NI 43-101 Technical Report Preliminary Economic Assessment of the Whabouchi Lithium Deposit and Hydromet Plant

Prepared for Nemaska Lithium Inc.

Prepared by André Laferrière, P. Geo. SGS Canada Inc.

Yves Dessureault, Eng. Patrice Live, Eng. BBA Inc.

Nicolas Skiadas, Eng. Noël Journeaux, P. Geo., Eng. Journeaux Assoc.

Gary H.K. Pearse, M.Sc., P. Eng. Equapolar Consultants Limited

Ann Lamontagne, Eng. Lamont Inc.

Isabelle Larouche, Eng. Alain Michaud, Eng. Michel Bilodeau, Eng. Céline Charbonneau, Eng. M. Sc.

Effective Date: October 2, 2012 Issue Date: November 16, 2012 Revised Date: February 27, 2013

Equapolar Consultants Limited

Project Number: 2012-014 Nemaska Lithium Inc. NI 43-101 Technical Report Preliminary Economic Assessment

Prepared for

Nemaska Lithium Inc. 450, Gare-du-Palais Street, 1st floor Quebec (Quebec) G1K 3X2 Canada

Prepared by:

Met-Chem Canada Inc. 555, boul. René-Lévesque Ouest, 3e étage Montréal (Québec) H2Z 1B1

February 2013 QPF-009-12/B

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IMPORTANT NOTICE

This Report was prepared as a National Instrument 43-101 Technical Report for Nemaska Lithium Inc. (“Nemaska”) by Met-Chem Canada Inc. (“Met-Chem”). The quality of information, conclusions and estimates contained herein is consistent with the level of effort involved in Met-Chem’s services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this Report. This Report is intended for use by Nemaska subject to the terms and conditions of its contract with Met-Chem. This Report can be filed as a Technical Report with Canadian Securities Regulatory Authorities pursuant to National Instrument 43-101, Standards of Disclosure for Mineral Projects. Except for the purposes legislated under Canadian securities laws, any other uses of this Report by any third party are at that party’s sole risk.

February 2013 QPF-009-12/B

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DATE AND SIGNATURE PAGE – CERTIFICATES

Effective Date: October 2, 2012 Issue Date: November 16, 2012 Revised Date: February 27, 2013

February 2013 QPF-009-12/B

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CERTIFICATE OF AUTHOR

Patrice Live, Eng.

To Accompany the Report entitled:

“NI 43-101 Technical Report Preliminary Economic Assessment of the Whabouchi Lithium Deposit and Hydromet Plant”

Effective Date: October 2, 2012

Issue Date: November 16, 2012

Revised Date: February 27, 2013

I, Patrice Live, Eng., do hereby certify that:

1) I am Manager, Mining with BBA with an office at 630, Rene-Levesque West, Suite 1900, Montreal, Quebec, H3A 4V5; 2) I graduated from Laval University in 1976; 3) I am a registered member of the Order of Engineers of Quebec (#38991); 4) I have worked as a mining engineer continuously since my graduation from university; 5) I have read the definition of “qualified person” set out in the National Instrument 43-101 and certify that, by reason of my education, affiliation with a professional association, and past relevant work experience, I fulfill the requirements to be an independent qualified person for the purposes of NI 43-101; 6) I am responsible for Sections 15.0, 16.1 to 16.6, 16.8 to 16.10, 21.1.3 and 21.3.1 to 21.3.2 of this Technical Report; 7) I have had no prior involvement with the properties that are the subject of the Technical Report; 8) I have visited the site on August 26, 2011; 9) I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this report; 10) Neither I, nor any affiliated entity of mine, is at present under an agreement, arrangement or understanding or expects to become an insider, associate, affiliated entity or employee of Nemaska Lithium Inc., or any associated or affiliated entities; 11) Neither I, nor any affiliated entity of mine, own directly or indirectly, nor expect to receive, any interest in the properties or securities of Nemaska Lithium Inc., or any associated or affiliated companies;

CERTIFICATE OF QUALIFIED PERSON

I, Nicolas Skiadas, Eng., do hereby certify that

1) I am a Project Manager in the consulting firm Journeaux Associates Division of Lab. Journeaux Inc. 801 Bancroft, Pointe-Claire, Quebec, Canada, H9R 4L6; 2) I graduated from McGill University with M. Eng. in 1982. 3) I am a registered member of the Order of Engineers of Quebec (117881). 4) I have worked as a Civil Engineer continuously since my graduation from university B. Eng. in 1977 and M. Eng. in 1982, Civil Engineering and applied mechanics, McGill University. 5) I have read the definition of “qualified person” set out in the National Instrument 43-101 and certify that, by reason of my education, affiliation with a professional association, and past relevant work experience, I fulfill the requirements to be an independent qualified person for the purposes of NI 43-101. 6) I am responsible for the coordination of the complete Technical Report and for the preparation of Sections 16.7, 20.1.5 and 20.5 of the report entitled “NI 43-101 Technical Report Preliminary Economic Assessment of the Whabouchi Lithium Deposit and Hydromet Plant”, effective date October 2, 2012, issue date: November 16, 2012 and revised date: February 27, 2013. 7) I have had no prior involvement with the properties that are the subject of the Technical Report. 8) I have visited the site on July 26 and 27, 2011. 9) I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this report. 10) Neither I, nor any affiliated entity of mine, is at present under an agreement, arrangement or understanding or expects to become an insider, associate, affiliated entity or employee of Nemaska Lithium Inc., or any associated or affiliated entities. 11) Neither I, nor any affiliated entity of mine, own directly or indirectly, nor expect to receive, any interest in the properties or securities of Nemaska Lithium Inc., or any associated or affiliated companies. 12) Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three years from Nemaska Lithium Inc., or any associated or affiliated companies.

13) I have read NI 43-101 and Form 43-101F1 and have prepared the technical report in compliance with NI 43-101 and Form 43-101F1; and have prepared the report in conformity with the generally accepted Canadian Mining Industry practice and, as of the date of the 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.

This Thursday, February 28, 2013, Montreal, Quebec.

Nicolas Skiadas, Eng., P. Eng., M. Eng. Project Manager Journeaux Associates Division of Lab Journeaux Inc. BBA Project Number: 3073002

CERTIFICATE OF QUALIFIED PERSON

I, Noel L. Journeaux, Eng., do hereby certify that:

1) I am President in the consulting firm Journeaux Associates Division of Lab. Journeaux Inc. 801 Bancroft, Pointe-Claire, Quebec, Canada, H9R 4L6; 2) I graduated from Purdue University with M.S.C.E. Civil Engineering in 1962. 3) I am a registered member of the Order of Engineers of Quebec (14341). 4) I have worked as a Civil Engineer continuously since my graduation from Queens University - B.A.Sc. Engineering Geology (1960); Purdue University - M.S.C.E. Civil Engineering (1962). 5) I have read the definition of “qualified person” set out in the National Instrument 43-101 and certify that, by reason of my education, affiliation with a professional association, and past relevant work experience, I fulfill the requirements to be an independent qualified person for the purposes of NI 43-101. 6) I am responsible for the coordination of the complete Technical Report and for the preparation of Section 16.3.1 of the report entitled “NI 43-101 Technical Report Preliminary Economic Assessment of the Whabouchi Lithium Deposit and Hydromet Plant”, effective date October 2, 2012, issue date: November 16, 2012 and revised date: February 27, 2013. 7) I have had no prior involvement with the properties that are the subject of the Technical Report. 8) I have visited the site on August 24, 2011. 9) I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this report. 10) Neither I, nor any affiliated entity of mine, is at present under an agreement, arrangement or understanding or expects to become an insider, associate, affiliated entity or employee of Nemaska Lithium Inc., or any associated or affiliated entities. 11) Neither I, nor any affiliated entity of mine, own directly or indirectly, nor expect to receive, any interest in the properties or securities of Nemaska Lithium Inc., or any associated or affiliated companies. 12) Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three years from Nemaska Lithium Inc., or any associated or affiliated companies. 13) I have read NI 43-101 and Form 43-101F1 and have prepared the technical report in compliance with NI 43-101 and Form 43-101F1; and have prepared the report in conformity with the generally accepted Canadian Mining Industry practice and, as of the date of the 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.

This Thursday, February 28, 2013, Montreal, Quebec.

Noel L. Journeaux, Eng., P. Geo., P. Eng., M.S.C.E., F-A.S.C.E. President Journeaux Associates Division of Lab. Journeaux Inc. BBA Project Number: 3073002

CERTIFICATE OF QUALIFIED PERSON

Gary H. K. Pearse, M. Sc, P. Eng.

To Accompany the Report entitled:

''NI 43-101 Technical Report Preliminary Economic Assessment of the Whabouchi Lithium Deposit and Hydromet Plant"

Effective Date: October 2,2012

Issue Date: November 16,2012

Revised Date: February 27, 2013

I, Gary H.K. Pearse, M. Sc. P. Eng.,

1) Am employed as an Engineer/GeologistlMineral Economist by Equapolar Consultants Ltd. of Unit 101-174 Stanley Ave., Ottawa, ON, Canada KIM IP1; 2) Am the person responsible for the preparation of Sections, 8.0 and 13.2.2 of this study entitled: "N! 43-101 Technical Report Preliminary Economic Assessment of the Whabouchi Lithium Deposit and Hydromet Plant"; 3) Hold the degrees of B. Sc. Geological Engineering CU of Manitoba 1961) and M. Sc. Economic Geology (U of Manitoba 1969); also completed the diploma course in Mineral Economics in the faculty of Mining Engineering (McGill University 1973) and have practiced continuously since; 4) Am in good standing as a member of The Association of Professional Engineers and Geoscientists of the Province of British Columbia, licence #7459 and a member of The Minerals, Metals and Materials Society of the AIME; 5) Have over 45 years experience as a geological engineer, economic geologist and mineral economist, work including geological survey work in Canada and Nigeria, exploration and project development in Canada, USA, Nigeria, Benin, Togo and Tanzania and serving as a government mineral economist; 6) Have for much of my career worked as a rare metals and industrial minerals consultant, in project management, feasibility, metallurgical research and development, and doing market studies based on in-depth knowledge of the mineral-based manufacturing industry; 7) Have undertaken valuations, designed dimension stone quarry and tin-tantalum-pyrochlore pegmatite operations, have calculated reclamation bonds, and published widely on technical and economic topics (over] 00 publications); 8) Have read the definitions of "qualified person" set out in the National Instrument 43-101 and certify that by reason of my education, affiliation with a Professional Association and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of Nl 43-101; 9) Personally inspected the Whabouchi Lithium Deposit property during August 10 and 11, 2010, April 12, 2011 and July 26-27,2011; 10) Am not aware of any material fact or material change with respect to the subject matter of this technical report which is not reflected in the report, which the omission to disclose would make the technical report misleading; 11) Am independent of the issuers applying all of the tests in Sections, 8.0 and 13.2.2 of N! 43-101 Technical Report; 12) Have read National Instrument 43-101 and Form 43-101F1, and the technical report has been prepared in compliance with this instrument and form; 13) I consent to the filing of the Technical Report with any stock exchange or any regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Date: February 28,2013 ,Ottawa, Ontario

The Association of Profes~~~maH~figmee nd Geoscientists of the Province of British Columbia, licence #7459. '

CERTIFICATE OF AUTHOR

To Accompany the Report entitled “NI 43-101 Technical Report Preliminary Economic Assessment of the Whabouchi Lithium Deposit and Hydromet Plant” prepared for Nemaska Lithium Inc. effective as of October 2nd, 2012, issued on November 16th, 2012 and revised on February 27th, 2013.

I, Isabelle Larouche, Eng., do hereby certify that: 1) I am a Mineral Processing Engineer with Met-Chem Canada with an office at suite 300, 555 René-Lévesque Blvd. West, Montréal, Canada; 2) I am a graduate from Laval University with B. Eng. in Materials and Metallurgy Engineering in 2006; 3) I am a registered member of “Ordre des Ingénieurs du Québec” (142262); 4) I have worked as a Mineral Processing Engineer continuously since my graduation from university; 5) I have read the definition of “qualified person” set out in the National Instrument 43-101 and certify that by reason of my education, affiliation with a professional association and past relevant work experience, I fulfil the requirements to be an independent qualified person for the purposes of NI 43-101; 6) I have participated in the preparation of this technical report and part of the sections 13.2 (except 13.2.2), 17.2 and 21.3.8; 7) I have not visited the potential Hydromet Plant project site; 8) I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this report; 9) Neither I, nor any affiliated entity of mine, is at present, under an agreement, arrangement or understanding or expects to become, an insider, associate, affiliated entity or employee of Nemaska Lithium Inc., or any associated or affiliated entities; 10) Neither I, nor any affiliated entity of mine, own, directly or indirectly, nor expect to receive, any interest in the properties or securities of Nemaska Lithium Inc., or any associated or affiliated companies; 11) Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three (3) years from Nemaska Lithium Inc., or any associated or affiliated companies; 12) I have read NI 43-101 and Form 43-101F1 and have prepared the technical report in compliance with NI 43-101 and Form 43-101F1; and have prepared the report in conformity with generally accepted Canadian mining industry practice, and as of the date of the 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.

This 27th day of February 2013.

Isabelle Larouche (signed) ______Isabelle Larouche, Eng. Met-Chem Canada Inc.

CERTIFICATE OF AUTHOR

I, Alain Michaud, Eng., do hereby certify that: 1) I am Manager, Estimation with Met-Chem Canada with an office at suite 300, 555 René-Lévesque Blvd. West, Montréal, Canada; 2) I am a graduate from École Polytechnique de Montréal with B.Eng. in Mechanical Engineering in 1986; 3) I am a registered member of “Ordre des Ingénieurs du Québec” (41788); 4) I have worked as an Estimator in the Mining Industry for the last 8 years; 5) I have read the definition of “qualified person” set out in the National Instrument 43-101 and certify that by reason of my education, affiliation with a professional association and past relevant work experience, I fulfil the requirements to be an independent qualified person for the purposes of NI 43-101; 6) I have participated in the preparation of this technical report for the section 21.2; 7) I have not visited the potential Hydromet project site; 8) I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this report; 9) Neither I, nor any affiliated entity of mine, is at present, under an agreement, arrangement or understanding or expects to become, an insider, associate, affiliated entity or employee of Nemaska Lithium Inc., or any associated or affiliated entities; 10) Neither I, nor any affiliated entity of mine, own, directly or indirectly, nor expect to receive, any interest in the properties or securities of Nemaska Lithium Inc., or any associated or affiliated companies; 11) Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three (3) years from Nemaska Lithium Inc., or any associated or affiliated companies; 12) I have read NI 43-101 and Form 43-101F1 and have prepared the technical report in compliance with NI 43-101 and Form 43-101F1; and have participated in the preparation of the report in conformity with generally accepted Canadian mining industry practice, and as of the date of the 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.

This 27th, day of February 2013.

Alain Michaud (signed) ______Alain Michaud, Eng. Manager, Estimation Met-Chem Canada Inc.

CERTIFICATE OF AUTHOR

I, Michel L. Bilodeau, Eng., do hereby certify that: 1) I am a retired (June 2009) Associate Professor from the Department of Mining and Materials Engineering of McGill University, 3450 University St., Montréal, QC, Canada H3A 2A7, and have taught on a contract basis the mineral economics course of the mining engineering program; 2) I am a graduate of École Polytechnique de Montréal with a B. Eng. in Geological Engineering (1970), and of McGill University with a M. Sc. (App.) in mineral exploration (1972) and a Ph.D. in mineral economics (1975); 3) I am a registered member of “Ordre des Ingénieurs du Québec” (#23799); 4) I have taught continuously in the areas of engineering economy, mineral economics and mining project feasibility studies in the mining engineering program dispenses by McGill University since my graduation from university, and have carried out in the capacity of independent consultant several assignments related to the economic/financial analysis of mining projects; 5) I have read the definition of “qualified person” set out in the National Instrument 43-101 and certify that by reason of my education, affiliation with a professional association and past relevant work experience, I fulfil the requirements to be an independent qualified person for the purposes of NI 43-101; 6) I am responsible for section 22 of this technical report; 7) I have neither visited the mine site nor the hydro-metallurgical plant site; 8) I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this report; 9) Neither I, nor any affiliated entity of mine, is at present, under an agreement, arrangement or understanding or expects to become, an insider, associate, affiliated entity or employee of Nemaska Lithium Inc., or any associated or affiliated entities; 10) Neither I, nor any affiliated entity of mine, own, directly or indirectly, nor expect to receive, any interest in the properties or securities of Nemaska Lithium Inc., or any associated or affiliated companies; 11) Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three (3) years from Nemaska Lithium Inc., or any associated or affiliated companies; 12) I have read NI 43-101 and Form 43-101F1 and have prepared the technical report in compliance with NI 43-101 and Form 43-101F1; and have prepared the report in conformity with generally accepted Canadian mining industry practice, and as of the date of the 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.

This 27th day of February 2013.

Michel Bilodeau (signed) ______Michel L. Bilodeau, Eng. Economic/Financial Analyst Consultant for Met-Chem, Inc. OIQ #23799

CERTIFICATE OF AUTHOR

To Accompany the Report entitled “NI 43-101 Technical Report Preliminary Economic Assessment of the Whabouchi Lithium Deposit and Hydromet Plant” prepared for Nemaska Lithium Inc. effective as of October 2, 2012, issued on November 16, 2012 and revised on February 27, 2013.

