Gold has prepared a stakeholder and aboriginal engagement plan in order to meet all interested parties and identify, address and mitigate any negative social and economic effects of the Lamaque Project which may be of concern. This plan will also include the collection of primary information regarding a number of identified information gaps relating to socio-economic factors and infrastructure facilities in the area. Integra Gold has a corporate website on which it posts relevant project information, and it issues periodic news releases. As well, the Company has informed stakeholders and held 3 public meetings to present the project and gather concerns. A compilation of the meetings indicates general concerns as noise,

www.innovexplo.com


vibration, traffic circulation, environment protection and visual impact. At this time it has not entered into agreements with any of those stakeholders. In Québec, there is no legal obligation to consult with stakeholders regarding mining closure plans. Nevertheless, Integra Gold considers it best practice to engage the interested parties to determine the intended land use of mining sites after closure. Due to the proximity of the Lamaque Project to Val-d’Or, the latter is an important stakeholder to consult regarding zoning and future land use plans in order to close and rehabilitate the site in compliance with the municipality’s land use requirements. Also, Integra Gold intends to engage the snowmobile clubs network, the Forestel Hotel, the Bourlamaque Village and the Cité de l’Or in order to provide the Company with additional input from directly adjacent stakeholders to the design of the eventual end-use of the Lamaque Project. It is notable that the recent amendments to the Mining Act mandate the creation of project-specific monitoring committees to ensure that all commitments made by the Proponent are implemented. Such committees will stay in place until the completion of the lease holder’s site rehabilitation and restoration plan.

www.innovexplo.com


21. CAPITAL AND OPERATING COSTS

21.1. Capital costs

The preproduction costs are estimated at $69.2M, net of production revenue received during the second year of the preproduction period ($37.4M) (Table 21.1). Preproduction capital costs include surface infrastructure (site preparation, roads, power lines, and water lines), installation of modular buildings for offices and garages (mechanical and electrical shops, stockroom), mining infrastructure at the North and South sites, mobile equipment, development and capitalized operating costs, owner costs (closure costs in line with required financial guarantees, company staff costs, and indirect costs) as summarized in the following tables. Preproduction capital costs are minimal given that there is no need to build processing and tailings facilities, and that mineralization is spatially close to surface. 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.

Table 21.1 – Capital cost estimate Description Preproduction Sustaining Total

Surface infrastructure* 12.9 M$ 4.7 M$ 17.7 M$ Mining infrastructure * 6.9 M$ 3.1 M$ 10.1 M$ Mobile equipment ** 14.8 M$ 17.3 M$ 32.2 M$ Develop. & capitalized operating costs ** 55.6 M$ 39.0 M$ 94.6 M$ Owner’s costs 16.3 M$ 2.6 M$ 18.9 M$ Offsetting capitalized revenue (37.4) M$ (37.4) M$

Total 69.2 M$ 66.8 M$ 136.0 M$

* contingency 20%, ** contingency 10%.

Integra Gold is also studying a scenario that would involve delaying the development of the South Ramp by 12 to 18 months in order to reduce up-front capital cost requirements and use cash flow from the North Ramp to fund development of the South Ramp.

21.1.1. Surface infrastructure

Surface infrastructure includes site preparation, access roads, installation, buildings, and the electrical distribution and communication system. A 20% contingency has been applied (Table 21.2).

www.innovexplo.com


Table 21.2 – Surface infrastructure cost estimate

All costs were estimated using budget quotes provided by suppliers and based on existing comparable projects. The site preparation costs were estimated at $1,543,635 for the North Ramp and $2,133,934 for the South Ramp, for a grand total of $3,677,569. These amounts include the costs of surface acquisition, deforestation, on-site roads, ramp portals, excavation, site infrastructure arrangement, escapeway setups, diesel tanks, used oil storage and the grouting of existing DDH. Building costs include the office, the main and secondary garages, the warehouse, the dry, the communication system (IT), and the protection for the powder-magazine. For the main buildings, it was assumed that a down payment of 25% would be paid on mining equipment with the balance paid over 5 years at a 6% interest rate. The cost of the electrical distribution & communication system for surface and underground includes the power distribution, cables and connectors, instrument and communication, lighting and accessories. Mining infrastructure Mining infrastructure includes water management and distribution, ventilation and air heating, mine dewatering, and compressed air distribution. A 20% contingency has been applied (Table 21.3).

Table 21.3 – Mining infrastructure cost estimate Description Preproduction Sustaining Total

Water management and distribution 2.6 M$ 2.6 M$ Ventilation and air heating 3.0 M$ 1.3 M$ 4.3 M$ Mine dewatering 0.5 M$ 1.1 M$ 1.5 M$ Compressed air distribution 0.9 M$ 0.9 M$ 1.6 M$

Total 6.9 M$ 3.1 M$ 10.1 M$

Description Preproduction Sustaining Total

Site preparation and installation 3.7 M$ 3.7 M$ Buildings 4.8 M$ 3.4 M$ 8.1 M$ Electrical distribution & comm. system 4.5 M$ 1.4 M$ 5.8 M$

Total 12.9 M$ 4.7 M$ 17.7 M$

www.innovexplo.com


Water management and distribution costs include the water treatment plant and its arrangement, the settling pond and surface piping. Ventilation and air heating costs include escapeway accommodation and underground main fan setup, the heating system, secondary fans and rigid conduit for long drive excavation. Mine dewatering includes the development pumping system and the main pumping station. Compressed air distribution includes equipment and installation.

21.1.2. Mobile equipment

Mobile equipment includes all surface and underground equipment. A 10% contingency has been applied to budgetary quotes obtained from equipment suppliers. For the economic evaluation, it was assumed that a down payment of 25% would be paid on mining equipment with the balance paid over 5 years at a 6% interest rate. Residual value was limited to 25% to 35% of original depending on years of use.

21.1.3. Development and capitalized operating costs

Development and capitalized operating costs include all preproduction development and main development (ramp, escapeway) from Years 1 to 2 of the production period (Table 21.4).

Table 21.4 – Development and capitalized operating costs Description Preproduction Sustaining Total

Definition drilling 0.8 M$ 0.8 M$ Stope development 10.3 M$ 10.3 M$ Mining cost 11.0 M$ 11.0 M$ Main development (sustaining only) 23.5 M$ 39.0 M$ 62.5 M$ Energy cost 4.4 M$ 4.4 M$ Milling and transportation 5.7 M$ 5.7 M$

Total 55.6 M$ 39.0 M$ 94.6 M$

Development and capitalized operating costs include definition drilling, stope development (10% contingency), mining cost (10% contingency), milling and transportation, and energy.

21.1.4. Owner’s costs

Owner’s costs include the financial guarantee reimbursement, NSR payment environment-related costs, staff and material required for preproduction. Owner’s costs amount to $16.3M for preproduction and $2.6M for sustaining.

www.innovexplo.com


Table 21.5 – Owner’s costs Description Preproduction Sustaining Total

Integra Gold staff 13.1 M$ 13.1 M$ Environment 1.4 M$ 1.4 M$ Financial guarantee reimbursement 1.8 M$ 0.6 M$ 2.4 M$ NSR 2.0 M$ 2.0 M$

Total 16.3 M$ 2.6 M$ 18.9 M$

21.1.5. Capitalized revenue

During the 24-month preproduction period, it is anticipated that 28,000 ounces of gold will be produced, providing revenue of $37.4M (US$1275/oz and CAD/USD of 1.05). The preproduction revenue was capitalized.

21.2. Operating costs

Operating costs are summarized below for the production period (Table 21.6). Given that this PEA presents a toll milling scenario and that Integra Gold has the ability to process mineralized material recovered during the preproduction and development stage, revenue generated from these ounces has been included in forecasted cash flows. A total of approximately 28,000 ounces are anticipated to be produced during Year 2 of the preproduction phase.