I, Céline M. Charbonneau, Eng., do hereby certify that: 1) I am Project Manager with Met-Chem Canada with an office at suite 300, 555 René-Lévesque Blvd. West, Montréal, Canada; 2) I am a graduate from École Polytechnique de Montréal with B.Eng. in Geological Engineering in 1985; 3) I am a registered member of “Ordre des Ingénieurs du Québec” (41764); 4) I have worked as a Geological Engineer or Project Manager continuously since my graduation from university; 5) I have read the definition of “qualified person” set out in the National Instrument 43-101 and certify that by reason of my education, affiliation with a professional association and past relevant work experience, I fulfil the requirements to be an independent qualified person for the purposes of NI 43-101; 6) I have participated in the preparation of this technical report and part of the sections 1.0, 2.0, 18.10, 19.0, 20.7, 21.0, 21.3, 24.3, 25.0 and 26.0; 7) I have visited the potential Hydromet project site; 8) I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this report; 9) Neither I, nor any affiliated entity of mine, is at present, under an agreement, arrangement or understanding or expects to become, an insider, associate, affiliated entity or employee of Nemaska Lithium Inc., or any associated or affiliated entities; 10) Neither I, nor any affiliated entity of mine, own, directly or indirectly, nor expect to receive, any interest in the properties or securities of Nemaska Lithium Inc., or any associated or affiliated companies; 11) Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three (3) years from Nemaska Lithium Inc., or any associated or affiliated companies; 12) I have read NI 43-101 and Form 43-101F1 and have prepared the technical report in compliance with NI 43-101 and Form 43-101F1; and have prepared the report in conformity with generally accepted Canadian mining industry practice, and as of the date of the 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.

This 27th day of February2013.

C. M. Charbonneau (signed) ______Céline M. CHARBONNEAU, Eng. M.Sc, Project Manager Met-Chem Canada Inc.

Nemaska Lithium Inc. NI 43-101 Technical Report Preliminary Economic Assessment Page i

TABLE OF CONTENTS

1.0 SUMMARY...... 1 1.1 General ...... 1 1.2 Geology ...... 1 1.3 Mineral Resource Estimate ...... 1 1.4 In-Pit Mineral Resources Estimate ...... 2 1.5 Mineral Processing and Testing ...... 3 1.6 Mine Method and Planning ...... 6 1.7 Spodumene, Lithium Hydroxide-Monohydrate and Lithium Carbonate Recovery ...... 7 1.8 Project Infrastructure ...... 9 1.9 Market Studies ...... 10 1.10 Environmental Permits ...... 10 1.11 Capital and Operating Costs ...... 10 1.12 Economic Analysis ...... 12 1.13 Conclusions and Recommendations ...... 14 2.0 INTRODUCTION ...... 15 2.1 Scope of Study ...... 15 2.2 Study Contributors ...... 15 2.3 Effective Date and Declaration ...... 18 2.4 Sources of Information ...... 19 2.5 Site Visit ...... 19 2.6 Units and Currency ...... 20 2.7 Acknowledgment ...... 22 3.0 RELIANCE ON OTHER EXPERTS ...... 23 4.0 PROPERTY DESCRIPTION AND LOCATION ...... 24 4.1 Location ...... 24 4.2 Property Ownership and Agreements ...... 25 4.3 Royalties Obligations ...... 28 4.4 Permits and Environmental Liabilities ...... 28 5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...... 29 5.1 Accessibility ...... 29 5.2 Physiography ...... 29 5.3 Climate ...... 29 5.4 Local Resources and Infrastructures ...... 29 5.5 Surface Rights ...... 30 6.0 HISTORY ...... 31 6.1 Regional Government Surveys...... 31 6.2 Mineral Exploration Work ...... 31 7.0 GEOLOGICAL SETTING AND MINERALIZATION ...... 33 7.1 Regional Geology ...... 33 7.2 Property Geology ...... 34 7.3 Mineralization ...... 36 February 2013 QPF-009-12/B

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8.0 DEPOSIT TYPES ...... 37 8.1 Origin and Features of Rare Metal Pegmatites ...... 37 8.2 Stacked Sill Structure ...... 38 8.3 Syntectonic Mobile Zone Feeder Dykes ...... 38 8.4 Mafic Host Rocks ...... 38 8.5 The Whabouchi Pegmatite ...... 39 9.0 EXPLORATION ...... 40 10.0 DRILLING ...... 41 11.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY ...... 44 11.1 Sample Procedure and Sample Security ...... 44 11.2 Sample Preparation and Analysis ...... 45 11.3 Quality Assurance and Quality Control Procedure ...... 46 11.4 Analytical Standards ...... 46 11.5 Analytical Blanks ...... 49 11.6 Core Duplicates ...... 50 11.7 Nemaska Pulp Re-Analysis ...... 51 11.8 Specific Gravity ...... 53 11.9 Conclusion ...... 54 12.0 DATA VERIFICATION ...... 55 13.0 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 58 13.1 Spodumene Concentrate ...... 58 13.2 Hydrometallurgical Process ...... 89 14.0 MINERAL RESOURCE ESTIMATES ...... 96 14.1 Exploratory Data Analysis ...... 97 14.2 Geological Interpretation ...... 100 14.3 Resource Block Modeling ...... 101 14.4 Grade Interpolation Methodology ...... 102 14.5 Mineral Resource Classification ...... 103 14.6 Mineral Resource Estimation ...... 105 14.7 Block Model Validation ...... 108 14.8 Comments about the Mineral Resource Estimate ...... 109 15.0 MINERAL RESERVE ESTIMATE ...... 110 16.0 MINING METHODS ...... 111 16.1 Block Model Validation ...... 111 16.2 Pit Optimization ...... 111 16.3 Engineered Pit Design ...... 113 16.4 Dilution and Mineralized Material Loss ...... 117 16.5 In-Pit Resources ...... 120 16.6 Mine Planning ...... 121 16.7 Waste Rock and Tailings Management ...... 134 16.8 Mine Operations ...... 135 16.9 Fleet Requirements ...... 139 16.10 Mine Manpower Requirements ...... 140 17.0 RECOVERY METHODS ...... 143 17.1 Concentrator Plant ...... 143

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17.2 Hydrometallurgical Processing Plant ...... 156 18.0 PROJECT INFRASTRUCTURE ...... 167 18.1 Whabouchi General Site Plan ...... 167 18.2 Control System ...... 170 18.3 Communication System (Local and External) ...... 171 18.4 Heating, Ventilation and Air Conditioning ...... 173 18.5 Fuel Storage Facilities ...... 173 18.6 Water Supply and Fire Protection ...... 173 18.7 Power Supply and Distribution ...... 174 18.8 Effluent Water Treatment ...... 175 18.9 Camp Accommodations ...... 176 18.10 Valleyfield Plant Infrastructure ...... 177 19.0 MARKET STUDIES AND CONTRACTS ...... 183 19.1 Lithium Hydroxide and Lithium Carbonate ...... 183 19.2 Contract ...... 195 20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT ...... 196 20.1 Environmental Studies ...... 196 20.2 Social Studies ...... 205 20.3 Permitting ...... 208 20.4 Anticipated Impacts ...... 209 20.5 Closure Plan ...... 212 20.6 Conclusion ...... 218 20.7 Hydrometallurgical Plant ...... 218 21.0 CAPITAL AND OPERATING COSTS ...... 220 21.1 Whabouchi Capital Costs ...... 221 21.2 Hydrometallurgical Plant Capital Costs ...... 229 21.3 Operating Costs ...... 235 22.0 ECONOMIC ANALYSIS ...... 258 22.1 General ...... 258 22.2 Assumptions ...... 258 22.3 Financial Model and Results ...... 260 22.4 Sensitivity Analysis ...... 264 23.0 ADJACENT PROPERTIES ...... 268 24.0 OTHER RELEVANT DATA AND INFORMATION ...... 270 24.1 Project Schedule ...... 270 24.2 Whabouchi Construction ...... 272 24.3 Valleyfield Construction ...... 274 25.0 INTERPRETATION AND CONCLUSIONS ...... 275 25.1 Conclusion ...... 275 25.2 Risks Evaluation ...... 275 26.0 RECOMMENDATIONS ...... 277 26.1 For Whabouchi Site ...... 277 26.2 For Hydrometallurgical Plant ...... 277 27.0 REFERENCES ...... 278

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LIST OF TABLES

Table 1.1 – Estimated Mineral Resources (0.4% Li2O Cut-off grade) ...... 2 Table 1.2 – Estimated In-Pit Mineral Resources (0.4% Li2O Cut-off Grade) ...... 3 Table 1.3 – Summary of the Capital Costs Estimate ...... 11 Table 1.4 – Summary of the Operating Costs Estimate ...... 11 Table 1.5 – Summary of the Life of Project Production, Revenues and Costs ...... 12 Table 1.6 – Summary of Financial Indicators ...... 12 Table 2.1 – Study Contributors ...... 16 Table 2.2 – List of Abbreviations ...... 20 Table 4.1 – List of the Property Mineral Titles ...... 26 Table 7.1 – Summary of the Different Lithologies Occurring in the Area ...... 35 Table 10.1 – Drilling Completed by Nemaska at Whabouchi ...... 41 Table 10.2 – Channel Sampling Done by Nemaska at Whabouchi ...... 41 Table 11.1 – Sets Values for the Li-LG and Li-HG Standards ...... 48 Table 11.2 – Summary Statistics of Li-LG and Li-HG Standards ...... 48 Table 11.3 – 2010 Pulps Re-analysis Comparison by Drill Hole Mineralized Intervals ...... 53 Table 11.4 – SGS Geostat SG Measurements Statistical Parameters ...... 53 Table 11.5 – Nemaska SG Measurements Statistical Parameters ...... 54 Table 12.1 – Comparative Statistics for the Check Sampling Results ...... 56 Table 12.2 – Check Sampling Comparison by Drill Hole Mineralized Intervals ...... 57 Table 12.3 – Final Drill Hole Database ...... 57 Table 13.1 – Head Grade of 2010 Metallurgical Sample (Assays Reported In Percent) ...... 60 Table 13.2 – Head Grade Semi-Quantitative X-Ray Diffraction Results ...... 60 Table 13.3 – Semi-Quantitative X-Ray Diffraction Results ...... 62 Table 13.4 – Modal Analysis of -10 Mesh Samples ...... 63 Table 13.5 – Modal Analysis of K80 -212 µm Samples ...... 64 Table 13.6 – Normalized Mass of Li Minerals in K80 -212 µm Samples ...... 65 Table 13.7 – Bulk Modal Analysis of Master Composite Sample ...... 67 Table 13.8 – Normalized Liberation Mass of Li Minerals for Master Composite Sample ...... 68 Table 13.9 – Grindability Test Summary ...... 70 Table 13.10 – Grindability Test Statistics ...... 71 Table 13.11 – WRA of Variability Samples ...... 72 Table 13.12 – Detailed Mass Balance – Mine Representative Sample ...... 84 Table 13.13 – Mine Representative Sample Summary Mass Balances – 2 and 4-Stage Scenarios ...... 85 Table 13.14 – Phase 1 Pilot Plant Feed Analysis ...... 91 Table 13.15 – Membrane Electrolysis Feed Solution Composition ...... 93 Table 14.1 – Range of Li2O Analytical Data for Mineral Resource Estimation ...... 97 Table 14.2 – Statistics for the 2-metre Composites for Li2O ...... 98 Table 14.3 – Resource Block Model Parameters ...... 101 Table 14.4 – Parameters Used for the Whittle Pit Optimization ...... 106 Table 14.5 – Whabouchi Deposit Updated Mineral Resource Estimate ...... 107 Table 14.6 – Comparative Statistics for Assays, Composites and Block Model Datasets ...... 108 Table 16.1 – Recommended Slope Angles for Whabouchi Project ...... 114 Table 16.2 – Summary of Dilution Results ...... 120 Table 16.3 – Final In-Pit Resources...... 121 Table 16.4 – Nemaska Lithium Final Mine Plan ...... 123 Table 16.5 – Drilling Parameters and Assumptions ...... 137 Table 16.6 – Blasting Parameters ...... 138 Table 16.7 – Complete Mining Fleet ...... 141 Table 17.1 – Process Design Basis ...... 143 Table 17.2 – Hydrometallurgical Design Criteria - Summary ...... 157 Table 18.1 – Estimated Total Project Power Demand ...... 174 Table 19.1 – Lithium Demand By Compound – Forecast 2011-2025 ...... 185

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Table 19.2 – Lithium Consumption By Application – Forecast 2011 - 2025 ...... 186 Table 19.3 – Possible Scenarios For Future Lithium Demand ...... 189 Table 19.4 – Lithium Demand Forecast – Sebsitivity Analysis ...... 191 Table 19.5 – Lithium Reserves and Resources (Tonnes Li) ...... 193 Table 20.1 – State of Studies for the Whabouchi Natural Environment Baseline (July 2012) ...... 196 Table 20.2 – “Normal” Operation Conditions Water Flow Estimates ...... 205 Table 20.3 – Total Restoration Costs ...... 216 Table 20.4 – Costs for Restoration of the Accumulation Areas ...... 217 Table 20.5 – Annual Guarantee Payments ...... 217 Table 21.1 – Summary of the Capital Costs Estimate ...... 220 Table 21.2 – Whabouchi Estimated Initial Capital Costs (M$) ...... 221 Table 21.3 – Foreign Exchange Rates ...... 222 Table 21.4 – Direct Cost Currency Split ($ x 1,000) ...... 222 Table 21.5 – Labor Rates Used for Capital Cost Estimation ...... 225 Table 21.6 – Factors Affecting Labour Productivity ...... 226 Table 21.7 – Summary of Capital Cost Estimate ...... 230 Table 21.8 – Summary of the Operating Costs Estimate ...... 236 Table 21.9 – Total Estimated Average Operating Costs for Production of Concentrate ...... 236 Table 21.10 – Annual Mine Equipment Operating Costs (Excluding Fuel Costs) (M$) ...... 238 Table 21.11 – Annual Mine Equipment Operating Costs (Fuel Costs) (M$) ...... 239 Table 21.12 – Blasting Costs by Category ...... 240 Table 21.13 – Complete Personnel Yearly Salaries (M$) ...... 241 Table 21.14 – Summary of Mining Unit Operating Costs by Category ...... 243 Table 21.15 – Summary of Mine Operating Costs (M$) ...... 244 Table 21.16 - Estimated Operating Costs of Coarse Tailings Material ...... 245 Table 21.17 – Process Plant Operating Cost by Sector...... 245 Table 21.18 – Process Plant Manpower ...... 248 Table 21.19 – Process Plant Electrical Power Costs ...... 249 Table 21.20 – Consumables Costs for Crushing ...... 249 Table 21.21 – DMS Consumables Costs ...... 250 Table 21.22 – Grinding Consumables Costs ...... 250 Table 21.23 – Reagent Consumables Costs ...... 251 Table 21.24 – Other Consumables Costs ...... 252 Table 21.25 – G&A Operating Costs by Sector ...... 252 Table 21.26 – G&A Manpower ...... 253 Table 21.27 – Process Plant Operating Costs by Sector ...... 253 Table 21.28 – Camp Operating Costs by Sector ...... 254 Table 21.29 – Average Operating Cost Estimate ($/year) ...... 255 Table 21.30 – Total Personnel Requirement...... 255 Table 21.31 – Average Operating Cost Estimate (CAD$/year) ...... 257 Table 21.32 – Hydrometallurgical Plant Operating Cost Estimate (CAD$/tonne product) ...... 257 Table 22.1 – Macro-Economic Assumptions ...... 258 Table 22.2 – Technical Assumptions ...... 259 Table 22.3 – Project Evaluation Summary – Base Case ...... 260 Table 22.4 – Cash Flow Statement ...... 262

LIST OF FIGURES Figure 1.1 – Sensitivity of Project NPV @ 8% (After Tax) ...... 13 Figure 1.2 – Sensitivity of Project IRR (After Tax) ...... 13 Figure 4.1 – Property General Location ...... 24 Figure 4.2 – Property Location with Near-by Infrastructure ...... 25 Figure 4.3 – Map of the Property Mineral Titles ...... 27 Figure 7.1 – Regional Geology Map ...... 33 Figure 7.2 – Local Geological Map ...... 34

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Figure 7.3 – Map of the Property Geology with Drill Holes Location ...... 36 Figure 10.1 – Plan View of the Drilling at Whabouchi ...... 42 Figure 10.2 – Longitudinal View of the Drilling at Whabouchi ...... 42 Figure 10.3 – Section 400 mE and 900 mE Showing Drill Holes and Envelopes (Looking West) ...... 43 Figure 11.1 – Plots of the Variation of the Li-HG and Li-LG Standards with Time ...... 49 Figure 11.2 – Plot of Variance of Analytical Blanks with Time ...... 50 Figure 11.3 – Correlation Plots for Core Duplicates ...... 51 Figure 11.4 – Correlation Plot of the Pulps Re-analysis for 2010 ...... 52 Figure 12.1 – Correlation Plot of 2010 and 2011 Independent Check Samples ...... 56 Figure 13.1 – Summary of the PEA Study Metallurgical Test Work ...... 58 Figure 13.2 – Summary of Mineralogical Test Work ...... 61 Figure 13.3 – Comparison of Mineral Distributions Between -10 Mesh and K80 -212 µm Samples...... 65 Figure 13.4 – Mineral Release Curves of Li Minerals, Microcline, and Muscovite for Master Composite Sample ... 68 Figure 13.5 – Grade vs. Recovery Curves for Li Minerals of Master Composite Sample ...... 69 Figure 13.6 – Heavy Liquid Test Results – HL1 ...... 73 Figure 13.7 – Test Flow sheet for HL2 ...... 75 Figure 13.8 – Test Flow sheet for HL4 ...... 77 Figure 13.9 – Correlation between Slime Mass Fraction and Lithium Loss ...... 78 Figure 13.10 – Effect of Different Collectors on Spodumene Flotation ...... 80 Figure 13.11 – Locked-Cycle Test Flow sheet – LCT1 ...... 82 Figure 13.12 – DMS Processing Flow sheet – Mine Representative Sample ...... 84 Figure 13.13 – Proposed Process Flow Sheet ...... 88 Figure 14.1 – Histogram of 2-metre Composites for Li2O ...... 98 Figure 14.2 – Plan View Showing the Spatial Distribution of the Composites ...... 99 Figure 14.3 – Longitudinal View Showing the Distribution of the Composites (Looking North) ...... 99 Figure 14.4 – Final 3-D Wireframe Envelopes in Plan View ...... 100 Figure 14.5 – Final 3-D Wireframe Envelopes in Longitudinal View (Looking North) ...... 101 Figure 14.6 – View of the Search Ellipsoids Used for the Different Interpolation Passes ...... 102 Figure 14.7 – Block Model Interpolation Results in Plan View ...... 103 Figure 14.8 – Block Model Interpolation Results in Longitudinal View (Looking North) ...... 103 Figure 14.9 – Block Model Final Classification in Plan View ...... 104 Figure 14.10 – Block Model Final Classification in Longitudinal View (Looking North) ...... 105 Figure 14.11 – Long Section Showing Optimized Pit Outline and Final In-Pit Resource Block Model (Looking North Grid) ...... 108 Figure 16.1 – 2D LG Pit Shell ...... 113 Figure 16.2 – 2D Pit Design ...... 115 Figure 16.3 – 3D Pit Design and LG Pit Shell with Resources (≥ 0.4 % Li2O) ...... 115 Figure 16.4 – Cross-Section View Pit Design, East 437.5 m ...... 116 Figure 16.5 – Cross-Section View Pit Design, East 752.5 m ...... 116 Figure 16.6 – Cross-Section View Pit Design, East 972.5 m ...... 117 Figure 16.7 – Example of Mining Dilution Polygon ...... 118 Figure 16.8 – Plan View of Mining Polygon, Bench 297.5 m ...... 118 Figure 16.9 – Plan View of Mining Polygon, Bench 257.5 m ...... 119 Figure 16.10 – Plan View of Mining Polygon, Bench 212.5 m ...... 119 Figure 16.11 – Plan View of Mining Polygon, Bench 167.5 m ...... 119 Figure 16.12 – Isometric View of 4 Benches Selected to Perform Mining Polygon Method ...... 120 Figure 16.13 – Starter Pit Shown Against Engineered Pit Design ...... 122 Figure 16.14 – Plan View Mine Plan Pre-Production ...... 124 Figure 16.15 – Plan View Mine Plan Year 2 ...... 125 Figure 16.16 – Plan View Mine Plan Year 4 ...... 126 Figure 16.17 – Plan View Mine Plan Year 7 ...... 127 Figure 16.18 – Plan View Mine Plan Year 9 ...... 128 Figure 16.19 – Plan View Mine Plan Year 11 ...... 129 Figure 16.20 – Plan View Mine Plan Year 13 ...... 130 Figure 16.21 – Plan View Mine Plan Year 15 ...... 131

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Figure 16.22 – Plan View Mine Plan Year 17 ...... 132 Figure 16.23 – Plan View Mine Plan Year 18 ...... 133 Figure 16.24 – Waste Rock Piles (Phases A and B) ...... 136 Figure 16.25 – Personnel Trend over LOM ...... 142 Figure 17.1 – PF and Mass Balance ...... 147 Figure 17.2 – PF and Mass Balance at 1.49 % Li2O ...... 148 Figure 17.3 – Water Balance ...... 149 Figure 17.4 – General Site Plan (Scale of 1/7,500) ...... 151 Figure 17.5 – General Site Plant (Scale of 1/1,250) ...... 152 Figure 17.6 – Lithium Hydrometallurgical Plant Simplified Flow sheet ...... 158 Figure 18.1 – Overall General Site Layout and Access Plan ...... 179 Figure 19.1 – Total Lithium Consumption by Application ...... 184 Figure 19.2 - Lithium Demand By Compound (2011) ...... 185 Figure 19.3 – Lithium Hydroxide Demand By Application – 2011 ...... 187 Figure 19.4 – Lithium Hydroxide Demand By Application – 2025 ...... 188 Figure 19.5 – Lithium Carbonate Demand By Application – 2011 ...... 188 Figure 19.6 – Lithium Carbonate Demand By Application – 2025 ...... 189 Figure 19.7 – Lithium Demand – Sensibility Analysis (tonnes LCE) ...... 190 Figure 19.8 – Lithium Chemicals By Source – 2011 ...... 192 Figure 19.9 – Potential Production Capacity From Newcomers ...... 194 Figure 20.1 – Regional and Local Study Area ...... 199 Figure 22.1 – Before-Tax NPV8%: Sensitivity to Capital Expenditure, Operating Cost and Price ...... 265 Figure 22.2 – Before-Tax IRR: Sensitivity to Capital Expenditure, Operating Cost and Price ...... 265 Figure 22.3 – After-Tax NPV8%: Sensitivity to Capital Expenditure, Operating Cost and Price ...... 266 Figure 22.4 – After-Tax IRR: Sensitivity to Capital Expenditure, Operating Cost and Price ...... 267 Figure 23.1 – Location Map Showing Adjacent Mineral Properties ...... 268 Figure 24.1 – Project Schedule ...... 271

LIST OF APPENDICES Appendix A – Drawings

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1.0 SUMMARY

1.1 General

The Whabouchi property is located in the James Bay area of the Province of Quebec, approximately 30 km east of the Nemaska community and 300 km north-northwest of the town of Chibougamau. The Property is accessible by the Route du Nord, the main all- season gravel road linking Chibougamau and Nemaska, and crossing the Property near its center. The Nemiscau airport is 18 km west of the Property.

The Property is composed of one (1) block containing 33 map-designated claims covering a total of 1,716 ha. The claims are 100% owned by Nemaska and were acquired via a purchase agreement with Victor Cantore Group, a purchase agreement with Golden Goose Resources Inc. and directly by map designation. The ten (10) claims acquired from Golden Goose are subject to a 2% NSR royalty and the claims acquired from Victor Cantore and four (4) of the seven (7) claims acquired by map designation are subject to a 3% NSR royalty in favour of Victor Cantore Group.

1.2 Geology

The Whabouchi property is located in the northeast part of the Superior Province of the Canadian Shield craton, in the Lac des Montagnes volcano-sedimentary formation which is principally composed of metasediments and mafic and ultramafic amphibolites. A spodumene-bearing pegmatite swarm occurs in the center of the Property and is composed of a series of sub-parallel and general sub-vertical pegmatites up to 130 m wide in total. The mineralized pegmatite swarm has a generally NE-SW orientation, extends 1.3 km along strike and reaches a depth of more than 300 m below surface. The lithium mineralization occurs mainly in medium to large spodumene crystals (up to 30 cm in size) but petalite also occurs, averaging less than 2% in the deposit. Muscovite also contains minor lithium.

Nemaska completed exploration programs in 2009, 2010 and early 2011 on the Property. A total of 143 surface channels for 944 samples and 115 drill holes for 22,085 m were completed by Nemaska. In addition to drilling, 14 line-km of ground magnetic surveying covering the main mineralized occurrence and 670 line-km of helicopter-borne magnetic surveying covering the Property were completed.

1.3 Mineral Resource Estimate

SGS Geostat completed the mineral resource update using the digital database supplied by Nemaska which included channel data from trenches and drill hole data completed by Nemaska since 2009. The database used to produce the mineral resource estimate was

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derived from a total of 460 channels and diamond drill holes and contained the collar, survey, lithology, and analytical results information.

The mineral resource estimate was derived from a computerized resource block model. Once the modeling was complete, the analytical data contained within the wireframe envelopes is normalized to generate fixed length analytical composites. The composite data was used to interpolate the grade of blocks regularly spaced on a defined grid that fills the 3-D wireframe envelopes. An optimized pit shell model using the pit optimization software Whittle was produced using the completed block model. The interpolated blocks located below the bedrock/overburden interface and within the optimized pit shell comprise the mineral resources.

Table 1.1 shows a summary of the estimated mineral resource for a cut-off grade of 0.4% Li2O.

Table 1.1 – Estimated Mineral Resources (0.4% Li2O Cut-off grade) Li O Grade Category Tonnes 2 (%) Measured 11,294,000 1.58 Indicated 13,785,000 1.50 Measured and Indicated 25,079,000 1.54 Inferred 4,401,000 1.50

1.4 In-Pit Mineral Resources Estimate

BBA prepared the in-pit mineral resources estimate for the PEA Report, based on a conventional open pit operation. The mining engineering work required for the Study, such as the pit optimization, engineered pit design, mine planning, and economic analysis is based on the resource block model prepared by SGS Geostat (“Geostat”) and only used the blocks classified as either a measured or indicated resource in order to generate revenue.

After multiple simulations on in-pit resources estimation, the concentrator cut-off grade was established at 0.4% Li2O. The average grades, using cut-offs ranging from 0.4% Li2O and 0.8% Li2O are very close and therefore the 0.4% Li2O has been selected. Table 1.2 shows a summary of the estimated In-Pit mineral resources for a cut-off grade of 0.4% Li2O.

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Table 1.2 – Estimated In-Pit Mineral Resources (0.4% Li2O Cut-off Grade)

Category Tonnes Li2O Grade (%) Measured 10,197,000 1.530 Indicated 9,442,000 1.455 Total 19,639,000 1.49 Waste Inferred 377,000 Rock 56,646,000 Overburden 2,356,000 Total Stripping 59,379,000 Stripping Ratio 3.02 : 1

1.5 Mineral Processing and Testing

1.5.1 Concentrator Process Testing

Bench scale testing was carried out on core from four (4) NQ holes drilled by Nemaska through the central portion of the main pegmatite in order to obtain approximately one (1) tonne of mineralized core exclusively for metallurgical test work. The core was sent by truck to the SGS Lakefield Laboratories in July of 2010. The head grade of the sample was reported as 1.72% Li2O. The material collected from the pegmatite was 5.5% higher in Li2O than the average grade determined by Geostat and used in their resource estimation. This was considered to be within acceptable limits and appropriate for the metallurgical work for the PEA Study. The following tests were carried out: • Mineralogy; • Crushing and Grinding Tests; • Dense Medium Test Work, and • Bench Scale Flotation Tests. Dense Media Separation (“DMS”) pilot testing was carried out on two (2) separate samples from the Whabouchi deposit. The first being a 25 tonnes blasted outcrop sample and the second was a 5 tonnes mine representative sample consisting of a blend of outcrop and drill core materials. Testing was carried out in several stages consisting of crushing, scrubbing, screening, DMS, magnetic separation, and dewatering. Based on test data during processing of the blasted sample, the mine representative sample was processed using a simplified flow sheet and was considered to approximate the final future DMS plant flow sheet.

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Pilot scale flotation tests were carried out on the as-received blasted outcrop and mine representative samples, as well as the DMS middling products obtained by processing the blasted outcrop and mine representative samples. Based on DMS and flotation pilot testing at a head grade of 1.61% Li2O, the following was concluded: • 25.5% of the total lithium was recovered in the DMS concentrate, while 65.7% of the lithium reported to the flotation circuit; • The flotation Li recovery was 80%, for an overall lithium recovery of 78%;

• The DMS concentrate grade was 6.2% Li2O, while the flotation concentrate grade was 5.9% Li2O. The combined concentrate grade was 6.0% Li2O, and • The DMS concentrate represented 31.5% of the total concentrate and contained 32.7% of the total recovered lithium, while the flotation concentrate represented 68.5% of the total concentrate and contained 67.3% of the total recovered lithium. Sedimentation and filtration tests were carried out at SGS Lakefield. Based on the filtration tests, it was concluded that a pressure filter would be required. The flotation tailings will be combined with the DMS tailings and dry stacked.

1.5.2 Pyrometallurgical Testing

Prior to the hydrometallurgical test work at SGS Lakefield, the Whabouchi spodumene concentrate was roasted at Feeco’s pilot plant facilities in Green Bay, Wisconsin, U.S.A. Fine (flotation) and coarse (DMS) concentrates from the original concentration flow sheet were shipped from SGS Lakefield to Feeco’s laboratory. Two (2) blends of concentrates were prepared using the as-received coarse material and the dried fine material. A first blend was prepared, containing 75% fines and 25% coarse. The second blend contained 50% fines and 50% coarse.

Roasting converts the spodumene’s crystalline structure from alpha to beta. The conversion temperature is about 1,025 ºC. The beta phase is reactive with sulfuric acid and produces lithium sulphate, which is amenable to leaching. A paddle mixer was used to blend the beta spodumene with 93% sulfuric acid.

1.5.3 Hydrometallurgical Testing

The hydrometallurgical testing was done in three (3) phases at SGS Lakefield. Additional test work was completed by Ameridia. a) Hydrometallurgical Testing – Phase 1 The Phase 1 consisted of concentrate leach and all purification steps (primary and secondary impurity removal and ion exchange). The test program was carried out in

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November 2011 by SGS Lakefield. The two (2) blends tested during the pyrometallurgy test work were provided to feed the pilot plant (75/25 and 50/50).

The objectives of the concentrate leach and the Primary Impurity Removal (“PIR”) were to dissolve lithium sulphate and remove the major impurities (Fe, Al, Si, Mn and Mg). Solids were mixed with site water in a 50:50 ratio and agitated for 30 minutes.

The objectives of the Secondary Impurity Removal (“SIR”) were to precipitate Ca, Mg and Mn impurities from the PIR filtrate. Sodium carbonate (Na2CO3) was added in the third tank to convert all the remaining divalent impurities to insoluble carbonates.

The objective of the Ion Exchange (“IX”) circuit is to further reduce the calcium and magnesium tenors from the SIR discharge to 10 mg/L each. The IX circuit consisted of three (3) columns packed with a cationic resin which is selective towards divalent and trivalent metal ions. The process consisted in a lead/lag/regeneration operation. b) Hydrometallurgical Testing – Phase 2 and Phase 3 The Hydrometallurgical (Phase 2) test program was carried out in December 2011 by SGS Lakefield. Phase 2 consisted of Membrane Electrolysis (“ME”) test work and subsequent crystallization to produce lithium hydroxide monohydrate. The objective of the electrolysis process is to produce a lithium hydroxide (LiOH) solution from a high purity lithium sulphate (Li2SO4) solution. The objective of the crystallization process is to produce high quality solid lithium hydroxide monohydrate (LiOH-H2O) from the lithium hydroxide (LiOH) solution generated through membrane electrolysis.

The Hydrometallurgical (Phase 3) test program was carried out in March 2012 by SGS Lakefield. Phase 3 consisted of the production of lithium carbonate. The objective of the test work was to prove that lithium hydroxide conversion to lithium carbonate can be an effective method of producing high purity lithium carbonate. The lithium carbonate (Li2CO3) production process consists of two (2) stages, the first is referred to as Lithium hydroxide (LiOH) Carbonization (“LC”) and the second is called lithium bicarbonate decomposition (“DC”). During carbonization (LC), carbon dioxide gas is reacted with lithium hydroxide at room temperature to form lithium carbonate. Once all lithium hydroxide is carbonized, an excess of carbon dioxide converts lithium carbonate to lithium bicarbonate. During decomposition (DC), the solution is heated to near boiling. Lithium bicarbonate formed in the first stage is decomposed to insoluble lithium carbonate and carbon dioxide.

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1.5.4 Electrodialysis Test Work

Electrodialysis using Bipolar Membranes (“EDBM”) was tested as an alternative method of producing lithium hydroxide. Ameridia was asked to conduct EDBM testing in June 2012.

The objective of the test work was to confirm the suitability of the EDBM technology for Nemaska’s solution, measure current efficiency as well as resulting base and acid concentrations.

A total of eight (8) trials were performed. A base concentration of 2 N (47.8 g LiOH/L) and an acid concentration of 1.5 N (73.5 g H2SO4/L) were achieved at an average current efficiency of 58%. Higher concentrations were achieved but affected the current efficiency, which dropped to below 30%.

Based on the results, EDBM appears to be a good technology for a large scale LiOH production facility.

1.6 Mine Method and Planning

The Study is based on a conventional open-pit operation at 1,095,000 tpy or 3,000 tpd using drilling, blasting, loading and hauling mining methods. The geotechnical parameters used for the pit design were provided by Journeaux Assoc. and are a bench face angle of 75° and inter-ramp angle of 56° and bench heights of 10 m, resulting in 4-m berms. The in-pit haulage roads are 22 m wide in order to accommodate two-way traffic for the 46-tonne trucks. A single-lane 16 m-wide ramp is used for the lower levels of the pit. All in-pit ramps have been restricted to a maximum gradient of 10%. The pit design dimensions are a length of 1,250 m, a width of 320 m and a depth of 190 m. BBA performed calculations, using the block model and the mining polygon method, to determine the dilution and mining recovery for the engineered pit design. The resulting dilution was 4.5% at 0.34% Li2O, and the mineralized material loss was 4.5%.

Mining operations will be conducted 24 hours day, seven (7) days week and 365 days per year. Drilling of mineralized material and waste will be performed by a fleet of two (2) diesel drilling rigs. The mineralized material and waste zones will be drilled with 6½” diameter holes, 5 m spacing, and 5 m burden. Blasting will be executed under contract with an explosive company that will supply blasting materials and technology, and ensure the delivery and storage of explosive products. Blasting will be accomplished using 100% emulsion type explosives with an average density (in the hole) of 1.25 g/cm3.