Table 21.6 – Summary of total operating costs

21.2.1. Definition drilling

InnovExplo has estimated the cost of definition drilling at $5.00/tonne with an additional $1.00/tonne 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 $11.7M or $6.00 per tonne milled. On a per-metre basis, for an average $75/m unit

Description Total cost Unit cost (Years 1-5) ($/t) ($/oz)

Definition drilling 11.7 M$ 6.00 24.58 Stope development 34.9 M$ 17.86 73.16 Mining cost 115.3 M$ 58.94 241.51 Integra Gold Staff 38.8 M$ 19.82 81.21 Energy cost 15.0 M$ 7.64 31.31 Milling and transportation 89.4 M$ 45.69 187.21 Environment 4.1 M$ 2.08 8.54

Total 309.3 M$ 158.04 647.53

www.innovexplo.com


cost for drilling and sampling, about 37,000m of drilling will be done on an annual basis during production period.

21.2.2. Stope development

The unit cost for stope preparation stands at $17.86 per tonne milled (based on tonnage milled assigned to production), and consists of 61% for the 4 m x 3 m drifts and ramp excavation or enlargement, 33% for drifts smaller than 3.5 m x 3 m, and 5% for raises. A 10% contingency was added to drift and raise development. Development costs include material (explosive, ground support, piping installed) and manpower.

21.2.3. Mining costs

Mining costs include stoping, auxiliary equipment operating costs and manpower. A 10% contingency has been applied (Table 21.7).

Table 21.7 – Mining costs Description Total cost Unit cost (Years 1-5) ($/t) ($/oz)

Stoping $76.8 M 39.27 160.88 Auxiliary equipment operating cost $3.6 M 1.83 7.52 Manpower $34.9 M 17.84 73.11

Total $115.3 M 58.94 241.51

Stoping costs includes material and manpower for room and pillar, long-hole and material and maintenance for haulage. The cost for material handling is estimated to range from $8.70/tonne to $13.10/tonne, including material, maintenance and manpower. Long-hole stoping costs amount to $20.24/tonne for the North Ramp and the Triangle Zone, and $17.93/tonne for No. 4 Plug. Room and pillar stoping costs amount to $61.38/tonne. The auxiliary equipment operating cost includes all services equipment, such as graders, tractors, etc. Manpower includes service workers, supervision, construction miners, scoop and truck operator and trainer. All these costs include 10% contingency.

21.2.4. Integra Gold staff

The staff and associated salaries include administration, technical services, site security, mechanical and electrical personnel. Salaries were evaluated based on experience and other projects near Val d’Or. To account for benefits, 33% was added and depending on the job, bonuses of 5% to 20% were also included. The estimate of the department’s general operating cost was based on comparable mine operating budget. The average cost for Integra Gold staff is $19.82 per tonne milled.

www.innovexplo.com


21.2.5. Energy cost

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 of surface equipment is already included in unit cost (development, mining and haulage) or department general operating cost. The estimated average annual electrical consumption for Years 1-4 is 31,649,584 kWh, representing an annual cost of $2,099,871 (Table 21.8). The electrical consumption cost is based on Hydro-Québec M rate. The estimated annual propane consumption for Years 1-4 is 2,170,870 litres per year, amounting to $1,302,522 per year at a price of $0.60/litre (budget quotation from Propane Nord-Ouest). The propane tank rental cost $7,800/year. As shown in Table 21.8, the estimated total annual energy cost is $3.4M, representing an average of $7.49 per tonne milled for years 1-4.

Table 21.8 – Yearly energy cost (average for Years 1-4) Description Yearly cost Electricity $2,099,871 Propane $1,302,522 Propane tank rental $7,800 Total: $3,410,193

21.2.6. Milling and transportation

The LOM average tonnage is approximately 1,480 tpd, and varies between 1,280 and 1,650 tpd depending on the period (based on 312 operating days per year). This production rate is consistent with potential milling options in the immediate area, and may change during actual production depending on which processing facility is used. This includes resource extraction from both ramps thereby minimizing undue pressure put on any one point of production and reducing potential bottlenecks while mining. The Company was able to identify which mills are best suited for material from the Lamaque Project, which has assisted in the determination of the $45.69 per tonne milling and transportation assumption. The unit cost for milling is estimated at $35.00/t, and at $10.69/t for truck loading and transportation.

21.2.7. Environment

The environmental operating cost is based on similar operation. Manpower cost are already included in Integra Gold staff. The actual environmental cost cover annual monitoring for noise, vibration, effluent water quality and underground water based on actual regulation. Water treatment cost was evaluated based on projected pumping rate and a unit cost of $0.50/m3. The cost for management and disposition of waste and hazardous material is included. The average costs for environment is estimated to $2.08 per tonne milled. The yearly estimated environmental cost is $977,680 (Table 21.9).

www.innovexplo.com


Table 21.9 – Yearly environment cost Description Yearly cost Personnel $378,600Air & noise quality monitoring $120,000 Water treatment $330,000Waste management $90,000Others $59,080Total: $977,680

21.2.8. Capitalized Opex

The operating costs incurred during the preproduction period ($46,596,373) were capitalized.

21.2.9. Taxes and Royalties

The Lamaque Project is subjected to the following taxes: Québec mining rights; Federal and provincial taxes.

The Lamaque Property is subjected to an NSR royalty consisting of 2% applied on net smelter revenue. At commercial production, with a $1M payment for the North Ramp and $1M for the South Ramp, the NSR Royalty will decrease to 1% .

www.innovexplo.com


22. ECONOMIC ANALYSIS

22.1. Financial Analysis

An after-tax model was developed for the Lamaque Project. All costs are in 2013 Canadian dollars with no allowance for inflation or escalation. The Lamaque Project is subject to the following taxes:

- Québec mining tax rate of 16% (2014 rate); - Income tax rate of 26.9% (federal and provincial).

The Lamaque Property is subject to a royalty in favor of Teck, equal to 2% of NSR and the buyout of 1% of the NSR for $2M is included in the economic model. 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. The following parameters were considered in the financial analysis (Table 22.1):

An average gold price of $US1,275/oz and an exchange rate of 1.05CAD/1USD (lower than 3-yr trailing average as of February 28, 2014).

Milling recovery: o Parallel Zone: 97% o Fortune zone: 95% o Triangle zone: 90% o No. 4 Plug zone: 86% o Resulting in an average recovery of 92.2% for the entire Project.

Royal Mint fees of $3/oz. Royalty of 2% of Net Profit for preproduction, and a payment of $2M and a

royalty payment of 1% of Net Profit for the production period. 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. The resulting main parameters and cash flow analysis are presented in Table 22.1. The economic analysis for the Lamaque Project is presented in Table 22.2.

www.innovexplo.com


Table 22.1 – Cash flow analysis summary

Parameters Results




Average annual gold production (ounces/year):

112,400




Preproduction period (years) 2

Mine life (years) 4.25

Cash Cost per Au Ounce (CA$/oz) 665

Cash Costs and Sustaining Cost per Au Ounce (CA$/oz) 805

PRE-TAX




AFTER-TAX




www.innovexplo.com


Table 22.2 – Economic analysis for the Lamaque Project (figures in Canadian dollars)

Q1 Q2 Q3 Q4 Total

PRODUCTION

North Zone

PARALLELE

Development (t) - - - - 59,573 32,987 6,275 18,925 - 117,760

Grade (g/t) - - - - 6.70 5.99 13.68 7.08 - - 6.94

Long Hole (t) - - - - - 106,476 164,098 38,287 73,079 - 381,940

Grade (g/t) - - - - - 7.59 6.56 5.31 9.09 - 7.21

Room and pillar (t) - - - - 19,257 78,044 79,689 62,457 26,338 - 265,785

Grade (g/t) - - - - 6.25 7.63 9.54 9.71 9.75 - 8.80


Grade (g/t) 6.59 7.36 7.69 7.89 9.26 7.72

FORTUNE

Development (t) 16,254 25,880 0 42,134

Grade (g/t) 5.18 4.98 0.00 5.06

Long Hole (t) 15,600 76,771 65,071 157,442

Grade (g/t) 5.20 5.11 5.14 5.13

Room and pillar (t) 4,095 14,268 5,848 24,211

Grade (g/t) 5.71 6.20 7.15 6.35

Total (tonne milled) 35,949 116,919 70,919 223,787

Grade (g/t) 5.25 5.21 5.31 5.25

South Zone

TRIANGLE

Development (t) 28,431 34,711 25,353 - - 88,495

Grade (g/t) 7.95 8.05 10.17 - - 8.63

Long Hole (t) 4,000 28,197 26,785 53,779 - 112,761

Grade (g/t) 6.49 6.44 7.36 10.57 - 8.63

Room and pillar (t) 13,230 118,105 170,687 134,840 31,323 468,185

Grade (g/t) 11.83 9.44 12.96 14.53 15.57 12.67


Grade (g/t) 8.95 8.71 11.97 13.40 15.57 11.45

PLUG 4

Development (t) 3,529 25,101 9,287 37,916

Grade (g/t) 3.59 5.06 3.72 4.59

Long Hole (t) 13,415 290,550 80,815 384,781

Grade (g/t) 6.16 5.40 5.88 5.53

Room and pillar (t)

Grade (g/t)

Total (tonne milled) 3,529 38,516 299,837 80,815 422,697

Grade (g/t) 3.59 5.44 5.35 5.88 5.44

TOTAL North and South

Total (tonne milled) 124,491 398,520 512,364 463,724 501,496 80,815 2,081,410

Grade (g/t) 7.46 7.97 9.35 9.25 6.76 5.88 8.19

Parallel Recovery (%) 97% 97% 97% 97% 97% 97% 97%

Fortune Recovery (%) 95% 95% 95% 95% 95% 95% 95%

Triangle Recovery (%) 90% 90% 90% 90% 90% 90% 90%

Plug 4 Recovery (%) 86% 86% 86% 86% 86% 86% 86%

Gold Produced (oz) 28,001 95,493 143,257 127,046 98,666 13,148 505,611

Tonne milled assigned to capital

Gold Produced assigned to capital (oz)

Tonne milled assigned to production

Grade (g/t)

Gold Produced (oz)

REVENUE

Gold Price ($US/oz) $1,275 $1,275 $1,275 $1,275 $1,275 $1,275 $1,275 $1,275 $1,275 $1,275 $1,275

Exchange rate ($CAN/$US) 1.05 1.05 1.05 1.05 1.05 1.05 1.05 1.05 1.05 1.05 1.05

Gold Price ($CAN/oz) $1,339 $1,339 $1,339 $1,339 $1,339 $1,339 $1,339 $1,339 $1,339 $1,339

Gross Revenue $37,486,227 $127,840,608 $191,785,293 $170,083,255 $132,089,419 $17,602,404 $676,887,205Mint (cost 3,00$ per oz) $84,003 $286,478 $429,771 $381,139 $295,999 $39,445 $1,516,834

NSR $748,044 $1,275,541 $1,913,555 $1,697,021 $1,317,934 $175,630 $7,127,726Capitalized revenue $37,402,224 -$37,402,224

Net Revenue $126,278,589 $189,441,967 $168,005,095 $130,475,486 $17,387,329 $631,588,465

OPERATING EXPENDITURES

North Zone

Definition drilling $0 $0 $0 $0 $472,980 $1,305,042 $1,716,062 $1,419,532 $1,022,016 $0 $5,935,631Stope development $0 $0 $48,812 $154,475 $7,421,088 $3,484,002 $2,936,379 $9,046,444 $0 $0 $23,091,199Mining cost $389,976 $389,976 $389,976 $389,976 $4,480,832 $13,049,018 $15,292,181 $12,706,201 $9,255,583 $0 $56,343,720Integra staff $684,011 $684,011 $684,011 $684,011 $4,541,103 $4,774,264 $4,817,410 $4,650,290 $3,522,575 $0 $25,041,685Energy cost $175,500 $175,500 $175,500 $175,500 $1,476,595 $1,738,868 $1,842,839 $1,844,561 $1,210,200 $61,250 $8,876,315Milling and transportation $0 $0 $0 $0 $3,601,743 $9,937,895 $13,067,811 $10,809,733 $7,782,652 $0 $45,199,833Environment $74,704 $74,704 $74,704 $74,704 $475,777 $468,675 $510,881 $492,699 $428,134 $0 $2,674,980

South Zone

Definition drilling $0 $0 $0 $0 $273,966 $1,086,078 $1,358,123 $1,362,809 $1,986,959 $484,891 $6,552,826Stope development $0 $0 $10,887 $17,857 $2,669,068 $3,586,494 $2,980,196 $11,347,649 $1,561,037 $0 $22,173,187Mining cost $315,226 $315,226 $315,226 $315,226 $3,651,450 $13,638,995 $17,315,679 $16,169,438 $14,599,997 $3,320,486 $69,956,950Integra staff $602,907 $602,907 $602,907 $602,907 $3,383,127 $4,559,156 $4,515,808 $4,682,860 $5,275,475 $1,991,259 $26,819,314Energy cost $165,464 $165,464 $165,464 $165,464 $1,525,491 $1,795,654 $1,856,538 $1,833,263 $1,928,457 $843,667 $10,444,924Milling and transportation $0 $0 $0 $0 $2,086,251 $8,270,484 $10,342,105 $10,377,794 $15,130,696 $3,692,442 $49,899,772Environment $74,704 $74,704 $74,704 $74,704 $374,903 $434,005 $466,799 $484,981 $549,546 $244,420 $2,853,470

$0Capitalized operating cost -$2,482,493 -$2,482,493 -$2,542,191 -$2,654,824 -$36,434,372 -$46,596,373

Total operating costs $0 $0 $0 $0 $0 $68,128,629 $79,018,810 $87,228,253 $64,253,326 $10,638,415 $309,267,433Op. cost/tonne $CAN $0 $170.95 $154 $188 $128 $132 $158Op. cost/oz $CAN $0 $713 $552 $687 $651 $809 $648Op. cost/tonne $US $0.00 $162.81 $146.88 $179.15 $122.02 $125.37 $151Op. cost/oz $US $0 $679 $525 $654 $620 $771 $617

Operating Cash Flow $0 $0 $0 $0 $0 $58,149,960 $110,423,157 $80,776,841 $66,222,159 $6,748,914 $322,321,032

CAPITAL EXPENDITURES

North Zone

Owner's cost $758,715 $758,715 $758,715 $758,715 $5,016,879 $8,051,738Capitalized revenue -$21,577,403 -$21,577,403NSR $1,000,000 $1,000,000Development $565,477 $1,392,748 $2,141,579 $2,110,858 $24,403,985 $10,103,579 $9,306,109 $0 $0 $50,024,333Mobile Equipment $3,786,233 $324,390 $376,640 0 $2,788,987 $2,650,707 $2,283,582 $2,283,582 $1,044,184 $15,538,305Site preparation and installation $1,543,635 $0 $0 0 $1,543,635Buildings $1,528,365 $0 $0 0 $867,767 $528,987 528,987 $528,987 $108,153 $4,091,247Water management and distribution - Environment $1,335,446 $0 $0 0 $1,335,446Ventilation and Air heating $764,363 $0 $0 0 $726,596 $435,264 $365,753 $28,564 $2,320,539Electrical distribution & Communication system $2,025,047 $0 $0 0 $78,540 $433,429 $152,459 $12,930 $2,702,405Mine dewatering $103,506 $0 $0 0 $41,402 $257,941 $239,345 $642,194Compressed Air distribution $197,790 $0 $0 0 $241,441 $107,798 $107,798 $107,798 $51,254 $813,879