Mineralized material, overburden and waste production will be accomplished using a fleet of hydraulic shovels with a 6 m3 bucket capacity and 46-tonne capacity haul trucks. The truck fleet size was calculated using a mechanical availability of 88% in the earlier years, transitioning to 83% in the later years and an equipment utilization factor of 95%. February 2013 QPF-009-12/B

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A maximum of two (2) shovels (in year 6) and a maximum of seven (7) trucks (in year 8) will be necessary to support the mine productivity level of 1.095 Mtpy of run of mine mineralization (“ROM”). An additional wheel loader with a 4 m3 bucket will assist in the loading and be also for standby in case of main equipment break down. In addition to this production equipment a number of primary equipment (dozers, graders, etc.) and support equipment (fuel and service trucks, pick-ups, etc.) will be required. All major equipment will be leased.

The fine and coarse mine waste will be mixed with the concentrator tailings to produce a single waste stream, for co-disposal on the waste dumps.

1.7 Spodumene, Lithium Hydroxide-Monohydrate and Lithium Carbonate Recovery

1.7.1 Concentrator

The spodumene concentrator is located near the open pit mine. The concentrator is designed to produce a nominal 213,000 tonnes of spodumene concentrate per year. The ROM mineralized material will be transported to the primary jaw crusher and conveyed to the secondary and tertiary cone crushers. The crushed mineralized material is upgraded in a two-stage dense media circuit to produce a coarse spodumene concentrate and middlings product that is ground and treated by flotation. The DMS middlings product is ground in a rod mill and then combined with screened 0.5 mm material (fine) from the DMS circuit and further ground in a ball mill to P80 200 microns. This ground product forms the feed for the flotation circuit. The flotation circuit consists of desliming and mica pre-flotation followed further desliming and spodumene flotation. Mica concentrate, and spodumene rougher flotation tails are combined and filtered in a pressure filter to a moisture content of about 13% and subsequently combined with DMS tailings and transported by truck to the mine waste stockpile. The spodumene flotation concentrate is filtered by a belt filter to about 5-6% moisture and combined with the DMS concentrate for transport by trucks to Chibougamau. Here it is transferred to a train with 18-24 railcars (90 tonnes each) for transport to CN load out at Valleyfield, some 700 km south of the site near Montreal, Quebec, for further processing; then transport to the hydrometallurgical plant by eight (8) railcars every day.

1.7.2 Hydrometallurgical Plant

The plant availability is estimated at 93%. The hydromet plant is designed to produce a nominal 20,700 tonnes of lithium hydroxide monohydrate crystals per year and a nominal 10,000 tonnes of lithium carbonate powder per year. The overall lithium recovery of the hydromet circuit is 88.6%. The concentrate is discharged from the railcar into a receiving hopper. A reclaim conveyor transports the concentrate to the hydrometallurgical plant where it feeds the spodumene conversion kiln. The kiln heats the spodumene to approximately 1,050°C. The high temperature converts the spodumene concentrate from

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the alpha crystalline structure to the beta crystalline structure. A flash cooler uses ambient air to cool the converted spodumene to approximately 200 °C.

The beta-spodumene is fed to the acid roaster. Sulphuric acid is sprayed onto the beta- spodumene in the roaster and the resulting reaction produces solid lithium sulphate and aluminum silicates. The temperature in the roaster must be maintained above 175 °C.

Lithium sulphate, being soluble in water at these conditions, will dissolve. The slurry discharges from the leach tank into the PIR tank. The discharge of the PIR tanks is pumped to the downstream belt filter. The slurry now contains approximately 19.3 g/L of dissolved lithium. The PIR residue cake, which consists mainly of aluminum silicate, is conveyed outside the plant building where it is stockpiled in anticipation of future marketable possibilities.

In the secondary impurity removal (SIR) tanks, the pH is further increased to precipitate even more dissolved metals as solid hydroxides and carbonates. The discharge slurry is pumped to a centrifuge to remove the solid impurities from the lithium solution. The solid residue is directed to the tailings. The solution is stored in the IX feed tank before it is pumped to the next cleaning process.

The final lithium sulphate solution cleaning step is performed by three (3) ion exchange columns in a round-robin configuration. Solution is fed to two (2) columns in series (the lead column and the lag column) while the third is being cleaned/stripped/regenerated. Clean lithium sulphate solution discharging from the columns is stored in the LiOH feed tank. Waste solutions from the acid strip and wash steps are collected in the IX residue tank and pumped to the tailings tank for disposal.

Membrane Electrodialysis is used to convert lithium sulphate (Li2SO4) to lithium hydroxide (LiOH). In total, 80 electrodialysis cells are required to convert the lithium sulphate solution to lithium hydroxide. A portion of the LiOH solution is directed toward the lithium carbonate production circuit to produce 10,000 t/y of lithium carbonate. The remaining solution is pumped to the LiOH-H2O crystallization circuit.

The LiOH-H2O crystallization circuit consists of a two-stage mechanical vapor recompression falling film evaporator followed by a single-effect steam-driven forced circulation crystallizer. The crystals are separated from the mother liquor in a centrifuge where they reach a moisture content of about 5% (w/w). The dewatered crystals are discharged to a fluid bed dryer before the final packaging step. A flash drying system using indirect heating with natural gas is planned to be installed. The dried crystals are stored in a bin located above the packaging system. A robot operated system is used to package the crystals into one (1) tonne bags.

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The Li2CO3 production process is carried out in two (2) steps. The first step is called the lithium hydroxide carbonization (LC). The second step is called the lithium bicarbonate decomposition step (DC). In the LC step carbon dioxide is injected at the bottom of the tank to react with lithium hydroxide and then with the precipitated lithium carbonate.

In the DC step, the lithium bicarbonate solution is heated to 95ºC. The tanks are jacketed and heated with steam. The DC decanter underflow joins the LC decanter underflow in the belt filter feed tank.

A vacuum belt filter is used to dewater the lithium carbonate slurry to 70 % solids (w/w). The belt filter cake falls by gravity to the flash dryer.

A flash drying system using indirect heating with natural gas is planned to be installed to dry the lithium carbonate final product. The dried product is transferred to the pulverization system. The jet mill uses high pressure air to pulverize the lithium carbonate to a P80 of about five (5) microns. The pulverized product is transferred to a bin located above the packaging system. A robot operated system is used to package the Li2CO3 into 25 kg bags.

1.8 Project Infrastructure

1.8.1 Whabouchi Site

The Route du Nord public road gives access to the existing base camp that will be used for both construction and operations. The road will have to be rerouted after some ten (10) years of mine operation. The camp site is about 12 km north of the mine and concentrator site and the Nemiscau airport is another 7 km further north. The planned infrastructures are: • Mine service and haul roads; • Service buildings that include gate house, administration office and mine management and engineering office; • Maintenance garage and warehouse. In addition to the buildings the following services will be constructed: • Fresh water supply (including fire protection); • Sewage treatment; • Power supply and distribution. 1.8.2 Valleyfield Site

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• Concentrate railcar unloading system; • Electrical substation; • Guard house; • Shipping warehouse, and • Tailings pond, aluminium silicate stock pile and gypsum pond. 1.9 Market Studies

Two (2) market study reports were prepared for Nemaska, one (1) by Roskill Consulting Group Limited (September 2012) and one by SignumBOX Inteligencia de Mercados (March 2012). The largest lithium markets are batteries (29% of overall consumption) and frits and glass (28%). The two (2) main lithium compounds used by the different industries are lithium carbonate and lithium hydroxide. The projected increase in annual demand of these lithium products from 2011 to 2025 is forecasted from 66,736 – 253,739 tonnes LCE and 27,533 – 183,303 tonnes LCE, respectively. The lithium consumption forecast for this period is that lithium hydroxide for batteries will increase from 3% in 2011 to 69% in 2025. Lithium carbonate will also see an increase for that application.

Sensitivity analysis of three pricing scenarios has been done over the same projection periods. Lithium hydroxide is the most sensitive compound to market conditions, because of its use as cathode material in batteries for hybrid and electric cars. The analysis shows a Compound Annual Growth Rate of 12.6% (pessimistic scenario) to 16.8% (optimistic scenario) for lithium hydroxide and 8.7% to 11.7% for lithium carbonate. The current price of lithium carbonate is in the range of US$ 6,000 - US$ 6,500 per tonne and for lithium hydroxide, in the range of US$ 8,500 - US$ 8,900 per tonne.

1.10 Environmental Permits

For the Whabouchi mine project, an Environmental and Social Impact Assessment (“ESIA”), including an Environmental and Social Baseline Study (“ESBS”) is ongoing at the site and should be completed before the end of the year.

For the Valleyfield plant, MDDEP has indicated that this part of the project will need only an Authorisation Certificate (“CA”) and not a complete Environmental and Social Impact Assessment (“ESIA”). The request to obtain this authorisation certificate is in preparation.

1.11 Capital and Operating Costs

1.11.1 Capital Cost Estimate

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sites. Each of the estimates will have a contingency added which is larger (15%) for the Valleyfield site compared with the Whabouchi site (10%) mainly because the hydromet plant estimate is considered to have still a Preliminary Economic Assessment (“PEA”) accuracy. The estimated Owner’s costs are also included. Table 1.3 shows the summary of the capital costs. A Working Capital, equal to two (2) months operating costs, has been included as well.

Table 1.3 – Summary of the Capital Costs Estimate Description Capital Costs CAD $M Whabouchi Site – Mine and Concentrator Total Direct Costs 110.9 Total Indirect Costs (incl. Owner’s Cost) 34.2 Contingencies 14.1 Sub Total Whabouchi Site 159.2 Valleyfield Site – Hydromet Plant Total Direct Costs 203.4 Total Indirect Costs 37.6 Contingencies 36.2 Sub Total Valleyfield Site 277.2 Mine Development Pre-Stripping 2.5 Trust Fund Rehabilitation First Payment 0.9 Working Capital 14.7 Total Capital Cost 454.5 1.11.2 Operating Costs Estimate

The Operating Costs has been estimated for the Whabouchi operations that include the mine and concentrator operations and the tailings and waste disposal. They are estimated based on the average over the life of the project. The mine operating costs include the costs for leasing the major equipment. The truck transport of the concentrate from the site to Chibougamau will be handled by a contractor and the transport to Valleyfield will be by train. The Operating Costs at the Hydromet Plant have also been estimated. A summary of the costs is shown in Table 1.4.

Table 1.4 – Summary of the Operating Costs Estimate

Operating Cost ($/t LiOH-H2O) Operating Cost ($/t Li2CO3) 3,400 3,495

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1.12 Economic Analysis

An economic analysis based on the production and cost parameters of the project has been carried out and the results are shown in Table 1.5. In the analysis, average selling prices of $8,000 for LiOH-H2O and $6,500 for Li2CO3 have been assumed.

Table 1.5 – Summary of the Life of Project Production, Revenues and Costs Description Life of Project Production – Mineralization 19,639,000 t

Life of Project Production – Concentrate @ 6.0% Li2O 3,828,000 t

Life of Project Production – LiOH-H2O product 366,000 t

Life of Project Production – Li2CO3 product 177,000 t Life of Project Revenue 4,083.1 M$ Life of Project Capital Costs 438.8 M$ Life of Project Operating Costs 1,875.2 M$ Life of Project Pre Tax Cash Flow 1,758.4 M$ Life of Project After Tax Cash Flow 1,123.8 M$ The analysis of these estimates shows the Financial Indicators summarized in Table 1.6.

Table 1.6 – Summary of Financial Indicators Description Pre Tax Payback Period (years) 3.9 NPV @ 6% 752.9 M$ NPV @ 8% 567.2 M$ NPV @ 10% 424.1 M$ Internal Rate of Return 23.3% After Tax Payback Period (years) 4.0 NPV @ 6% 455.1 M$ NPV @ 8% 330.5 M$ NPV @ 10% 234.1 M$ Internal Rate of Return 18.9% Figure 1.1 and Figure 1.2 show the sensitivity of the NPV and IRR, respectively, for variations of in Capital Costs, Operating Costs and Selling Price.

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Figure 1.1 – Sensitivity of Project NPV @ 8% (After Tax)

800

600 mil.)

($ 400

200 NPV

T ‐ A 0

‐200 ‐30 ‐20 ‐100 102030 RELATIVE VARIATION (%)

CAPEX OPEX PRICE

Figure 1.2 – Sensitivity of Project IRR (After Tax)

30.0

25.0

20.0 (%)

15.0 IRR

T ‐ A 10.0

5.0

0.0 ‐30 ‐20 ‐100 102030 RELATIVE VARIATION (%)

CAPEX OPEX PRICE

The PEA is preliminary in nature. There is no certainty that the conclusions reached in the PEA will be realized. Mineral resources that are not mineral reserves do not have demonstrated conomic viability.

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1.13 Conclusions and Recommendations

The parameters used in this Preliminary Economic Assessment include developing a 1,095 Mtpy open-pit mine using diesel hydraulic equipment, construction of a concentrator at the mine site (crushing, heavy media, grinding, flotation circuits) with a nominal capacity of 3,000 tpd of mineralized material at 90% availability and construction of a lithium compounds complex production plant at Valleyfield.

BBA has examined the technical and economic aspects of the Whabouchi mine project within the level of precision of a feasibility study and Met-Chem the Hydrometallurgical plant within the level of a Preliminary Economic Assessment. The current report is a Preliminary Economic Assessment (“PEA”) in conformance with the standards required by NI 43-101 and Form 43-101F1.

A computed cash flow analysis was developed based on the technical aspects and on metal price projections made for lithium hydroxide and lithium carbonate derived from two (2) market studies. As it stands, the Whabouchi deposit contains an economic Mineral Resource.

Consequently, Met-Chem and BBA conclude that the Whabouchi project is technically feasible as well as economically viable. The authors of this Technical Report consider the Whabouchi Project to be sufficiently robust to warrant moving it to the feasibility study stage.

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2.0 INTRODUCTION

The Whabouchi Property is located in the James Bay area of the Province of Quebec, approximately 30 km east of the community of Nemaska and 300 km north-northwest of the town of Chibougamau.

This NI 43-101 Report was prepared by BBA inc. (“BBA”) and Met-Chem Canada Inc. (“Met-Chem”) for Nemaska Lithium Inc. (“Nemaska”). The resource estimates were conducted in accordance with NI 43-101 Standards and Disclosure for Mineral Projects, Form 43-101F1 and BBA and Met-Chem standards. A Technical Report Updated Mineral Resources was performed in July 2011 by SGS Geostat (“Geostat”). This 2011 NI 43-101 report is an update of the NI 43-101 Technical Report from July 2010.

2.1 Scope of Study

BBA and Met-Chem have provided complete engineering and integration services for all aspects of the Whabouchi Lithium Preliminary Economic Assessment (“PEA”) Study, including the mine, mill, hydrometallurgical plant, infrastructure, tailings, capital costs, operating costs, and economic analysis.

This Technical Report presents extracts of the updated mineral resource estimate and summarizes the results of the PEA for the Project.

The following document was prepared to provide a Technical Report and PEA of the Whabouchi Lithium mineralization on the Whabouchi Property (“Property”) located in the James Bay area of the Province of Quebec, in compliance with the provisions of National Instrument 43-101 Standards of Disclosure for Mineral Projects.

This Report was prepared by BBA and Met-Chem at the request of Mr. Guy Bourassa, President and CEO of Nemaska Lithium Inc. Nemaska is a Quebec based company trading on the TSX Venture Exchange (TSX-V) under the symbol “NMX”, with its corporate office at: 450, Gare-du-Palais Street, 1st floor Quebec (Quebec) G1K 3X2 Canada

Tel.: 418 704-6038 Fax: 418 614-0627 2.2 Study Contributors

At the request of Nemaska, BBA and Met-Chem have been retained to prepare a NI 43-101 report for the Nemaska PEA Study with the participation of specialized

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consultants. Table 2.1 provides a detailed list of qualified persons as defined in Section 1.5 of NI 43-101 and their respective sections of responsibility.