South zone

Owner's cost $677,611 $677,611 $677,611 $677,611 $3,758,030 $6,468,474Capitalized revenue -$15,824,821 -$15,824,821NSR $1,000,000 $1,000,000Development $480,690 $480,690 $1,396,449 $2,503,541 $20,143,062 $10,005,450 $9,572,917 $0 $0 $0 $44,582,800Mobile Equipment $3,586,126 $0 $0 0 $3,970,424 $2,511,014 $2,731,811 $2,382,561 $974,907 $483,999 $16,640,840Site preparation and installation $2,133,934 $0 $0 0 $0 $0 $0 $0 $0 $0 $2,133,934Buildings $1,494,984 $0 $0 0 $867,767 $528,987 528,987 $528,987 $108,153 $0 $4,057,865Water management and distribution - Environment $1,275,446 $0 $0 0 $0 $0 $0 $0 $0 $0 $1,275,446Ventilation and Air heating $377,675 $0 $0 0 $1,102,898 $111,498 $378,817 $0 $0 $0 $1,970,888Electrical distribution & Communication system $1,962,392 $0 $0 0 $408,809 $368,117 $337,211 $52,356 $0 $0 $3,128,885Mine dewatering $41,402 $0 $0 0 $264,582 $149,412 $202,314 $223,179 $0 $0 $880,889Compressed Air distribution $197,790 $0 $0 0 $241,441 $107,798 $107,798 $107,798 $51,254 $0 $813,879

$0

$0

Total capital expenditures $24,836,623.21 $3,634,153.44 $5,350,993.26 $6,050,724.94 $27,520,386.46 $30,299,982.35 $26,843,889.94 $6,256,742.24 $2,337,903.38 $483,999.01 $133,615,398

Total Cost cost/oz $CAN $983 $1,031 $739 $736 $675 $846 $876

Total Cost cost/oz $US $936 $982 $704 $701 $643 $806 $834

Salvage $3,512,654 $7,310,539 $10,823,193

Financial guarantee reimbursement $1,200,000 $600,000 $600,000 -$2,400,000

Closure Costs $2,400,000 $2,400,000

Net cashflow -$26,036,623 -$3,634,153 -$5,350,993 -$6,050,725 -$28,120,386 $27,249,977 $83,579,267 $74,520,099 $67,396,910 $13,575,455 $197,128,827

Cumulative cashflow -$26,036,623 -$29,670,777 -$35,021,770 -$41,072,495 -$69,192,881 -$41,942,904 $41,636,363 $116,156,463 $183,553,373 $197,128,827

Estimated Mining and Income taxes -$1,009,589 -$467,064 -$733,398 -$905,617 $317,142 $6,506,239 $23,761,344 $23,638,406 $21,979,101 $191,231 $73,277,795

Cash Surplus After Taxes -$25,027,034 -$3,167,090 -$4,617,595 -$5,145,108 -$28,437,529 $20,743,739 $59,817,924 $50,881,693 $45,417,809 $13,384,223

Cumulative Cash flow After Taxes -$25,027,034 -$28,194,124 -$32,811,719 -$37,956,827 -$66,394,356 -$45,650,617 $14,167,307 $65,049,000 $110,466,808 $123,851,032

Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 TOTALNet cashflow -$41,072,495 -$28,120,386 $27,249,977 $83,579,267 $74,520,099 $67,396,910 $13,575,455 $197,128,827Estimated Mining and Income taxes $0Cash Surplus After Taxes -$37,956,827 -$28,437,529 $20,743,739 $59,817,924 $50,881,693 $45,417,809 $13,384,223 $123,851,032Cumulative Cash flow $0

Pre-tax NPV (5%) $146,006,626

Pre-tax IRR 51%

After-tax NPV (5%) $88,459,063

After-tax IRR 38%

OperationPre production

Year 4 Year 5Year -1

Year 0 Year 1 Year 2 Year 3

www.innovexplo.com


22.2. Sensitivity Analysis

Project risks can be identified in economic and non-economic terms. Key economics were examined by running cash flow sensitivities against:

Operating cost Capital cost Revenue Gold price, exchange rate, mill grade and mill recovery

Sensitivity analyses were performed on the Project’s pre-tax NPV (5%) and IRR, revenue, operating cost, and capital cost (Tables 22.3 and 22.5; Fig. 22.1 and 22.3). While project revenues are directly proportional to gold price, mill recovery and grade, the NPV (5%) and IRR project are highly sensitive to these factors (Tables 22.4 and 22.6; Fig. 22.2 and 22.4). However, NPV (5%) and IRR are moderately sensitive to operating and capital costs.

Table 22.3 – Sensitivity analysis of economical parameters, pre-tax NPV at 5% (millions $)

‐30% ‐20% ‐10%Base Case

scenario10% 20% 30%

Revenue ‐16.4 37.7 91.9 146.0 200.1 254.3 308.4

Opex 220.3 195.5 170.8 146.0 121.2 96.5 71.7

Capex 181.5 169.6 157.8 146.0 134.2 122.4 110.6

Figure 22.1 – Sensitivity analysis of economical parameters, pre-tax NPV at 5% (millions $)

www.innovexplo.com


Table 22.4 – Sensitivity analysis of grade, pre-tax NPV at 5% (millions $)

‐1.5 ‐1.0 ‐0.5

Base Case scenario

0.5 1.0 1.5

Grade 47.3 80.2 113.1 146.0 178.9 211.8 244.7

‐30% ‐20% ‐10% 0% 10% 20% 30%

0

50

100

150

200

250

300

‐1,5 ‐1 ‐0,5 BaseCase

scenario

0,5 1 1,5

NPV 5% (M$)

Grade

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

www.innovexplo.com


Table 22.5 – Sensitivity analysis of economical parameters, pre-tax IRR

‐30% ‐20% ‐10%

Base Case scenario

10% 20% 30%

Revenue ‐1% 18% 36% 51% 66% 80% 94%

Opex 70% 64% 58% 51% 45% 38% 30%

Capex 77% 67% 59% 51% 45% 40% 35%

0

0,2

0,4

0,6

0,8

1

‐0,3 ‐0,2 ‐0,1 Base Casescenario

0,1 0,2 0,3

IRR (%) Rev

Op

Cap

Figure 22.3 – Sensitivity analysis of economical parameters, pre-tax IRR

www.innovexplo.com


Table 22.6 – Sensitivity analysis on grade, pre-tax IRR

‐1.5 ‐1.0 ‐0.5

Base Case scenario

0.5 1.0 1.5

Grade (g/t) 6.69 7.19 7.69 8.19 8.69 9.19 9.69

IRR 22% 32% 42% 51% 61% 69% 78%

0,1

0,2

0,3

0,4

0,5

0,6

0,7

0,8

0,9

IRR (%)

Figure 22.4 – Sensitivity analysis on grade, pre-tax IRR

www.innovexplo.com


23. ADJACENT PROPERTIES

The production, resource and reserve estimates indicated below are of a historical nature and do not comply with NI 43-101. However, the authors believe that this information provides a conceptual indication of the potential of the area, which is pertinent to the Report. The qualified persons have been unable to verify the information and the information is not necessarily indicative of the mineralization on the property that is the subject of the Report. The Sigma-Lamaque Project is located immediately to the north of the Lamaque Property (see Figure 23.1). This project is now controlled by the Deustche Bank (formerly owned by Century Mining and presently non-operating). The resources and reserves on this project include the following: Table 23.1 – Compliant Reserve & Resources – Updated June 2009 Category Metric Tonnes Grade grams/tonne Gold Ounces

Proven & Probable Reserves 7,736,181 4.56 1,134,971

Measured & Indicated Resource 8,305,551 4.81 1,283,891

Inferred Resource 19,680,673 4.95 3,130,779 Information from Century Mining Corporation web site www.centurymining.com – January 14, 2010.

Table 23.2 – Historical production statistics from Sigma mine

Mine Operator Operating Period Production Figures

Tonnes Grade (g/t) Ounce (Au)

Sigma Mines 1937 to 1997 23,898,243 5.8 4,456,477

McWatters Mines

1997 to 2003 372,000 2.2 26,313

Century Mining

2005 1,112,746 1.6 57,242

2006 1,415,530 1.6 72,818

2007 1,155,937 1.5 55,747

2008 46719 3.2 4,807

2010 157,561 2.9 14,691

2011 (to end of April) 112,246 2.2 7,939

Total Production: 4,696,034 Reference: Micon International Ltd., August 2011

Contiguous and south of the Lamaque Property, a group of mining claims called the Airport property (now the Alexis Option), is owned by Aur Resources (optioned by Alexandria Minerals) which hosts auriferous mineralization associated with quartz and quartz-carbonate veins and veinlets. The most significant results found are 3.43 g/t Au over a length of 6.1 m and 6.65 g/t Au over 1.2 m (MRNQ File # GM-43367). Also reported are results of 1.24% Zn and 6.86 g Au/t over 0.76 m (MRNQ File # GM-50650).

www.innovexplo.com


Contiguous with and east of the Lamaque Property boundary, the seven-claim Aumaque property owned by Century is underlain by intermediate to felsic tuffs with some diorite and feldspar porphyry dykes and sills from the Val-d’Or Formation. Mineralization consists of auriferous quartz-pyrite-sphalerite-chalcopyrite-gold-silver veins hosted within an ENE-WSW vertical shear zone. In 1988 an Inferred Resource of 180,894 tonnes grading 8.57 g Au/t was reported (ref: MRNQ Deposit File # 32C04-0063). Since that time, more work has been conducted; however, the resource has not yet been confirmed. To the West is the Goldex Mine owned by Agnico-Eagle reporting 3,500,000 gold ounces in resources and reserves and directly to the west of Goldex there is the Osisko Canadian Malartic open pit gold mine reporting 10.71 million Proven and Probable gold ounces. The adjoining Cadillac, Malartic and Val-d’Or camps have produced over 45 million ounces of gold since the 1930s and presently encompass seven producing gold mines (15 million ounces in reserves) .

www.innovexplo.com


Figure 23.1 – Adjacent Properties

www.innovexplo.com


24. OTHER RELEVANT DATA AND INFORMATION

There is no other relevant information to be included in this Report.

www.innovexplo.com


25. INTERPRETATION AND CONCLUSIONS

The objective of InnovExplo’s assignment was to prepare a Preliminary Economic Assessment (“PEA”) on the potential economic viability of the Lamaque Project based on the Mineral Resource Estimate presented in an earlier report titled “NI 43-101 Technical Report on the Lamaque Property” published in November 2013 by Geologica Groupe-Conseil Inc. and GéoPointCom Inc. InnovExplo considers the present PEA study (and Resource Estimate herein) to be reliable and thorough, based on quality data, reasonable hypotheses, and parameters compliant with Regulation 43-101 and CIM standards regarding mineral resource estimates. Geology The Lamaque Project deposit is at an advanced stage of exploration and hosts significant gold mineralization. The Property is strategically positioned in an area known to be associated with gold mineralization. Geologica considers that the Lamaque Property has excellent potential for further discoveries of economic gold mineralization and that the fundamental control on mineralization is structural and associated with young or late intrusive plugs, dykes and sills of felsic to mafic composition. More detailed knowledge and understanding of the property-scale controls and structures will help guide and focus future drilling programs. Geologica believes that Integra Gold should continue to refine its understanding of the structural complexity to help interpret and define potentially mineralized shear and fault structures, both cutting across and parallel to stratigraphy. The magnetic data should help identify the presence of altered fractured intrusions of felsic to mafic composition. Thereafter, follow-up exploration, surveys, prospecting and drilling should be conducted. Geologica believes that significant exploration potential exists in the Triangle, Mylamaque, Sixteen, No. 6 Vein, No. 4 Plug and No. 5 Plug zones, as well as the No. 3 Mine area. Some definition drilling is warranted to categorize resources and test for lateral and depth extensions of the already recognized mineralized zones. Mineral Resource Estimate

The 2013 Mineral Resource Estimates for the Fortune, Parallel, No. 4 Plug and Triangle zones of the Lamaque Project, presented herein, were completed by GéoPointCom, using all available results as per the effective date of each zone. The main objective was to publish revised mineral resource estimates for the above mentioned zones. The mineral resources presented herein are not mineral reserves as they have no demonstrable economic viability. The result is a Mineral Resource Estimate with Indicated and Inferred resources for each of the four mineralized zones, modelled for underground mining. The resource estimates for the abovementioned zones are part of a NI 43-101 Technical Report prepared and supervised by Geologica Inc. and filed on SEDAR in November 2013. The effective dates of the Mineral Resource Estimates vary from zone to zone but the common publication date by news release is September 25,

www.innovexplo.com


2013. All details for these resource calculations are presented in Tables 14.1, 14.2 and 14.3. The November 2013 mineral resources report for the four zones were included in the current PEA. The 2014 Mineral Resource Estimates for the No. 6 Vein and Sixteen Zone, presented herein, were completed by GéoPointCom, using all available results as per the effective date of each zone. The publication date by press release is January 28, 2014. The result is Mineral Resource Estimates with Indicated and Inferred Resources (see Tables 14.1, 14.2 and 14.3). The mineral resources are not mineral reserves as they have no demonstrable economic viability. These mineral resources were not included in the current PEA. Total Indicated Resource Estimate by zone using a 3.00 g/t Au cut-off (Table 14.1)


No. 4 Plug 1,325,100 5.6 237,450

Fortune Zone 125,500 5.8 23,600

Parallel Zone 793,900 8.2 209,570

Triangle Zone 599,700 9.9 190,670

No. 6 Vein 389,400 6.4 79,550

Sixteen Zone 91,700 5.2 15,440

Total Indicated 3,325,300 7.1 756,280

Total Inferred Resources Estimate by zone using a 3.00 g/t Au cut-off (Table 14.2) Gold Deposit

Name Metric Tonnes Grade (g/t Au) Ounces

No. 4 Plug 0 0.0 0

Fortune Zone 252,300 5.6 45,220

Parallel Zone 153,400 17.5 86,050

Triangle Zone 332,300 12.9 137,600

No. 6 Vein 111,600 6.9 24,590


Total Inferred 851,400 10.8 293,710

www.innovexplo.com


Key parameters for the 2013 and 2014 Mineral Resource Estimates by zone (Table 14.3)


Min. search radius

Max. search radius

Min. no. samples

Max. no. samples

Capping grade

Cut Off grade

Min. true thickness

Compo-site

length

Effective date


10X10X15m 25m 90m 3 10 None 3 2m 0.5m 2012 11/08


5X5X5m 50m 50m 3 15 100


2012 05/24

No. 4 Plug

Ordinary Kriging

10X10X10m 35X50X

16m 60X60X

16m 3 10 300 3 ** 1.0m

2013 03/19


5X5X5m 25m 50m 2 10 80


2013 04/24

No. 6 Vein

Ordinary Kriging

10X10X10m 50X50X

50 100X100

X100 4 8

40 [progressive]

3 2m 1.0m 2012 08/17


10X10X10m 15X15X

15 60X60X

60 5 10 35 3 ** 0.7m

2013 11/18



Metallurgy Three series of laboratory testwork were carried out by ALS Metallurgy Kamloops in 2012 and 2013 using samples from the Lamaque Project (see section 13 for details and report citation). For the purpose of the PEA, WSP recommends using recovery numbers that were obtained from retention times of 48 hrs because the plants that have been examined for the various milling options might offer retention times of about that order of magnitude or likely less. Based on the results obtained from tests KM4025-05 to 08 and after adjusting leaching recoveries for a 48-hour retention time, using the leach kinetic curves, the following recoveries could be used if the only process used is cyanidation:




No. 4 Plug: 82.0% (represents 98.5% of recovery obtained with a retention time of 96 hrs)

The percentage of the 96-hour recovery for the Triangle Zone was lower than the one obtained with the others zones after 48 hrs (96.5% vs. >98%). To validate that, the ALS Kamloops laboratory repeated a cyanidation test with the same Triangle Zone composite and similar test conditions (KM4232).