Table 2.1 – Study Contributors Section Title of Section Qualified Person 1.0 Executive Summary Céline Charbonneau and related QPS 2.0 Introduction Céline Charbonneau and related QPS 3.0 Reliance on Other Experts Yves Dessureault, BBA Inc. 4.0 Property Description and Location André Laferrière, SGS Geostat Accessibility, Climate, Local 5.0 Resources, Infrastructure and André Laferrière, SGS Geostat Physiography 6.0 History André Laferrière, SGS Geostat 7.0 Geological Setting and Mineralization André Laferrière, SGS Geostat Gary H. K. Pearse, Equapolar 8.0 Deposit Types Consultants Limited 9.0 Exploration André Laferrière, SGS Geostat 10.0 Drilling André Laferrière, SGS Geostat Sample Preparation, Analysis and 11.0 André Laferrière, SGS Geostat Security 12.0 Data Verification André Laferrière, SGS Geostat 13.0 and Mineral Processing and Metallurgical Yves Dessureault, BBA Inc. 13.1 Testing – Spodumene Concentrate 13.2 and Mineral Processing and Metallurgical Isabelle Larouche, Met-Chem 13.2.1 Testing – Hydrometallurgical Process Canada Inc. Mineral Processing and Metallurgical Gary H. K. Pearse, Equapolar 13.2.2 Testing – Pyrometallurgical Test Work Consultants Limited Mineral Processing and Metallurgical 13.2.3 to Isabelle Larouche, Met-Chem Testing – Hydrometallurgical Test 13.2.6 Canada Inc. Work – Electrodyalysis Test Work 14.0 Mineral Resource Estimates André Laferrière, SGS Geostat 15.0 Not required for a PEA Mining Methods – Block Model 16.1, 16.2 Validation and Pit Optimization – Patrice Live, BBA Inc. and 16.3 Engineered Pit Design Mining Methods – Pit Slope 16.3.1 Noël Journeaux, Journeaux Assoc. Parameters Mining Methods – Dilution and 16.4 ,16.5 Mineralized Material Loss – In-Pit Patrice Live, BBA Inc. and 16.6 Resources – Mine Planning

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Section Title of Section Qualified Person Mining Methods – Waste Rock and 16.7 Nicolas Skiadas, Journeaux Assoc. Tailings Management Mining Methods – Mine Operations – 16.8 to 16.10 Fleet Requirements – Mine Manpower Patrice Live, BBA Inc. Requirements 17.0 and Recovery Methods – Concentrate Yves Dessureault, BBA Inc. 17.1 Process Design Basis Recovery Methods – Isabelle Larouche, Met-Chem 17.2 Hydrometallurgical Processing Plant Canada Inc. 18.0 to 18.9 Project Infrastructure – Whabouchi Yves Dessureault, BBA Inc. Céline Charbonneau, Met-Chem 18.10 Project Infrastructure – Valleyfield Canada Inc. Céline Charbonneau, Met-Chem 19.0 Market Studies and Contracts Canada Inc. Environmental Studies, Permitting and 20.0 and Social or Community Impact – Ann Lamontagne, Lamont Inc. 20.1 Environmental Studies Environmental Studies, Permitting and 20.1.5 Social or Community Impact –Run- Nicolas Skiadas, Journeaux Assoc. Off and Mine Water Management Environmental Studies, Permitting and Social or Community Impact – Social 20.2 to 20.4 Ann Lamontagne, Lamont Inc. Studies – Permitting – Anticipated Impacts Environmental Studies, Permitting and 20.5 Social or Community Impact – Closure Nicolas Skiadas, Journeaux Assoc. Plan Environmental Studies, Permitting and 20.6 Social or Community Impact – Ann Lamontagne, Lamont Inc. Conclusion Environmental Studies, Permitting and Céline Charbonneau, Met-Chem 20.7 Social or Community Impact – Canada Inc. Hydrometallurgical Plant Céline Charbonneau, Met-Chem 21.0 Capital and Operating Costs Canada Inc. Capital and Operating Costs – 21.1 Yves Dessureault, BBA Inc. Whabouchi Capital Cost Capital and Operating Costs – Alain Michaud, Met-Chem Canada 21.2 Hydrometallurgical Plant Capital Costs Inc. Capital and Operating Costs – Céline Charbonneau, Met-Chem 21.3 Operating Costs Canada Inc.

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Section Title of Section Qualified Person 21.3.1 to Whabouchi Operating Cost Yves Dessureault, BBA Inc 21.3.7 Hydrometallurgical Plant Operating Isabelle Larouche, Met-Chem 21.3.8 Costs Canada Inc. Michel Bilodeau, Met-Chem Canada 22.0 Economic Analysis Inc. 23.0 Adjacent Properties André Laferrière, SGS Geostat 24.0 to 24.2 Other Relevant Data and Information Yves Dessureault, BBA Inc Other Relevant Data and Information – Céline Charbonneau, Met-Chem 24.3 Valleyfield Construction Canada Inc. 25.0 Interpretation and Conclusions Céline Charbonneau and related QPS 26.0 Recommendations Céline Charbonneau and related QPS 27.0 References

2.3 Effective Date and Declaration

This Report is considered effective as of October 2, 2012 and is in support of the Nemaska press release, dated October 2, 2012, entitled “Nemaska Lithium Announces Positive Preliminary Economic Assessment (PEA) for Whabouchi Mine and Lithium Hydroxide/Carbonate Plant”.

The current Report provides an independent Technical Report for the PEA Study of the lithium mineralization of the Nemaska Deposit, in conformance with the standards required by NI 43-101 and Form 43-101F1. The estimate of mineral resources contained in this Report conforms to the CIM Mineral Resource and Mineral Reserve definitions.

BBA and Met-Chem are not insider, associate, or an affiliate of Nemaska and neither BBA and Met-Chem nor any affiliate has acted as advisor to Nemaska, its subsidiaries or its affiliates, in connection with this Project.

It should be understood that the mineral resources presented in this PEA Report are estimates of the size and grade of the deposits based on a number of drillings and samplings and on assumptions and parameters currently available. The level of confidence in the estimates depends upon a number of uncertainties. These uncertainties include, but are not limited to, future changes in metal prices and/or production costs, differences in size and grade and recovery rates from those expected, and changes in Project parameters. In addition, there is no assurance that the Project implementation will be realized.

The comments in this PEA Report reflect BBA and Met-Chem’s best judgment in light of the information available at the time of preparation. BBA and Met-Chem reserve the

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right, but will not be obligated, to revise this PEA Report and conclusions if additional information becomes known to BBA and Met-Chem subsequent to the date of this PEA Report.

2.4 Sources of Information

This Report is based, in part, on internal company technical reports and maps, published government reports, company letters and memoranda, and public information as listed in the “References”, Section 27 of this Report. Several sections from reports authored by other consultants have been directly quoted in this Report, and are so indicated in the appropriate sections.

The following documents detailing the Whabouchi Lithium Project were made available to BBA by Nemaska: • NI 43-101 Technical Report, Mineral Resource Estimation, Whabouchi Lithium Deposit, Nemaska Exploration Inc. – July 14, 2010; • NI 43-101 Technical Report, Updated Mineral Resources, Whabouchi Lithium Project, James Bay, Quebec, Nemaska Exploration Inc. – July 11, 2011; • Request For Proposal for the Pre-Feasibility Study for the Whabouchi Lithium Project – May 2011; • SignumBOX Inteligencia de Mercados. Lithium Minerals Market, Final Report – March 2012; • Roskill Consulting Group Limited, Battery Grade Lithium Hydroxide Market, Report, September 2012. 2.5 Site Visit

Mr. André Laferrière, M.Sc., P. Geo., a qualified person under the terms of NI 43-101, conducted a site visit to the Property on two (2) occasions: from March 10 to 12, 2010 and on May 4 and 5, 2011.

Mr. Gary H.K. Pearse, M. Sc., P. Eng., a qualified person under the terms of NI 43-101, conducted a site visit to the Property on three (3) occasions: August 10 and 11, 2010, on April 12, 2010, and on July 26 and 27, 2011.

Ms. Ann Lamontagne, Eng., a qualified person under the terms of NI 43-101, conducted a site visit to the Property on August 24, 2011 and several other visits during 2011 and 2012.

Mr. Yves Dessureault, Eng., a qualified person under the terms of NI 43-101, conducted a site visit to the Property on August 24, 2011.

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Mr. Patrice Live, Eng., a qualified person under the terms of NI 43-101, conducted a site visit to the Property on August 24, 2011.

Mr. Nicolas Skiadas, Eng., a qualified person under the terms of NI 43-101, conducted a site visit to the Property on July 26 and 27, 2011.

Mr. Noël Journeaux Eng., P. Geo, a qualified person under the terms of NI 43-101, conducted a site visit to the Property on August 24, 2011 and several other visits during 2011 and 2012.

Ms. Céline Charbonneau, Eng. M. Sc., a qualified person under the terms of NI 43-101, conducted a site visit to the potential hydrometallurgical Plant site in Valleyfield on August 16, 2012.

In addition to the site visits, BBA Inc. carried out a study of all relevant parts of the available literature and documented results concerning the Property and held discussions with technical personnel from Nemaska regarding all pertinent aspects of the Property. The reader is referred to these data sources, which are outlined in the “References”, Section 27 of this Report, for further details.

2.6 Units and Currency

In this Report, all currency amounts are Canadian Dollars (“CAD$”) unless otherwise stated, with commodity prices typically expressed in US Dollars (“US$”). Quantities are generally stated in Système international d’unités (“SI”) metric units, the standard Canadian and international practices, including metric tons (tonnes, t) for weight, and kilometres (km) or metres (m) for distance. Abbreviations used in this Report are listed in Table 2.2.

Table 2.2 – List of Abbreviations Symbol Description Symbol Description 2-D Two dimensions IRA Inter-Ramp Angle 3-D Three dimensions kg Kilograms ° Degree km Kilometres °C Degree Celsius kWh Kilowatt- hour % Percent Li Lithium

$ Dollar Li2O Lithium oxide µm Micrometres m Metres AI Abrasion Index m3 Cubic metre

BFA Bench Face Angle Na2O Sodium Oxide cm Centimetre Ta Tantalum

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Symbol Description Symbol Description Tonnes or ha Hectares Metric Tonnes t Canadian Environmental CEAA E East Assessment Agency Centre d’expertise hydrique du Environmental and Social Baseline CEHQ ESBS Québec Study Canadian Institute of Mining, Environmental and Social Impact CIM ESIA Metallurgy and Petroleum Assessment COMEV Evaluation Committee FOB Free on Board COMEX Review Committee of the JBNQA GOH Gross Operating Hours CWI Bond Crushing Work Index HG High Grade RWI Bond Rod Mill Index ID Identification Inverse distance to the power BWI Bond Work Index ID2 square Inductively Coupled Plasma DMS Dense Media Separation ICP-AES Atomic Emission Spectroscopy Inductively Coupled Plasma Mass ICP-MS LG Low Grade Spectroscopy Inductively Coupled Plasma ICP-OES LG-3D Lerchs-Grossman 3D algorithm Optical Emission Spectroscopy Ministère du Développement James Bay and Northern Quebec JBNQA MDDEP durable de l’Environnement et des Agreement Parcs Ministère des Ressources lbs Pounds MRNF Naturelles et de la Faune Mtpy Million of tonnes per year NSR Net Smelter Return MMU Mobile Manufacturing Units OB Overburden N North PEA Preliminary Economic Assessment PF Process Flow Sheet NQ Drill Core Size (47.6 mm) PFS Pre-Feasibility Study HQ Drill Core Size (63.5 mm) PP Preproduction HL Heavy Liquid Quality Assurance / Quality ppm Parts per million QA/QC Control Société de Développement de la ROM Run-of-Mine SDBJ Baie James RQD Rock Quality Designation SG Specific Gravity Système national de référence S South SNRC cartographique (SNRC) du Canada Synthetic Precipitation Leaching scfm Standard cubic feet per minute SPLP Procedure February 2013 QPF-009-12/B

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Symbol Description Symbol Description SR Stripping Ratio VLF Very Low Frequency t/y tonne per year t/d tonne per day t/m3 tonne per cubic metre m metre m3 cubic metre SIR Secondary Impurity Removal PIR Primary Impurity Removal CL Concentrate Leach Toxicity characteristic leaching TCLP W West procedure Table Jamésienne de Concertation TJCM WRA Whole Rock Analysis Method Minière UTM Universal Transverse Mercator XRD X-ray Diffraction 2.7 Acknowledgment

BBA and Met-Chem would like to acknowledge the support provided by Nemaska personnel during this assignment. The collaboration of Lamont Inc., Journeaux Assoc., SGS Geostat, SGS Lakefield, EEM, and TerraGeographical was also greatly appreciated.

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3.0 RELIANCE ON OTHER EXPERTS

BBA and Met-Chem prepared this Report using reports and documents as noted in Section 27 “References” of this Report. The authors wish to make clear that they are qualified persons only in respect to the areas in this Report identified in their “Certificates of Qualified Persons”, submitted with this Report to the Canadian Securities Administrators.

The authors wish to state that an independent verification of land title and tenure was not performed, nor has BBA Inc. and Met-Chem verified the legality of any underlying agreement(s) that may exist concerning the licenses or other agreement(s) between third parties, but has relied on Nemaska to have conducted the proper legal due diligence.

A draft copy of the report has been reviewed for factual errors by Nemaska. Any changes made as a result of these reviews did not involve any alteration to the conclusions made. Hence, the statement and opinions expressed in this document are given in good faith and in the belief that such statements and opinions are neither false nor misleading at the date of this Report.

The individuals listed below, who have contributed to this PEA Study Report, have extensive experience in the mining and metals industry or in a supporting capacity in the industry. They are not considered as QPs for the purpose of this NI 43-101 Report.

Enzo Palumbo and Francois Metallurgy and Recovery BBA Larouche, P. Eng. Methods

Two (2) market studies were prepared, one (1) by SignumBOX and the other by Roskill Consulting Group Ltd. (“Roskill”), independent consultants. Chapter 19 summarizes the key information from these studies about lithium markets. SignumBOX was mandated to perform a market study to evaluate potential target markets for the lithium concentrate. Roskill was mandated to evaluate the battery grade lithium hydroxide market. BBA and Met-Chem have reviewed the content of the market study reports and believe that they provide a reasonable overview of the past and current lithium minerals market as well as projections according to various recognized sources.

Drill core samples collected and prepared by Nemaska were submitted by Nemaska to SGS Minerals Services (Lakefield, Ontario, Canada), which is an accredited laboratory. Although BBA and Met-Chem have reviewed the test work results generated by SGS Lakefield and believes that they are generally accurate, BBA and Met-Chem are relying on SGS Lakefield as an independent expert.

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4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 Location

The Whabouchi Property is located in the James Bay area of the Province of Quebec, approximately 30 km west of the Nemaska community and 300 km north-northwest of the town of Chibougamau. The center of the Property is situated at about UTM 5,725,750 mN, 441,000 mE, NAD83 Zone 18. The Property is accessible by the Route du Nord, the main all season gravel road linking Chibougamau and Nemaska, and crossing the Property near its center. The Nemiscau airport is 18 km west of the Property. Figure 4.1 shows the general location of the Property.

Figure 4.1 – Property General Location

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Figure 4.2 shows the location of the Property with near-by existing infrastructures.

Figure 4.2 – Property Location with Near-by Infrastructure

4.2 Property Ownership and Agreements

The Property is composed of one (1) block containing 33 map-designated claims covering a total of 1,716 ha. The Company owns 100 % interest in the Property. Sixteen (16) claims were acquired from Victor Cantore Group (“Cantore claims”) on September 17, 2009, ten (10) claims were acquired from Golden Goose Resources Inc. (“Golden Goose”) on January 15, 2010 as part of a larger mining titles purchase agreement (594 claims forming the Lac Levac and Lac des Montagnes then properties), and seven (7) claims were acquired by map designation directly by the Company. All the claims are registered in the name of the Company. As of the date of this Report, all 33 claims are in good standing. The Whabouchi deposit is located on the Cantore claims. The expiry dates for the claims range from November 2, 2013 to January 24, 2014. The mining titles are listed in Table 4.1 and shown in Figure 4.3.

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Table 4.1 – List of the Property Mineral Titles Area Registration Expiration Renewals SNRC Title Title Holder (ha) Date Date Done 32O12 53 CDC 101251 2005-11-03 2013-11-02 3 Nemaska Lithium Inc. 32O12 53 CDC 101252 2005-11-03 2013-11-02 3 Nemaska Lithium Inc. 32O12 53 CDC 101253 2005-11-03 2013-11-02 3 Nemaska Lithium Inc. 32O12 53 CDC 2137247 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137248 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137249 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137250 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137251 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137252 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137253 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137254 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137255 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137256 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137257 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137258 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137259 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137260 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137261 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2137262 2007-11-26 2013-11-25 2 Nemaska Lithium Inc. 32O12 53 CDC 2202355 2010-01-21 2014-01-20 1 Nemaska Lithium Inc. 32O12 53 CDC 2202356 2010-01-21 2014-01-20 1 Nemaska Lithium Inc. 32O12 53 CDC 2202357 2010-01-21 2014-01-20 1 Nemaska Lithium Inc. 32O12 53 CDC 2141913 2008-01-24 2014-01-23 2 Nemaska Lithium Inc. 32O12 53 CDC 2141920 2008-01-24 2014-01-23 2 Nemaska Lithium Inc. 32O12 53 CDC 2141921 2008-01-24 2014-01-23 2 Nemaska Lithium Inc. 32O12 53 CDC 2141927 2008-01-24 2014-01-23 2 Nemaska Lithium Inc. 32O12 53 CDC 2141928 2008-01-24 2014-01-23 2 Nemaska Lithium Inc. 32O12 53 CDC 2141933 2008-01-24 2014-01-23 2 Nemaska Lithium Inc. 32O12 53 CDC 2141934 2008-01-24 2014-01-23 2 Nemaska Lithium Inc. 32O12 53 CDC 2203107 2010-01-25 2014-01-24 1 Nemaska Lithium Inc. 32O12 53 CDC 2203108 2010-01-25 2014-01-24 1 Nemaska Lithium Inc. 32O12 53 CDC 2203109 2010-01-25 2014-01-24 1 Nemaska Lithium Inc. 32O12 53 CDC 2203110 2010-01-25 2014-01-24 1 Nemaska Lithium Inc.

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Figure 4.3 – Map of the Property Mineral Titles

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At the date of this Technical Report, Nemaska has paid for the 16 Cantore claims, a total amount of $510,000 in cash and issued 4,000,000 common shares for a total amount of $958,000. Nemaska is committed to pay Cantore Group an additional $500,000 and to issue 500,000 additional common shares if an independent feasibility study confirms the feasibility of production of the property is obtained.

No value was assigned specifically on the ten (10) claims acquired from Golden Goose since they were part of the Lac Levac and Lac des Montagnes purchase agreement relating to the purchase of 594 claims.

4.3 Royalties Obligations

As described in Section 4.2, the Property is subject to two (2) separate agreements. The first concerns the ten (10) claims acquired from Golden Goose, where a 2% NSR is retained by Golden Goose, of which 1% can be repurchased by Nemaska for $1 million within the first three years from the acquisition. The second relates to 20 claims, the 16 claims that were acquired from Cantore Group plus four (4) of the seven (7) claims acquired by map designation. Cantore retains a 3% NSR on these 20 claims, of which 1% can be repurchased by Nemaska for $1 million. The Whabouchi deposit is located on the Cantore claims.