www.innovexplo.com


The results obtained indicate a gold recovery of 89.5% after a leaching time of 48 hours, which represents 98.5% of the gold recovered after 96 hrs. Still based on the same report, if the process involves gravity separation with cyanidation of the gravity tails, the results obtained in KM4015-01 to 04 (gravity testwork) and KM4025-25 to 28 (cyanidation of gravity tails) were used after calculating leaching recovery for a 48-hour retention time, using the factors above, derived from the leach kinetic curves of tests 05 to 08, to get the following recoveries that could be used for the PEA:


*The average of both tests (reports KM4025 and 4232) was used to calculate the recovery factor for the Triangle Zone (89.6% recovery after 48 hrs which is 97.5% of the 91.9% recovery after 96 hrs). However, the results obtained to date from the series of laboratory tests performed by ALS Metallurgy Kamloops are preliminary and further testwork will be needed to better study gravity concentration, determine the fineness of grind and reagents consumption that will enhance recoveries, and define the best flowsheet that could be used to optimize metallurgical performance as well as economics. A bulk sample run in the selected plant could be extremely useful to adjust the process developed in the lab to a plant-scale process. Environment The federal and provincial environmental acts and regulations were reviewed. An Environmental Impact Assessment (EIA) will be required under the Canadian Environmental Assessment Act (2012) because the proposed production will be greater than 600 tpd. A provincial EIA accompanied by a Bureau d’audiences publiques sur l’environnement (“BAPE”) public hearing may also be required, if production is 2,000 tpd or more. Other specific regulatory approvals will also be required, such as provincial Certificates of Authorization (“CofA”) for the planned water treatment facility, sinking of ramps and shafts, and others. An environmental baseline study was conducted, including the range of physical and biological characteristics of the site. Wetlands and watercourses, as well as terrestrial ecosystems, were described. This data will form a solid basis for the eventual EIA. Key operational elements bearing potential environmental consequences have been presented in the Report, including the storage and disposal of waste rock and the control of water run-off from it, the containment and treatment of mine water, and the overall site water management plan. The requirements of the federal Metal Mining Effluent Regulations and provincial Directive 19 will need to be observed. A mine closure plan (restoration and reclamation) will need to be developed and approved under the conditions of the Québec Mining Act before a mining CofA will

www.innovexplo.com


be issued. The various components of the proposed plan are discussed. The entire cost of the proposed plan must be guaranteed by Integra Gold. The socio-economic setting of the Lamaque Project has been described. The design of the Project is consistent with the zoning bylaws of Val d’Or. No protected areas or areas of archaeological interest are affected by the Project. A stakeholder and First Nations engagement and consultation plan has been prepared and is being implemented by Integra Gold This will eventually include consultations regarding the closure plan. Stakeholder issues raised to date include noise, vibration, traffic circulation, environmental protection and visual impact. Mining Mineralization at the Lamaque Project would be accessed via two separate ramps, or declines, located in the Parallel Zone to the north (the “North Ramp”) and in the Triangle Zone to the south (the “South Ramp”), approximately 2 km apart. Material would then be transported to an off-site mill for toll processing, thereby eliminating the need for the construction and permitting of a new mill and tailings facilities. For the purpose of this PEA, the mining plan for the Lamaque Project calls for a combination of conventional and mechanized mining. Two mining methods are proposed based on the vein geometry of the four deposits: long-hole and room-and-pillar. The approach in this study has been to force the application of long-hole mining where applicable. Waste material generated from drift development will be used to backfill part of the long-hole open stopes. The planned production is 2M tonnes of resources at a diluted grade of 8.19 g/t Au, representing 505,600 ounces of gold recovered. The Lamaque Project has a 4.25-year mine life.

www.innovexplo.com


Financial analysis The present PEA is preliminary in nature. 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 financial analysis indicates a payback period of 1.8 years. The after-tax Net Present Value (NPV) of the project is estimated at $88.5M with a discount rate of 5%. The after-tax Internal Rate of Return (IRR) is evaluated at 38%. The main results are summarized in the following table: Cash flow analysis summary (Table 22.1)

Cash flow analysis summary (Table 22.1)

Parameters Results




Average Annual Gold Production (ounces/year): 112,400




Preproduction Period (years) 2

Mine Life (years) 4.25

Cash Cost per Gold Ounce (CA$/oz) 665

Cash Costs and Sustaining Cost per Gold Ounce (CA$/oz) 805

PRE-TAX




AFTER-TAX



Payback Period (years) 1.8

www.innovexplo.com


25.1. Opportunities & Risks

Opportunities to improve the Lamaque Project economics include the following:

The Issuer is developing a scenario internally which would involve delaying the development of the South Ramp by 12 to 18 months in order to reduce up-front capital cost requirements and utilize cash flow from the North Ramp to fund development of the South Ramp.

Acquisition of a mill instead of toll milling would likely reduce LOM operating costs and allow Integra Gold greater security in meeting its future milling requirements.

The potential to utilize contract mining in order to reduce up-front capital requirements.

Production outlined in the PEA is limited to a vertical depth of 620 m at the Triangle Zone. A 2013 drill program intersected multiple high-grade zones below this level, to vertical depths of up to 1,000 m. The Triangle Zone also remains open to the south, east and west.

Drilling at the Triangle Zone intersected 13.89 g/t Au over 7.0 m, approximately 175 m down-dip from the Triangle Zone resource estimate. The ground in between the Triangle Zone has been subsequently tested in 2014 with assay pending.

The PEA is based on a mineral resource database cut-off date of April 24, 2013 and does not include the subsequent drilling (either infill or expansion) of approximately 39,235 m that was completed in late February 2014. Another phase of drilling (4 rigs) is currently underway, and this is also not included in the PEA.

The PEA does not include resources from the No. 6 Vein or the Sixteen Zone.

Significant mineralization has been identified at the No. 3 Mine and the No. 5 Plug targets. Integra Gold expects to have resource estimates completed for those zones in the second half of 2014. Should a resource be defined at these targets, they could be potentially mined from the North Ramp infrastructure.

Recent metallurgical testwork indicates a potential to further improve gold recoveries.

Risks requiring mitigation strategies include:

The Issuer’s future financial success depends on the ability to raise additional

capital from the issue of shares or the discovery of property which could be economically justifiable to develop. Such development could take years to complete and resulting income, if any, is difficult to determine. The sales value of any mineralization potentially discovered by the Issuer is largely dependent upon factors beyond the Issuer’s control, such as the market value of the products produced.

The resource exploration industry is an inherently risky business with significant capital expenditures and volatile metals markets. The marketability of any minerals discovered may be affected by numerous factors that are beyond the Issuer’s control and which cannot be predicted, such as market fluctuations, mineral markets and processing equipment, and changes to government regulations, including those relating to royalties, allowable

www.innovexplo.com


production, importing and exporting of minerals, and environmental protection.

This industry is intensely competitive and there is no guarantee that, even if commercial quantities are discovered, a profitable market will exist for their sale. The Issuer competes with other junior exploration companies for the acquisition of mineral claims as well for the engagement of qualified contractors. Metal prices have fluctuated widely in recent years, and they are determined in international markets over which the Issuer has no influence.

Exploration and development on the Issuer’s Property are affected by government regulations relating to such matters as environmental protection, health, safety and labour, mining law reform, restrictions on production, price control, tax increases, maintenance of claims, and tenure. There is no assurance that future changes in such regulations would not result in additional expenses and capital expenditures, decreasing availability of capital, increased competition, title risks, and delays in operations.

Management of construction/engineering and procurement schedules, costs, and cost containment.

Operating risks related to recruitment and training and performance of the underground workforce, specifically room-and-pillar miners.

Permitting risks. Crown pillar thickness and stability evaluation through geo-mechanics

characterization and stability analysis. Possibilities that the population does not accept the mining project.

www.innovexplo.com


26. RECOMMENDATIONS

The results from this prefeasibility study demonstrate that the Lamaque Project is technically and economically viable and InnovExplo recommends that Integra Gold continue to advance the project toward prefeasibility. InnovExplo recommends proceeding with the following steps in the continued development of the Lamaque Project.

1. Continue exploration and definition drilling at the Parallel, Triangle and

Fortune zones in 2014 in order to upgrade as many resources as possible from the Inferred resource category to the Indicated resource category, while continuing to increase the resource base laterally and at depth.

2. Update the resource estimates for all zones included in the PEA using all

new information generated since the latest database cut-off and evaluate their impact on Project economics (since April 24, 2013).

3. Complete exploration drilling and perform resource estimations on two of the

Project’s advanced exploration targets, the No. 3 Mine and No. 5 Plug, in order to integrate these areas into the future economic evaluations for the Project.

4. Commence a prefeasibility study that includes:

Hydrogeology study; Rock mass characterization and stope design; Crown pillar stability analysis; Revised mining plan using new resources; Trade-off analysis; Re-scheduling development of the 2 ramps to limit capital

requirement; Energy alternatives for underground air heating; Mineralized material and waste handling alternatives; and Access, possibly via shaft sinking, to the deeper part of the Triangle

and No. 4 Plug zones. Finalize the connecting scenario to the Hydro-Québec grid; Engineering for surface installation, electricity and mechanics

installations; Engineering for water treatment and management facilities; and Updated economic evaluation of capital expenditures and operating

costs.

5. Initiate and complete the permitting process for an underground exploration program. A complete prefeasibility study will likely require underground exploration, meaning a significant portion of Project permitting will be completed during the prefeasibility stage:

Apply for Certificate of Authorization under Québec’s jurisdiction; Apply for Project Description under Canada’s jurisdiction; Mineralization and waste characterization;

www.innovexplo.com


Hydrology study; Noise study; and Biology study.

6. Conduct a fourth phase metallurgical study in order to further improve gold

recoveries for the Triangle and No. 4 Plug zones.

7. Complete a formal information and consultation process in order to promote social acceptability of the Lamaque Project and the plans for its development.

To advance the project, InnovExplo estimates an exploration budget of approximately $6.9M is required as presented in Table 26.1.

Table 26.1 – Proposed work program and budget Item Cost

Exploration and definition drilling at the Parallel, Triangle and Fortune zones

3,500,000$

Update resource estimates 300,000$Complete exploration drilling and perform resource estimations on two of the Lamaque Project’s advanced exploration targets, the No. 3 Mine and No. 5 Plug.

1,000,000$

Prefeasibility study 1,350,000$

Initiate and complete the permitting process for an underground exploration program.

350,000$

Conduct a fourth phase metallurgical study 150,000$Complete a formal information and consultation process

250,000$

Total 6,900,000$

www.innovexplo.com


27. REFERENCES

AMEC 2014: “Étude de Base sur l’Environnement et le Milieu Socio- économique, Project Aurifère Lamaque, Val-d’Or, Québec”. AMEC Environnement & Infrastructure, Janvier, 2014.

Beauregard, A.J., Gaudreault, D. and D’Amours, C. November 2013, NI 43-101

Technical Report on the Lamaque property, for Integra Gold Corp. Beauregard, A.J., Gaudreault, D. and D’Amours, C. June 2011 (Amended Sept.

2012), NI 43-101 Technical Report on the Lamaque property, for Integra Gold Corp.

Bédard, P. 2014. Étude de Faisabilité infrastructure réseau, N/Réf: 121043 -0000-

9M- ER- 0001-_0, prepare par Stavibel, 9 janvier 2014, 11 p. Blecha, Matthew, Aug. 1985, Report on the First Program on the Lamaque property

for the Golden Pond Joint Venture. Report #950T. Teck Explorations Limited.

Blecha, Matthew, 1985, Proposed Exploration of the Villemaque property

Bourlamaque Township. Québec. Report #947T. Teck Explorations Limited. Blecha, Matthew, Jun. 1986, Report on the 1985/86 Exploration of the Lamaque

property Val-d’Or, Québec. Report #959T. Teck Explorations Limited. Blecha, Matthew, Mar. 1986, Teck-Golden Pond Joint Venture Proposed Exploration

and Development at Lamaque. Report #956T. Teck Explorations Limited. Boileau, Pierre, 2004, Induced Polarization Survey executed on the Lamaque Project

on behalf of Kalahari Resources Inc. Burrow, D.R., and E.T. Spooner, 1989, Relationship between Archean Gold Quartz

Vein-Shear Zone Mineralization and Igneous Intrusions in the Val-d’Or and Timmins Areas. Abitibi Subprovince. Canada. Econ Geol. Monograph 6. pp. 424-444.

Christopher, A., 1993, The Lamaque property Exploration Potential for Base Metals

Report No. 1235NB. Teck Exploration Ltd. Corfu, Fernando, 1993, The Evolution of the Southern Abitibi Greenstone Belt in light

of Precise U-Pb Geochronology. Econ. Geol. 88. pp. 1323-1340. Costmine – division of INFOMINE USA INC., 2013 – Mining Cost Service, Volume 1

and Volume 2; Publisher: Jennifer B. Leinart. Davis, Donald W., 2002, U-Pb geochronology of Archean metasedimentary rocks in

the Pontiac and Abitibi subprovinces. Québec. Constraints on timing. provenance and regional tectonics. Precambrian Research 115. pp. 97-117.

www.innovexplo.com


Derry, Michener, Booth & Wahl, 1988, Economic Evaluation of the Lamaque Mine. Val-d’Or, Québec prepared for Tundra Gold Mines Ltd.

Feng, Rui, Jianzhong & Robert Kerrich, 1993, Noble Metal Abundances and

Characteristics of Six Granitic Magma Series. Archean Abitibi Belt. Pontiac Subprovince: Relationship to Metallogeny and Overprinting of Mesothermal Gold Deposits. Econ. Geol. 88. pp. 1376-1401.

Golder Associates Limited, December 2013, PEA Level Review of Geomechanics

and Hydrology – DRAFT, Integra Gold – Lamaque Project, Ref. # 003-12-1221-0123-RA-RevD.

Hansen, Jens E., 1988, Tundra Gold Mines Limited, Lamaque Mine Gold

Exploration. Targets from Processed Ground Magnetic Data. Harris, G., Lewis, W.J., Mukhopadhyay, D. K., Patrick, G. and Lattanzi, C. R., 2011

Technical review of the Mining Plan/Operations and audit of the Resource and Reserve Estimates for the Lamaque Mine Project.

Hester, Brian, 1988, Lamaque Feasibility Study Geology and Reserves. DMBW. Inc. Hugon, Hervè, 1988, Structural Interpretation of the No. 35 Shears and No. 4 Plug

Teck-Tundra Joint Venture property. Val-d’Or, Québec. Karvinen, William., 1985, Geology of the Lamaque Mine property. Val-d’Or, Québec. Lang, B. (1994) Span design for entry type excavations. MASc Thesis, University of

British Columbia. Louvicourt Mining Management Co. Ltd., 1988, Evaluation of the Teck Corporation-

Lamaque property in Val-d’Or. Norman Lecuyer, Madon, Z. C., 1987, Carte Geologique. Drawing No. 6050.

sheets 1 to 6. 1 inch to 200 feet. Corporation Teck-Division Lamaque. Ministere de l’Energie et des Ressources, Government of Québec, 1990 Carte Géologiques des Gites Métalliferes des Districts de Rouyn-Noranda

et de Val-d’Or. Neumayr, P., S.G. Hagemann and J.-F. Couture, 2000, Structural setting, textures

and timing of hydrothermal vein systems in the Val-d’Or camp. Abitibi. Canada: Implications for the evolution of transcrustal. second- and third-order fault zones and gold mineralization. Can. J. of Earth Sci. 37. pp. 95-114.

Patton, T.C., 1988, Summary Report on the Lamaque property Val-d’Or, Québec for

Tundra Gold Mines Limited.