4.4 Permits and Environmental Liabilities

The main permit required to conduct exploration work on the Property is the forest intervention permit delivered by the provincial Ministère des Ressources Naturelles et de la Faune (“MRNF”). A certificate of authorisation from the Ministère du Développement durable de l’Environnement et des Parcs (“MDDEP”) is also necessary to conduct mechanical stripping covering more than 1,000 m3 of overburden. As of the date of this Report, the Company’s management confirmed having valid work permits and authorisations. To the knowledge of the author, there are no environmental liabilities pertaining to the Property.

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5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Accessibility

The Property is easily accessible via the Route du Nord road that crosses the Property near its center. This road is linking the town of Chibougamau, located approximately 300 km to the SSE, and leads to the community of Nemaska and the Route de la Baie- James road.

5.2 Physiography

The Property is characterized by a relatively flat topography with the exception of the local ridge where the more competent pegmatites occur. The elevation above sea level ranges from 275 m at the lowest point on the Property to 325 m at the top of the pegmatite ridge, with an average elevation of 300 m. Lakes and rivers cover approximately 15% of the Property area. The fauna in the area is typical of the taiga environment observed in the region with a mix of black spruce forest and peat moss- covered swamps. A portion of the Property was devastated by forest fires several years ago. There is no permafrost at this latitude and the overburden cover ranges in depth from 0 m near the ridge to 25 m in the south part of the Property.

5.3 Climate

The climate in the region is sub-arctic. This climate zone is characterized by long, cold winters and short, cool summers. Daily average temperature ranges from -20°C in January to +17°C in July. Break-up usually occurs in early June, and freeze-up in early November.

5.4 Local Resources and Infrastructures

The nearest infrastructure with general services is the Relais Routier Nemiscau Camp, located 12 km west of the Property, where the Company has setup its field office and core logging facilities. The community of Nemaska located 30 km west also has accommodation and general services. The area is serviced by the Nemiscau airport, serviced by regular Air Creebec flights and charter flights, and by cellular network from the principal Canadian service providers. There is no mining infrastructure on the Property.

Hydro-Québec owns several infrastructures and facilities in the area including the Poste Albanel and Poste de la Nemiscau electrical stations located approximately 20 km east and 12 km west from the Property, respectively. Electrical transmission lines connecting both stations run alongside the Route du Nord road and cross the Property near its center.

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5.5 Surface Rights

All claims comprising the Property are located on Crown Lands. There is no reason to believe that the Company will not be able to secure the surface rights to construct the infrastructures related to a potential mining operation.

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6.0 HISTORY

6.1 Regional Government Surveys

Numerous geological surveys and geoscientific studies have been conducted by the Québec Government in the James Bay area. Geological surveys in the 1960s (Valiquette 1964, 1965 and 1975) cover the entire property area. In 1998, the MRNF released the results of a regional lake bottom sediment survey completed in 1997.

6.2 Mineral Exploration Work

The first exploration work reported in the area, dating back to 1962 by Canico, included a lithium-bearing pegmatite found by the geologists of the Québec Bureau of Mines. That same year, Canico drilled two (2) packsack drill holes on the pegmatite, followed by three (3) diamond drill holes on the same pegmatite ridge in 1963. A total of 463.11 m were drilled. The best result obtained was 1.44% Li2O over 83.2 m (Elgring 1962).

No exploration is reported for the next ten (10) years. In 1973, James Bay Nickel Ventures (Canex Placer) performed a large-scale geological reconnaissance that covered the property (Burns 1973). From 1974 to 1982, the exploration work was exclusively reported by the Société de Développement de la Baie James (“SDBJ”), which mainly executed large scale geochemical surveys, followed by geological reconnaissance of the anomalies (Pride 1974, Gleeson 1975 and 1976). Two (2) exploration programs, one in 1978 and the other in 1980 were aimed at lithium exploration, with the evaluation of the Whabouchi spodumene-bearing pegmatite (Goyer et al. 1978, Bertrand 1978, Otis 1980, Fortin 1981, and Charbonneau 1982). No channel sampling or drill holes are reported. No work was conducted from 1982 to 1987.

In 1987, Westmin Resources completed an airborne Dighem III survey. A part of this survey was located immediately east of the property (McConnell 1987). In 1987-1988, Muscocho Exploration also completed ground Mag and VLF surveys that covered a major part of the property. The spodumene-bearing pegmatite gave a weak Mag and VLF response. The Muscocho Exploration efforts were oriented toward the search for massive sulphides. A program of 14 holes, 11 of them located on the southern part of the Whabouchi Property, was completed. Several arsenic anomalies were obtained, with a maximum of 3,750 ppm, as in Hole ML-88-8 (Brunelle 1987, Gilliatt 1987 and Zuiderveen 1988).

In 2002, while exploring for tantalum, Inco resampled the spodumene-bearing pegmatite, taking 11 channels and seven (7) grab samples. Inco obtained a best value of 0.026% Ta, and Li2O values ranging from 0.3% to 3.72% (Babineau 2002).

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In 2008, Golden Goose Resources visited and sampled the Valiquette (Ni) and chromite showings south of the Whabouchi Property (Beaupre 2008).

Nemaska, as part of the Qualifying NI 43-101 Technical Report, initiated its exploration work on the Property during the fall of 2009. During the site visit, several outcrops of spodumene-bearing pegmatite were observed and nine samples were collected and analyzed for Li2O. The highest and lowest results obtained during the site visit are the grab sample # 946511, with a value of 6.3% Li2O, and grab sample # 946508 at 1.18% Li2O (Théberge 2009). During the fall of 2009, a mechanical stripping and trenching program was conducted to expose and sample the main spodumene-bearing pegmatite along with a small drilling program designed to validate the historical results.

During 2010 and 2011, exploration work completed by Nemaska on the property included three (3) drilling campaigns, mechanical stripping, ground and airborne geophysics, a 50 tonne bulk sample and metallurgical testing. An initial mineral resource was estimated in May 2010 by SGS Geostat and was followed by an initial preliminary economic assessment of the project completed in March 2011 by Equapolar in collaboration with BBA.

The initial mineral resource estimate of the Whabouchi property, effective May 28, 2010, totaled 9.78 Mt grading 1.63% Li2O in the measured and indicated resources categories, with an additional 15.40 Mt grading 1.57% Li2O in the inferred resources category.

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7.0 GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology

The Whabouchi Property is located in the northeast part of the Superior Province of the Canadian Shield craton. The Superior Province extends from Manitoba to Quebec, and is mainly made up of Archean-age rocks. The general metamorphism is at the greenschist facies, except in the vicinity of intrusive bodies, where it reaches the amphibolite-to- granulite facies. In Quebec, the eastern extremity of the Superior Province has been classified into the following sub-provinces, from south to north: Pontiac, Abitibi, Opatica, Nemiscau, Opinaca, La Grande, Ashuanipi, Bienville and Minto (Hocq 1994). According to Card and Ciesielski (1986), the area covered by the Property is located in the Opinaca or Nemiscau sub- province. Figure 7.1 shows the position of the Property in the Superior Province.

Figure 7.1 – Regional Geology Map

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7.2 Property Geology

The Whabouchi Property is located in the Lac des Montagnes volcano-sedimentary formation and sits between the Champion Lake granotoïds and orthogneiss and the Opatica NE, which is made of orthogneiss and undifferentiated granitoïds. From the northwest to the southeast, the Property is underlain by the Champion Lake granitoïds, a grey oligoclase gneiss and then by the Lac des Montagnes formation. The Lac des Montagnes belt is approximately 7 km wide in the area, oriented northeast, and is principally composed of metasediments (quartz-rich paragneiss, biotite-sillimanite- staurotide schist and garnet-bearing schist) and amphibolites (mafic and ultramafic metavolcanics). These rocks are strongly deformed and cut by late granitoïds (leucogranites and biotite-bearing white pegmatites) (Valiquette 1975). Figure 7.2 shows the location of the Property relative to the Lac des Montagnes, the Champion Lake and Opatica NE formations. Table 7.1 summarizes the different lithologies occurring in the area.

Figure 7.2 – Local Geological Map

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Table 7.1 – Summary of the Different Lithologies Occurring in the Area Pleistocene and Moraines, eskers, alluvial deposits, reticulated peat bogs, morainic belts Holocene 11: Diabase 10: Pegmatites a) White with muscovite, tourmaline, garnet and magnetite b) Pink, with microcline 9: White and pink granite 8: Grey hornblende-oligoclase granite with phenocrist of pink microcline 7: Ultramafic rocks: Serpentinites, tremolite rocks 6: Hornblende-plagioclase gneiss 5: Metasomatic anthophyllite-cordierite rocks (mineralization susceptible) 4: Paragneiss or biotite schists; garnet-biotite schists; porphyroblastic schist: Garnet, sillimanite, biotite

PRECAMBRIAN Garnet, cordierite, biotite Garnet, andalousite, biotite Staurotide, sillimanite, andalousite, biotite Sillimanite, cordierite, andalousite, biotite Amphibole paragneiss 3: Quartz-rich paragneiss; sillimanite, sericite and quartz schist; impure quartzite 2: Pillowed metavolcanic amphibolites 1: Oligoclase gneiss

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Figure 7.3 – Map of the Property Geology with Drill Holes Location

7.3 Mineralization

The mineralization of economic interest observed in the Whabouchi pegmatites is principally lithium in spodumene, which when pure, contains 8% Li2O, with minor nobium and tantalum. Spodumene generally assays 7.6% – 8.0% Li2O depending on the degree of replacement largely by Na2O. Albite and microcline feldspars and high purity quartz generally occurring in pegmatites, are of interest for glass, ceramics and other use when sufficiently close to markets. Rubidium is also present in microcline (feldspar) and muscovite (mica) but this element is normally extracted from lepidolite, lithium-rubidium mica. In the present Study, only spodumene is being considered since metallurgical testing has been conducted only on this element.

Two (2) distinct phases are observed in the Whabouchi pegmatites: a spodumene-bearing phase comprising most of the pegmatite material and a lesser, white to pink barren quartz-feldspar pegmatite. The lithium mineralization occurs mainly in medium to large spodumene crystals (up to 30 cm in size) but petalite also occurs, averaging less than 2% in the deposit. Muscovite also contains minor lithium. The lithium mineralisation sampled from recent drill holes averages 1.62% Li2O and ranges up to 4.24% Li2O.

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8.0 DEPOSIT TYPES

This section is an extract of the 2011 Technical Report.

8.1 Origin and Features of Rare Metal Pegmatites

The interpretation of the pegmatite model is, in a major way that of the author, Gary H. K. Pearse, based on geological mapping, evaluation work, and development work on a number of major pegmatite deposits over many years.

The Whabouchi deposit is a lithium-bearing rare metal pegmatite. Emplacement of rare metal pegmatites is the last phase of the crystallization of a parent granite pluton. High- pressure residual fluids, with abundant water, silica, alumina, alkalis, and rich in rare elements and other volatiles from the crystallization of a pluton at modest depth, concentrate in the cupola or upper domed contact of the granite as it crystallizes. Under increasing pressure, this fluid dilates fractures in overlying rocks in a manner analogous to that of hydraulics in mechanical equipment, thereby providing feeder channels for emplacement of pegmatites. Progressive crystallization of the main rock-forming minerals out of the fluid enriches the final fluids in rare metals and the process culminates in the formation of rare metal pegmatites still under fluid pressure. A variety of types occur depending on the abundance and type of rare metals associated with the pluton and the physico-chemical conditions affecting the sequence of emplacement events.

Pegmatite petrologists classify the variety of types and subtypes by combinations of the following criteria: • Mineralogical-geochemical signatures; • Internal structure/zonation; • Pressure-temperature conditions of crystallization. The criteria are related through degree of fractionation, which arise from the chemical, temperature and pressure evolution of the pegmatite fluids over time and distance from the parent granite. The complex rare element pegmatites generally evolve as follows: at depth under high-pressure and temperature conditions, simple granite pegmatites of quartz, feldspar and mica crystallize in fractures above and within the solidified granite pluton. Above this level, columbo-tantalite minerals appear starting with high niobium compositions and progress to higher tantalum/niobium ratios where the complex pegmatites appear with lithium, cesium, and rubidium bearing minerals. Variations may appear, in which petalite is the dominant lithium mineral, often along pollucite, lepidolite, etc. Alternatively, spodumene dominates in a classification known as Albite-Spodumene pegmatite. Tantalum may occur in a variety of minerals and cassiterite may be present. A

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final, mariolitic or greisen phase at low pressure-temperature, may be present with lepidolite, quartz, tantalum-rich minerals, tin, topaz, etc.

Three (3) characteristics of the geological setting for rare metal pegmatites are common: • Emplacement in concordant stacked sills; • Presence of a compressed, near-vertical, syntectonic mobile zone that is the locus of pegmatite intrusion; • Host rocks most commonly are dominantly mafic volcanics often with intercalated metasediments and gabbroic rocks. 8.2 Stacked Sill Structure

Although a field of rare metal pegmatites is commonly termed a dyke swarm, the major economic ones are most commonly emplaced in concordant, shallow to medium dipping sills with one or more steeply dipping feeder dykes. The mechanism for emplacement of the rare metal pegmatite sills is as follows: stratification in volcanic-metasedimentary- gabbroic piles provides planes of weakness along contacts that facilitate entry of, and hydraulic dilation by, late-stage pressurized rare metal bearing fluid. The layering also provides a barrier or cap to the escape of the volatile fluids from which the final rare metal pegmatites crystallize. Zoning in the pegmatite results from a continuation of crystallization of the rock forming minerals from the cooler contacts inward in the dilated space-albite at the contact, dominantly K-feldspar with quartz-mica next, spodumene quartz with some feldpars and mica, and finally, a core of quartz (in the albite-spodumene type). This simple zoning is often made more complex by two (2) or more pulses of intrusion, albitisation and other replacement reactions.

8.3 Syntectonic Mobile Zone Feeder Dykes

The feeder dykes are near vertical and represent the conduit from depth connecting the pluton to the final rare metal pegmatite bodies. In most cases, shearing at the contacts of the dyke and mylonitisation and/or plastic deformation of the feeder pegmatite identifies this as a deeper, syntectonic mobile zone. In extreme cases, the feeder pegmatite may be stretched and result in the formation of boudinage structure (as occurs at the Moblan Lake “Southwest Dyke” in northern Quebec). The feeder dykes tend to be intermediate in composition in the fractionation chain.

8.4 Mafic Host Rocks

Virtually all pegmatite researchers make only passing reference or none at all to the host rocks of rare metal pegmatites. Their interest has been limited to contact geochemistry and mineralogy and to some extent, the structural setting. The fact that most often the host rocks are volcanic-metasedimentary-gabbroic piles indicates that these rocks are an

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important part of the genesis of rare metal pegmatite fields. Gary H. K. Pearse submits indeed that the presence of these host rocks may be the most important factor after parent granite composition in the genesis of rare metal pegmatites. The layering and ductility, particularly of mafic volcanics and gabbros, causes the pile to flex and stretch without fracturing, thereby confining the high pressure volatiles to permit final crystallization of the rare metal pegmatite. This mechanical behaviour is also conducive to selective emplacement along contact zones between units of the host rock, which accounts for the preponderance of sill-controlled pegmatite deposits. Were the host to be brittle and isotropic, the rocks would tend to fracture, the confinement required to permit coarse crystallization would be absent or greatly reduced, the volatiles would be essentially lost and the final product may be a more uniform feldspar-mica-spodumene rock with more subtle zoning or little zonation at all (examples are: the Thompson Brothers deposit at Wekusko Lake, Manitoba and the Quebec Lithium deposit near Amos, Quebec). In cases where the active feeder dyke propagates through brittle fractures quickly to the surface, the final phase would be an extrusive rhyolite.

8.5 The Whabouchi Pegmatite

The Whabouchi pegmatite is a highly fractionated, spodumene-rich pegmatite swarm, individual bodies of which display typical zoning to varying degrees - a comparatively thin albite wall zone at the contacts followed by a K-feldspar rich zone with lesser albite, quartz, mica, and little or no spodumene, followed by a spodumene-quartz-rich core zone (with variable feldspars and mica) making up more than 90% of the cross-section. The Whabouchi deposit lacks a quartz core which is one of the classic zoned pegmatite features. Insufficient stratigraphic work has been done on the host rocks to establish that the bodies are dominantly sills as in the classic case. The concordance of the bodies with the greenstone belt and the persistence of even thin pegmatite bodies over a 100 m or more on strike and at depth support this structural control. The drilled sections at 700E and 800E on the grid do appear to show this, in that the hanging wall of the main pegmatite zone is basalt and the footwall gabbro.

The south contact observed at several locations by the author in the field showed strong shearing and stretched basalt pillows indicating that the main pegmatite was emplaced in the mobile feeder zone, rather than in offshoot sills from the feeder. This suggests that the feeder dyke was controlled by the steeply dipping volcano-sedimentary-gabbro pile and that above the present level, an unknown layer or structure, now eroded away, sealed the system and allowed containment of the volatile pegmatite fluids for crystallization in the steeply dipping dilation.

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9.0 EXPLORATION

This section is an extract of the 2011 Technical Report.

The Company began working on the Property in October of 2009 with a first exploration program that lasted 25 days. During fall 2009 exploration program, mechanical stripping successfully exposed the spodumene-bearing pegmatites in 16 trenches spaced between 50 and 100 m apart and covering 1,000 m in strike length. From these trenches, 35 channels were cut and a total of 295 samples were collected for analysis. In addition to the trenching work, seven (7) diamond drill holes were completed including one hole abandoned for technical reasons. All successful drill holes have intersected pegmatites zones.

A second exploration program was conducted from January to April, 2010. During that program, 59 drill holes totalling 11,600 m were completed. In addition to drilling, 14 line-km of ground magnetic surveying covering the main mineralized occurrence and 670 line-km of helicopter-borne magnetic surveying covering the Property were completed. Later in May 2010, the Company completed 2,780 m of mechanical stripping of the south contact of the main mineralized zone with (16 trenches and seven (7) contact zones) and collected 649 channel samples. The stripping also allowed the mapping of the surface geology. A 1.2 km access road from the Route du Nord main road was constructed in 2010.

In late 2010, 23 drill holes were completed, with an additional 26 holes in early 2011 for a total of 9,500 m. In May 2011, a 50-tonne bulk sample was collected at surface.

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10.0 DRILLING

This section is an extract of the 2011 Technical Report.

A total of 115 drill holes were completed by Nemaska to define the mineral deposit. In addition to the drilling, extensive mechanical stripping on surface permitted the completion of more than 140 channels. Table 10.1 and Table 10.2 summarize the drilling and channel sampling completed by Nemaska to define the mineralized pegmatite intrusion.

Table 10.1 – Drilling Completed by Nemaska at Whabouchi Year Count Metres Drilled 2009 7 915 2010 82 15,670 2011 26 5,500 Total 115 22,085 Table 10.2 – Channel Sampling Done by Nemaska at Whabouchi Year Channels Total Samples 2009 35 295 2010 108 649 Total 143 944 All the drilling done by Nemaska used NQ and HQ coring size. HQ size was used to collect material for metallurgical testing. The samples collected for analysis represent approximately 37% of the drill core material. The drill holes are generally spaced 25 m to 50 m apart with azimuth ranging between N312° and N340° with an average of N330°. The dips range from 43° to 75° and average 49°. The deepest hole reaches 500 m below surface. The mineralized drill intersection ranges from near true thickness to 70% true thickness. The geometry of spodumene-bearing pegmatites is defined as a series of stacked dyke-shaped intrusions which include a thicker principal intrusion. Some pegmatite contains local rafts or xenoliths of the host rock which can be a few metres thick and hundreds of metres in length. Based on the information gathered from the drilling, the pegmatite intrusion is more than 1,300 metres in length and can be up to 90 metres thick. The intrusions are generally oriented N030°, dip to the south at an angle ranging between 80 and 85 degrees and are reaching depths of up the 500 metres below surface. Please refer to section 14 for the interpretation of the drill results based on the results of mineral resources estimate. Figure 10.1, Figure 10.2 and Figure 10.3 show the drill holes in plan view, on longitudinal sections and on representative cross sections.

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Figure 10.1 – Plan View of the Drilling at Whabouchi

Figure 10.2 – Longitudinal View of the Drilling at Whabouchi

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Figure 10.3 – Section 400 mE and 900 mE Showing Drill Holes and Envelopes (Looking West)

The author completed verification programs of the Project’s analytical data and considers that there is no known drilling, sampling or recovery factors that could materially impact the accuracy and reliability of the results. The data from the few historical drill holes reported on the project could not be validated and were not considered as part of the current mineral resource estimate.

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11.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY

This section is an extract of the 2011 Technical Report.

This section is based on information supplied by Nemaska and observations made during the independent verification programs conducted at the Project site by SGS Geostat from March 10 to 12, 2010 and on May 4 and 5, 2011.

11.1 Sample Procedure and Sample Security

The Whabouchi Project is located less than 12 km east of the Nemiscau Camp where the Project office and core logging facilities are located. The evaluation of the geological setting and mineralization on the Property is based on observations and sampling from surface (through geological mapping, grab and channel samples) and diamond drilling. The channel and drill core logging and sampling was conducted at the Property or at the nearby Project facilities. All samples collected by Nemaska during the course of the 2009, 2010 and 2011 exploration programs were sent to the Table Jamésienne de Concertation Minière (“TJCM”) preparation laboratory located in Chibougamau, Quebec. The 2009 and 2010 sample pulps were shipped to SGS Canada Inc. – Mineral Services (“SGS Minerals”) laboratory in Don Mills, Ontario, for analysis. The 2011 sample pulps were sent to ALS Canada Inc. – Chemex Laboratory (“ALS Chemex”) in North Vancouver, B.C., for analysis. The remaining drill core is stored at the Property site in covered metal core racks.

All channel samples and drill core handling were done on site with logging and sampling processes conducted by employees and contractors of Nemaska. The observations on lithology, structure, mineralization, sample number, and location were noted by the geologists and technicians on hardcopy and then recorded in a Microsoft Access digital database. Copies of the database are stored on external hard drive for security.

Channel samples were collected from two (2) diamond saw cuts (typically 4 cm in width and 4 cm in depth). Each sample is generally one (1) m long and broken directly from the outcrop, identified and numbered then placed in a new plastic bag. Drill core of NQ and HQ size was placed in wooden core boxes and delivered twice daily by the drill contractor to the Project core logging facilities at the Nemiscau Camp. The drill core was first aligned and measured by a technician for core recovery. The core recovery measurements were followed by the RQD measurements. After a summary review of the core, it was logged and sampling intervals were defined by a geologist. Before sampling, the core was photographed using a digital camera and the core boxes were identified with Box Number, Hole ID, and by using “From” and “To” aluminum tags. Due to the hardness of the pegmatite units, the recovery of the channel material and the drill core was generally very good, averaging more than 95%.

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Sampling intervals were determined by the geologist, marked and tagged based on observations of the lithology and mineralization. The typical sampling length is one (1) m but can vary according to lithological contacts between the mineralized pegmatite and the host rock. In general, one (1) host rock sample was collected from each side that contacts the pegmatite. The NQ drill core samples were split into two (2) halves with one (1) half placed in a new plastic bag along with the sample tag; the other half was replaced in the core box with the second sample tag for reference. The third sample tag was archived on site. HQ size drill core was collected for a portion of the 2011 program for metallurgical purposes. The first half of the HQ drill core was selected for metallurgical testing. The second half was split in two (2) quarters, one (1) quarter placed in a new plastic bag along with the sample tag and the remaining quarter was replaced in the core box with the second sample tag for reference. The samples were then catalogued and placed in rice bags or pails, for shipping. The sample shipment forms were prepared on site with one (1) copy inserted with the shipment, one (1) copy sent by email to TJCM, and one (1) copy kept for reference. The samples were transported on a regular basis by Nemaska’s employees or contractors by pick-up truck directly to the TJCM facilities in Chibougamau. At the TJCM laboratory, the sample shipment was verified and a confirmation of shipment reception and content was emailed to Nemaska’s project manager.

SGS Geostat validated the exploration processes and core sampling procedures used by Nemaska as part of an independent verification program. The author concluded that the drill core handling, logging and sampling protocols are at conventional industry standard and conform to generally accept best practices. The author considers that the samples quality is good and that the samples are generally representative. Finally, the author is confident that the system is appropriate for the collection of data suitable for the estimation of a NI 43-101 compliant mineral resource estimate.

11.2 Sample Preparation and Analysis

Channel and drill core samples collected during the 2009, 2010 and 2011 exploration programs were transported directly by Nemaska representatives to the TJCM laboratory facilities in Chibougamau, Quebec for sample preparation. The submitted samples were pulverized at the TJCM laboratory to respect the specifications of the analytical protocol and then shipped to SGS Minerals or ALS Chemex for analysis. The author visited the TJCM facilities on March 10, 2010.

All samples received at TJCM were inventoried and weighted prior to being processed. Drying was done to samples having excess humidity. Sample material was crushed to 80-85% passing 2 mm using jaw crushers. Ground material was split using a split riffle to obtain a 275-300 g sub-sample. Sub-samples were then pulverized using a two- component ring mill (ring and puck mill) or a single component ring mill (flying disk mill) to 85-90% passing 200 mesh (75 µm). The balance of the crushed sample (reject)

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was placed into the original plastic bag. The pulverized samples were finally sent to SGS Minerals or ALS Chemex using Canada Post secured delivery services.

The majority of the 2009 and 2010 analyses were conducted at the SGS Minerals laboratory located in Don Mills, Ontario, which is an ISO/IEC 17025 laboratory accredited by the Standards Council of Canada. There are two (2) analytical methods used for the pulverized samples from the Whabouchi Project. The first analytical method used by SGS Minerals is the 55-element analysis using sodium peroxide fusion followed by both Inductively Coupled Plasma Optical Emission Spectrometry (“ICP- OES”) and Inductively Coupled Plasma Mass Spectrometry (“ICP-MS”) finish (SGS code ICM90A). This method uses 10 g of the pulp material and returns different detection limits for each element and includes 10 ppm lower limit detection for Li. The ICM90A analytical method was conducted at the beginning of the 2009-2010 exploration program to verify the content of other elements in the mineralization. The second method processed 20 g of pulp material and used the mineralization grade sodium peroxide fusion with ICP-OES finish methodology with a lower detection limit of 0.01% Li (SGS code ICP90Q). The ICP90Q analytical method was used at the beginning of the exploration program on samples analysed by ICM90A returning values greater than 0.3% Li. The ICP90Q method for Li was later used on a more systematic basis. Analytical results were sent electronically to Nemaska and results were compiled in an MS Excel spreadsheet by the project manager.

The 2010 pulp reanalysis and the 2011 analyses were conducted at ALS Chemex using the mineralization grade lithium four-acid digestion with Inductively Coupled Plasma – Atomic Emission Spectrometry (“ICP-AES”) (ALS code Li-OG63). The Li-OG63 analytical method used 4 g of pulp material and returned a lower detection limit of 0.01% Li.

11.3 Quality Assurance and Quality Control Procedure

Above the laboratory quality assurance quality control (“QA/QC”) routinely implemented by SGS Minerals and ALS Chemex using pulp duplicate analysis, Nemaska developed an internal QA/QC protocol consisting in the insertion of analytical standards, blanks and core duplicates on a systematic basis with the samples shipped to the analytical laboratories. In 2010, the Company also sent pulps from selected mineralized intersection to ALS Chemex for reanalysis. No pulp reanalysis was performed by the Company in 2011. The author did not visit the SGS Minerals or ALS Chemex facilities or conduct an audit of the laboratories.

11.4 Analytical Standards

Two (2) different standards were used by Nemaska for the internal QA/QC program: one (1) low grade lithium (“Li-LG”) and one (1) high grade lithium (“Li-HG”) standards.

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Both standards were custom made reference materials coming from historical drill core from the Whabouchi deposit itself. The preparation for the standards material has been conducted by TJCM using the same sample preparation protocol used for the regular Whabouchi samples. Each standard inserted weight between 90 and 120 g. In order to evaluate their expected values, Li-HG and Li-LG standards have been analysed six (6) times, each at the SGS Mineral Services laboratory in Don Mills, Ontario, and five (5) times each at the ALS Chemex laboratory in North Vancouver, British-Colombia. Both facilities are accredited ISO/IEC 17025 laboratories. The analytical protocol used at SGS Minerals is the mineralization grade sodium peroxide fusion with ICP-OES finish (SGS code ICP90Q). The analytical protocol used at ALS Chemex is the mineralization grade lithium four-acid digestion with Inductively Coupled Plasma – Atomic Emission Spectrometry (“ICP-AES”) finish (ALS code Li-OG63).

For the Li-LG standard, the analytical results returned from SGS Minerals for the six (6) samples averaged 0.46% Li versus an average of 0.45% Li for the five (5) samples submitted to ALS Chemex. For the Li-HG standards, the average of the six (6) samples returned 0.72% Li versus an average of 0.71% Li for the five (5) samples processed at ALS Chemex. Each laboratory showed relatively consistent analytical results from one (1) sample to another for each standard analysed. The averages for each standard also showed a good correlation between SGS Minerals and ALS Chemex. The amount of data is not statistically significant to calculate standard deviation (“Std. Dev.”) parameters which can be used to determine the success/failure of standards. Table 11.1 shows the results for each standard using both analytical protocols.

The insertion of the analytical standards Li-LG and Li-HG did not begin until drill hole WHA 10-15. After that, one (1) standard was inserted in the sample series at a rate of one (1) every 25 regular samples, alternating between the Li-LG and Li-HG standards. A total of 169 Li HG and 169 Li-LG standards were analysed during the 2010 and 2011 exploration campaigns, representing 3.8% of the samples analysed. In order to determine the QC warning (±2x Std. Dev.) and QC failure (±3x Std. Dev.) intervals for the Li-LG and Li-HG standards, the Std. Dev. parameters returned from the 169 Li-HG and 169 Li- LG analytical results are considered.

From the 169 Li-HG standards analysed, 14 standards fall outside the QC Warning interval and four (4) standards fall outside the QC Failure interval. After reviewing the four (4) failures, they are considered acceptable as the value falls within 12-15% of the expected value for Li-HG. From the 169 Li-LG standards analysed, six (6) fall outside the QC Warning interval and one (1) is considered a failure as it falls outside the QC Failure interval. After reviewing the only failure, it is considered acceptable as it returned 13% of the expected value for Li-LG. Table 11.2 reports the statistics of the Li-LG and Li-HG standards. Figure 11.1 shows plots of the variation of both standards with time.

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Table 11.1 – Sets Values for the Li-LG and Li-HG Standards SGS Minerals - ICP90Q Analytical Method Low Grade Standard (Li-LG) Low Grade Standard (Li-LG) Sample Li (%) Sample Li (%) Li-LG 1 0.47 Li-HG 1 0.72 Li-LG 2 0.46 Li-HG 2 0.72 Li-LG 3 0.46 Li-HG 3 0.72 Li-LG 4 0.46 Li-HG 4 0.71 Li-LG 5 0.46 Li-HG 5 0.71 Li-LG 6 0.46 Li-HG 6 0.72 Average 0.46 Average 0.72 ALS Chemex - Li-OG63 Analytical Method Low Grade Standard (Li-LG) Low Grade Standard (Li-LG) Sample Li (%) Sample Li (%) Li-LG 1 0.44 Li-HG 1 0.72 Li-LG 2 0.44 Li-HG 2 0.69 Li-LG 3 0.45 Li-HG 3 0.71 Li-LG 4 0.47 Li-HG 4 0.72 Li-LG 5 0.44 Li-HG 5 0.72 ALS Chemex - Li-OG63 Analytical Method Low Grade Standard (Li-LG) Low Grade Standard (Li-LG) Average 0.45 Average 0.71 Averages for Both SGS Minerals and ALS Chemex Methods Standard Li (%) Standard Li (%) Li-LG 0.46 Li-HG 0.71

Table 11.2 – Summary Statistics of Li-LG and Li-HG Standards Expected Expected QC QC Standard Count Observed Li (%) Li (%) (%) Warning Failure Average Average Std. Dev. Li-LG 169 0.46 0.47 0.017 102% 6 1 Li-HG 169 0.71 0.73 0.027 102% 14 4

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Figure 11.1 – Plots of the Variation of the Li-HG and Li-LG Standards with Time

11.5 Analytical Blanks

Nemaska implemented the insertion of analytical blanks in the sample series as part of their internal QA/QC protocol. The blank samples, which are made of coarse silica lumps, are inserted at every 20 samples in the sample series, at the beginning of the sample preparation procedure by TJCM before shipping to ALS Chemex. The analytical blanks used in 2010 were made of pre-pulverized silica instead of coarse lumps. The 2010 procedure was not considered adequate since the analytical blanks were inserted by TJCM after the sample preparation procedure. The QA/QC procedure was updated by the Company for 2011 and is now considered adequate.

A total of 255 analytical blanks were analysed during the 2009 and 2010 exploration programs and an additional 208 blanks were analysed for the 2011 program for a grand total of 463 results for analytical blanks available. From the 463 blanks analysed, 100% of them returned less than 0.05% Li, which is five (5) times the methods detection limit. Figure 11.2 shows a plot of the variation of the analytical blanks with time.

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Figure 11.2 – Plot of Variance of Analytical Blanks with Time

11.6 Core Duplicates

Sample duplicates were inserted at every 20 samples in the sample series as part of Nemaska’s internal QA/QC protocol. The sample duplicates correspond to a quarter NQ or HQ core from the sample left behind for reference, or a representative channel sample from the secondary channel cut parallel to the main channel. Figure 11.3 shows correlation plots for the core duplicates.

For 2010, a total of 254 duplicates results analysed by SGS Minerals are available. From the 254 core duplicates analysed in 2010, 91.3% of assay pairs with a grade higher than 0.05% Li (five (5) times the method detection limit) reproduced within ±20% and 92.7% of assay pairs with a grade higher than 0.1% Li reproduced within ±20%. The sign test for the 2010 duplicates analysed by SGS Minerals does not show any bias (30% original < duplicate, 33% original > duplicate, and 37% original = duplicate).

For 2011, a total of 58 duplicate results analysed by SGS Minerals and 145 duplicate results analysed by ALS Chemex are available. From the 58 core duplicates analysed by SGS Minerals, 96.6% of assay pairs with a grade higher than 0.05% Li (five (5) times the method detection limit) reproduced within ±10% and 96.4% of assay pairs with a grade higher than 0.1% Li reproduced within ±10%. The sign test for the 2011 duplicates analysed by SGS Minerals does not show any bias (40% original < duplicate, 34% original > duplicate, and 26% original = duplicate). From the 145 core duplicates analysed at ALS Chemex, 90.0% of assay pairs with a grade higher than 0.05% Li (five

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(5) times the method detection limit) reproduced within ±20% and 89.7% of assay pairs with a grade higher than 0.1% Li reproduced within ±20%. The sign test for the 2011 duplicates analysed by ALS Chemex does not show any significant bias (38% original < duplicate, 54% original > duplicate, and 8% original = duplicate).