Perrault, Guy, Pierre Trudel and Paul Bedard, 1984, Auriferous Halos Associated with the Gold Deposits at Lamaque mine. Québec. Econ Geol. 79. pp. 227-238.

www.innovexplo.com


Poirier, S. and Tremblay, A., 2013, Review of custom milling options – Lamaque

Property Project, February 18, InnovExplo. Ravenelle, J.-F. and Nagy, C., October 2013, Internal report – No.4 Plug Structural

Geology Study and Three-Dimensional Modelling – SRK Consulting.

Risto, Richard W., Alain J. Beauregard & D. Gaudreault, 2004, Technical report on the Lamaque property, Val-d’Or, Québec for Kalahari Resources Inc. by Watts, Griffis and McOuat Limited.

Robert, Francois and Alex C. Brown, 1984, Progressive Alteration associated with

gold-quartz-tourmaline veins at the Sigma Mine. Abitibi greenstone Belt. Quebec. Econ Geol. 79. pp. 393-399.

Robert, Francois and Alex C. Brown, 1986, Archean Gold Forming Quartz Veins at

the Sigma Mine. Abitibi Greenstone Belt. Québec: Part I. Geologic Relations and Formation of the Vein System. Econ Geol. 81. pp. 578-592.

Robert, Francois and Alex C. Brown, 1986 ,Archean Gold Forming Quartz Veins at

the Sigma Mine. Abitibi Greenstone Belt. Québec: Part II. Vein Paragenesis and Hydrothermal Alteration. Econ Geol. 81. pp.593-616.

Roulston, D, and Shouldice, T., 2013, ALS Metallurgy - Metallurgical test work on the

Lamaque Deposit – Lamaque Project, KM3876, May 29, 2013.

Roulston, D, and Johnston, H., 2013, ALS Metallurgy - Metallurgical test work on the Lamaque Deposit – Lamaque Project, KM3569, April 3, 2013.

Roulston, D, and Johnston, H., 2014, ALS Metallurgy - Metallurgical test work on the

Lamaque Deposit – Lamaque Project, KM4025, January 16, 2014. Roulston, D, and Johnston, H., 2014, ALS Metallurgy - Additional Cyanidation Test

Work on the Lamaque Deposit – Lamaque Project, KM4232, March 5, 2014. Sauvé, Pierre, Laszlo Imreh & Pierre Trudel, 1993 Description des gîtes d’or de la région de Val-d’Or, MM-91-03. Québec. Scammell, D.R., Jan. 1988, Report on the No. 4 Plug Lamaque property. Val-d’Or,

Québec, Report #: 1004T. Teck Explorations Limited. Scammell, D.R., Aug. 1988, Teck-Tundra Joint Venture report on the Geology and

Reserves of the #35 Vein (No. 5 Plug) Bourlamaque Township, Québec. Report 996T. Teck Explorations Limited.

www.innovexplo.com


Scammell, D.R., Oct. 1988, Teck-Tundra Joint Venture Report on the Geology and Reserves of the North Shear and South Shear. No. 5 Plug. Lamaque property. Report #998T. Teck Explorations Limited.

Scammell, D.R., Oct. 1988, Teck-Golden Pond Venture Summary of the Fourth

Program Lamaque property. Report #997T. Teck Explorations Limited.

Scammell, D.R., & S. Frostad., Nov 1988, Proposed Surface Drilling Program Lamaque property Teck-Tundra/Teck-Golden Pond Joint Ventures. Report #999T. Teck Explorations Limited.

Scammell, D.R., Dec. 1988, Report on the Lamaque Main Mine and No. 2 Mine. Val-

d’Or, Québec for Tundra Gold Mines Limited. Scammell & Associates Inc.

Scammell, D.R., Dec. 1988, Report on the Lamaque Main Mine and No. 2 Mine. Val-d’Orm Québec for Tundra Gold Mines Limited. Scammell & Associates Inc.

Scammell, D.R., Mar. 1989, Golden Pond/Teck – Tundra/Teck J.V. report on the

Drilling of the Magnetic Anomalies (Phase I) and other Targets. Report #1009T. Teck Explorations Limited.

Scammell, D.R., May 1989, Report on the No. 4 Plug Lamaque property. Val-d’Or,

Québec. Volume I of II. Report #1012T. Teck Explorations Limited. Scammell, D.R., May 1989, Teck-Tundra J.V. Lamaque-Sigma Drilling Proposal

Lamaque Mine. Val-d’Or, Québec. Report #1017T, Teck Explorations Limited.

Scammell, D.R., May 1989, Preliminary Drill Report on the West Plug Lamaque

property. Val-d’Or, Québec. Report #10118T. Teck Explorations Limited.

Scammell, D.R., Aug. 1989, Tundra/Golden Pond-Teck J.V. Report on the Base Metal Exploration Proposal for the Lamaque property. Report #1024T. Teck Explorations Limited.

Scott, Craig R., Wulf U. Mueller and Pierre Pilote, 2002, Physical volcanology.

stratigraphy. and lithogeochemistry of an Archean volcanic arc: evolution from plume-related volcanism to arc rifting of SE Abitibi Greenstone Belt. Val-d’Or, Canada. Precambrian Research 115. pp. 223-260.

Sutcliff, R.H., C.T. Barrie, D.R. Burrows & G.P. Beakhouse, 1993, Plutonism in the

Southern Abitibi Subprovince: A Tectonic and Petrographic Framework. Econ. Geol. 88. pp. 1359-1375

Taner, Mehmet F. & Pierre Trudel, 1991 Gold Distribution in the Val-d’Or Formation and model for the formation of

the Lamaque-Sigma mines. Val-d’Or, Québec. Can J. Earth Sci. 28. pp. 706-720.

Val-d’Or SAGAX, 1997, Image 2D for Teck Exploration Ltd. Lamaque Project

Bourlamaque Township, Québec 97-N277. (Boileau. Pierre).

www.innovexplo.com


Watts, Griffis and McOuat; “Technical report on the Lamaque property, Val-d’or,

Quebec for Kalahari resources inc”; October, 7 2004; 139P. Williams, B.R., 1988, Lamaque No. 10 Vein Mining No.1 Vein with No. 10 Vein. Teck

Corporation.

Wilson, H.S., 1948, Lamaque Mine in Structural Geology of Canadian Ore Deposits A Symposium. Canadian Institute of Mining and Metallurgy, pp. 882-891.

www.innovexplo.com


APPENDIX I

UNITS, CONVERSION FACTORS, ABBREVIATIONS

www.innovexplo.com


Units

Units in the Report are metric unless otherwise specified. Precious metal content is reported in grams of metal per metric ton (g/t Au or Ag), unless otherwise stated. Tonnage figures are dry metric tons (“tonnes”) unless otherwise stated. Ounces are troy ounces.

Abbreviations

°C degrees Celsius oz troy ounces

ha hectares avdp avoirdupois pound

g grams st short ton

kg kilograms oz/t ounces per short ton

mm millimetres t metric ton (tonne)

cm centimetres Mt millions of metric tonnes

m metres g/t grams per metric ton

km kilometres tpd metric tons per day

masl metres above sea level ppb parts per billion

’ or ft ft ppm parts per million

cfm cubic ft per minute cps counts per second

m3/min cubic metres per minute hp horsepower

Mbs megabytes per second Btu British thermal units

$ or C$ or CAD Canadian dollars kV/kVA kilovolts/kilovolt-amps

US$ or USD American dollars MPa mega pascals

Conversion factors for measurements

Imperial Unit Multiplied by Metric Unit

1 inch 25.4 mm 1 foot 0.305 m 1 acre 0.405 ha

1 ounce (troy) 31.103 g 1 pound (avdp) 0.454 kg

1 ton (short) 0.907 t 1 ounce (troy) / ton (short) 34.286 g/t

TECHNICAL REPORT AND PRELIMINARY ECONOMIC … · Richmont Mines at Francoeur and Beaufor mines,...

Documents

Transcript of TECHNICAL REPORT AND PRELIMINARY ECONOMIC … · Richmont Mines at Francoeur and Beaufor mines,...