Figure 11.3 – Correlation Plots for Core Duplicates

11.7 Nemaska Pulp Re-Analysis

As part of Nemaska’s 2010 QA/QC protocol, pulps from 192 mineralized core samples were sent for re-analysis to ALS Chemex. No pulp re-analysis was performed by the Company in 2011 at the time of writing the report for the updated mineral resource estimate. The re-analysed samples represent continuous mineralized intervals of different lengths selected from eight (8) drill holes (WHA-10-08, 11, 15, 21, 22, 28, 38 and 44). Figure 11.4 shows a correlation plot of the re-analysed pulps for SGS Minerals vs. ALS Chemex. Table 11.3 contains a comparison of the weighted average grade for each mineralized interval by hole. The pulp re-analysis returned higher Li values for SGS Minerals for 145 samples (or 76% of the samples reanalysed) compared to 23 samples (or 12%) returning lower Li values for SGS Minerals and 24 samples (or 13%) shows identical values for both laboratories. The SGS Minerals Li grades show a relative difference averaging 5.3% higher than ALS Chemex. As observed, seven (7) mineralized February 2013 QPF-009-12/B

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intervals out of a total of eight (8) show a higher weighted average grade for the SGS Minerals analysis. The relative difference of the weighted average grades for the different holes range from -1.4% to +8.9%. The results of the pulp re-analysis program conducted by the Company show a potential positive small analytical bias toward SGS Minerals analytical data, which could possibly be explained by the differences in analytical methodologies from one laboratory to another. Although SGS Geostat considers that the potential analytical bias observed in the pulps re-analysis results is significant enough to be investigated in more detail, the grade differences observed between the two (2) laboratories can be considered acceptable for a mineral resource estimate. The author recommends an in-depth comparison of the different analytical methods used by each laboratory be completed and that an additional pulp re-analysis of mineralized samples be carried out in order to verify the grade differences outlined herein.

Figure 11.4 – Correlation Plot of the Pulps Re-analysis for 2010

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Table 11.3 – 2010 Pulps Re-analysis Comparison by Drill Hole Mineralized Intervals Relative Hole ID From To Length Weighted Average Grade SGS ICP90Q ALS Li-OG63 Difference (m) (m) (m) Li2O (%) Li2O (%) (%) WHA-10-08 53.9 70.0 16.1 1.67 1.65 1.2% WHA-10-11 119.0 134.5 15.5 1.19 1.20 -1.4% WHA-10-15 111.0 198.5 87.5 1.44 1.38 3.8% WHA-10-21 191.8 209.8 18.0 1.71 1.58 7.2% WHA-10-22 32.9 47.0 14.1 1.82 1.72 5.4% WHA-10-28 120.2 126.0 5.8 1.29 1.25 3.0% WHA-10-38 214.6 230.7 16.1 2.20 2.00 8.9% WHA-10-44 93.4 116.2 22.8 1.91 1.81 5.2% 11.8 Specific Gravity

As part of the 2010 and 2011 independent data verification programs, SGS Geostat conducted specific gravity (“SG”) measurements on the 74 mineralized core samples collected from drill holes WHA-09-07, WHA-10-25, WHA-10-79, and WHA-11-96. The measurements were performed by the water displacement method using a graduated cylinder (SG = weight in air in kg/volume of water displaced in litre) on representative half NQ core and quarter HQ core pieces weighting between 0.42 kg and 2.28 kg with an average of 0.53 kg. The resulting measurements reported an average SG value of 2.70 t/m3, (Table 11.4).

Table 11.4 – SGS Geostat SG Measurements Statistical Parameters SG Measurements

(t/m3) Mean (t/m3) 2.70 Count 74.0 Standard Deviation 0.08 Rel. Std. Deviation (%) 2.80 Minimum (t/m3) 2.55 Median (t/m3) 2.70 Maximum (t/m3) 2.88 In 2011, Nemaska also conducted SG measurements of mineralized core samples. The measurements were conducted at TJCM laboratory in Chibougamau, by the water displacement method using weight in vs. weight in water. The SG measurements were done on samples from 24 different drill holes from 2009, 2010 and 2011. The resulting measurements reported an average SG value of 2.72 t/m3 (Table 11.5). Based on the

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available SG measurement datasets, an SG value of 2.70 t/m3 was selected as the average SG for the mineralization for the Whabouchi deposit.

Table 11.5 – Nemaska SG Measurements Statistical Parameters SG Measurements (t/m3) Mean (t/m3) 2.72 Count 73 Standard Deviation 0.03 Rel Std Deviation (%) 1.18 Minimum (t/m3) 2.65 Median (t/m3) 2.72 Maximum (t/m3) 2.87 11.9 Conclusion

Nemaska implemented an internal QA/QC protocol by regularly inserting reference materials (standards and blank) and core duplicates in the samples stream. The Company also conducted in the 2010 re-analysis of selected pulps in a second laboratory, as part of their QA/QC protocol.

SGS Geostat completed a review of the sample preparation and analysis including the QA/QC analytical protocol implemented by Nemaska for the Project. The author visited the sample preparation facilities at TJCM on March 10, 2010 and the Whabouchi Property from March 10 to 12, 2010 and again on May 4 and 5, 2011 to review the Company sample preparation procedures.

A review of the QA/QC analytical results for the standards, blanks and core duplicates did not highlight any analytical issues. The re-analysis of the pulps completed by the Company in 2010 outlined a potential small analytical bias from the results of the re- analysis of pulps from selected mineralized intervals. SGS Geostat recommends investigating this potential analytical bias which could potentially be caused in part by the different analytical methodologies used in the two (2) laboratories. SG measurements were completed in 2010 and 2011 on mineralized core samples to estimate an average bulk density value for the Whabouchi deposit.

The author is of the opinion that the sample preparation, analysis and QA/QC protocol used by Nemaska for the Whabouchi Project follow generally accepted industry standards and that the Project data is of a quality sufficient.

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12.0 DATA VERIFICATION

As part of the 2010 and 2011 data verification programs, SGS Geostat completed independent analytical checks of drill core duplicate samples taken from Nemaska 2009, 2010 and 2011 diamond drilling programs. SGS Geostat also conducted verification of the laboratories analytical certificates and validation of the Project digital database supplied by Nemaska for errors or discrepancies.

During a site visit conducted from March 10 to 12, 2010, a total of 35 mineralized core duplicates were collected from holes WHA-09-07 and WHA-10-25 by the author and submitted for Li analysis at the SGS Minerals laboratory in Don Mills, Ontario. The core duplicates were processed using the same analytical protocol used by Nemaska during the 2009 and 2010 drilling programs (code ICP90Q) except that the sample preparation has been done directly at the SGS Mineral Services and not at the TJCM laboratory. During the 2011 site visit, an additional 40 independent check samples were collected from holes WHA 10-079 and WHA-11-096. The core samples from these holes were originally analysed by ALS Chemex. The check analyses were performed at SGS Minerals laboratory in Don Mills.

The comparative results for the individual check samples are displayed in the correlation graph in Figure 12.1. Table 12.1 shows the comparative statistics for the 2010 and 2011 check samples. Table 12.2 reports the check sampling weighted average results for continuous mineralized intervals sampled for each drill hole.

On an individual sample basis, there are some divergences between the results of the check samples from 2010 versus 2011. In 2010, the original samples returned a higher grade than the check samples. In 2011, the original samples returned a lower grade than the check samples. In both years, the average relative grade differences between the original and the check samples range between 3% and 19%, which can be considered acceptable for core duplicates, considering the coarse nature of the spodumene mineralization generally observed at Whabouchi. The weighted average grades between the original and the check samples outline similar results except for hole WHA-10-25. For WHA-10-25, the original samples returned a higher weighted average than the check samples with a relative grade difference of 37%, which can be considered excessive for core duplicates, although most of the samples in the hole have relatively low grades compared to the other holes. The other three (3) holes returned an acceptable relative grade difference ranging between 1.1% and 6.0%. For the 2011 check sampling results, the pulps for the check samples were sent to ALS Chemex for reanalysis. The objective is to validate the significant variable observed between the original and the check samples. The reanalysis of the 2011 check samples completed at ALS Chemex for holes WHA-10- 079 and WHA-11-096 returned comparable results versus the original check samples

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analyzed at SGS Minerals. This final check validates the results of the independent check sampling conducted in 2011.

Figure 12.1 – Correlation Plot of 2010 and 2011 Independent Check Samples

Table 12.1 – Comparative Statistics for the Check Sampling Results Average Original > Original < Original = Relative Hole ID Count Check Check Check Percent Difference WHA-09-07 23 52% 43% 4% -4% WHA-10-25 12 75% 17% 8% -19% WHA-10-079 20 25% 75% 0% 7% WHA-11-096 20 35% 65% 0% 3%

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Table 12.2 – Check Sampling Comparison by Drill Hole Mineralized Intervals Relative Weighted Average Length Grade Hole ID From (m) To (m) (m) Original Check Difference Li2O (%) Li2O (%) (%) 105.94 124.51 18.6 1.77 1.84 -3.7% WHA-09-07 133.50 136.70 3.2 2.88 2.91 -1.1% WHA-10-25 30.50 42.50 12.0 0.47 0.32 36.8% WHA-10-079 113.3 133 19.7 1.75 1.65 6.0% WHA-11-096 215 235 20 1.80 1.77 1.4% The digital drill hole database supplied by Nemaska has been validated for the following data field: collar location, azimuth, dip, hole length, survey data, lithology and analytical values. The validation returned only minor discrepancies located in survey data, lithology and analytical values, which were communicated to Nemaska and corrected in the final drill hole database.

As part of the data verification of the Project, the analytical data from the database has been validated with the values from the laboratories analytical certificates. No errors were noted during the validation.

The final database includes the channel samples collected in 2009 and 2010 from surface trenches and the drilling data from the 2009, 2010 and 2011 drilling programs. The final drill hole with reported analytical results included in the database is WHA-10-112. The few historical drill hole and channel analytical data were not considered for the current mineral resource estimate. Table 12.3 lists the data contained in the final drill hole database. The author is in the opinion that the final drill hole database is adequate to support a mineral resource estimate.

Table 12.3 – Final Drill Hole Database Metres Survey Lithology Assays % Assayed Type Count Drilled Records Records Records Metres Channel 344 6,806 0 855 944 14% Drill hole 116 22,168 449 1,925 8,039 37% Total 460 28,974 449 2,780 8,983

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13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

Mineral processing and metallurgical testing were performed to evaluate the potential of spodumene concentrate production and lithium hydroxide (Li2O-H2O) and lithium carbonate (Li2CO3) production separately. The spodumene concentrate production test work is presented in Section 13.1. The hydrometallurgical production of Li2O-H2O and Li2CO3 test work is presented in Section 13.2.

13.1 Spodumene Concentrate

This section gives a review of bench scale and pilot scale metallurgical test work carried out in 2010 and 2011 at SGS Lakefield Research Limited (“SGS Lakefield”) to evaluate the amenability of Nemaska spodumene pegmatite for the production of a 6.0% Li2O spodumene concentrate. Some of the bench scale test work results were reported in the July 2011 NI 43-101 Technical Report of the Updated Mineral Resources Estimate. This Report provides an update of the bench scale tests carried out during and since the Updated Mineral Resources Estimate Report and provides a summary of the pilot scale test work carried out for the current PEA Study. The following Figure 13.1 illustrates the bench scale and pilot testing that was done in order to propose the PEA final flow sheet.

Figure 13.1 – Summary of the PEA Study Metallurgical Test Work

Benchscale Testwork Pilot Testwork

BenchscaleTestwork Pilot Testwork Samples Samples 25 tonne blasted outcrop 4 NQ holes 5 tonne blend of outcrop + drill core

Mineralogy DMS Testwork

Crushing and Flotation Testwork Grinding Testwork

Dense Media Sedimentation and Filtration Testwork Testing

Bench Scale Proposed Flowsheet Flotation Tests

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13.1.3 2010 and 2011 Mineral Processing and Metallurgical Bench Scale Test work a) Test Work Samples Four (4) NQ holes were drilled by Nemaska through the central portion of the main pegmatite in order to obtain approximately one (1) tonne of mineralized core exclusively for metallurgical test work. The core was sent by truck in a single palletized crate to the SGS Lakefield in July of 2010. The entire sample was stage- crushed to 6 mm (1/4”). The head grade of the sample was reported as 1.72% Li2O (Table 13.1).

The material collected from the pegmatite was 5.5% higher in Li2O than the average grade determined by Geostat and used in their resource estimation. This was considered to be within acceptable limits and appropriate for the metallurgical work for the PEA study. Table 13.2 below gives the mineralogical assay and indicates that spodumene makes up 21.1% of the sample.

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Table 13.1 – Head Grade of 2010 Metallurgical Sample (Assays Reported In Percent)

Sample Li2O SiO2 Al2O3 Fe2O3 MgO CaO Na2O K2O TiO2 P2O5 MnO Cr2O3 V2O5 LOI Sum Cut A 1.70 74.5 16.0 0.99 0.11 0.49 3.31 2.67 0.01 0.15 0.09 0.02 ˂0.01 0.06 98.4 Cut B 1.74 76.0 16.4 0.99 0.14 0.50 3.39 2.71 0.02 0.15 0.09 0.02 ˂0.01 -0.08 100 Average 1.72 75.3 16.2 0.99 0.13 0.50 3.35 2.69 0.02 0.15 0.09 0.02 -- -0.01 99.2

Note: Li2O is determined by ICP-OES and the other oxides by WRA

Table 13.2 – Head Grade Semi-Quantitative X-Ray Diffraction Results Head Sample Cut A Mineral (wt %) Quartz 32.4 Albite 26.5 Microcline 9.4 Muscovite 8.5 Fluorapatite 0.8 Spodumene 21.1 Siderite 1.3 Total 100.0

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b) Mineralogical Test Work Figure 13.2 below summarizes the mineralogical test work done with the mineralized material. The objective of mineralogical test work is to predict the amount of liberated mineral of interest at varying size distribution.

Figure 13.2 – Summary of Mineralogical Test Work

XRD QUEMSCANTM

Stage Crushed at 4 NQ Drill Core Six Drill Core QUEMSCANTM -212 µm Stage Crushed to Samples 6 mm -10 mesh O/S Stage Crushed to K80 at 425 µm Screened at Five Size Blending = Master XRD -212 µm Composite Fraction +425 to - QUEMSCANTM 38 µm U/S

Six (6) drill core samples, named MW, ME, SE, PP, and WP, were submitted to SGS Lakefield as 10 mesh material for mineralogical analysis.

From each -10 mesh composite sample, a portion was submitted for XRD and QEMSCANTM analysis. Another portion from each composite was stage-crushed to -212 µm and submitted for QEMSCANTM. A third portion from each composite sample (with the exception of WP, which was considered waste) was screened at 212 µm to temporarily remove the fine particles. The oversized material was stage- crushed to K80 425 µm. The -212 µm material and stage-crushed oversized material were then blended to create a master composite sample. The master composite sample was then split into five (5) size fractions ranging from +425 µm to -38 µm and submitted for XRF and QEMSCANTM analysis.

The XRD and QEMSCANTM analyses for each of the -10 mesh composite samples were in close agreement. For the QEMSCANTM analysis, it was difficult to distinguish between the spodumene and petalite minerals, and therefore, the minerals were combined and reported as Li Minerals. The XRF results are provided in Table 13.3 and the QEMSCANTM results are shown in Table 13.4. MW, SE, and WP samples contained no petalite. Sample WP was also low in lithium and was considered as waste.

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Table 13.3 – Semi-Quantitative X-Ray Diffraction Results

(1) MW -10 m (2) ME -10 m (3) SW -10 m Mineral (wt %) (wt %) (wt %) Quartz 36.9 36.2 32.6 Albite 26.7 25.1 21.9 Spodumene 14.4 20.2 12.8 Muscovite 11.7 9.1 10.0 Microcline 9.8 6.0 9.9 Beryl 0.5 - - Petalite - 3.4 12.8 Kaolinite - - -

Total 100.0 100.0 100.0

(4) SE -10 m (5) PP -10 m (6) WP -10 m Mineral (wt %) (wt %) (wt %) Quartz 33.8 35.3 29.8 Albite 28.8 25.5 37.3 Spodumene 15.5 10.4 - Muscovite 7.6 9.6 5.5 Microcline 14.3 11.6 26.3 Beryl - - - Petalite - 7.5 - Kaolinite - - 1.0 Total 100.0 99.9 99.9

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Table 13.4 – Modal Analysis of -10 Mesh Samples Survey Nemaska Exploration Project 12240-001/MI5011-MAY10 MW -10 Sample ME -10 m SW -10 m SE -10 m PP -10 m WP -10 m m Fraction -1,700 µm -1,700 µm -1,700 µm -1,700 µm -1,700 µm -1,700 µm Mass Size Distribution 100.0 100.0 100.0 100.0 100.0 100.0 (%) Particle Size 171 144 173 258 127 128 Sample Sample Sample Sample Sample Sample Na- 24.2 24.4 20.7 26.0 23.0 39.1 Feldspar Quartz 32.9 34.1 36.1 30.5 32.0 25.1 Microcline 17.9 10.4 15.2 16.8 15.6 29.3 Li Minerals 19.1 25.6 19.5 22.2 24.7 1.8 Muscovite 5.1 4.5 6.4 3.2 3.6 2.9 Mineral Ta-Nb- Mass (%) 0.0 0.2 0.1 0.0 0.1 0.0 Minerals Garnet 0.4 0.4 0.9 0.8 0.3 0.8 Biotite 0.0 0.0 0.0 0.0 0.0 0.1 Apatite 0.0 0.1 0.1 0.2 0.1 0.0 Other 0.3 0.3 0.9 0.3 0.7 0.9 Total 100.0 100.0 100.0 100.0 100.0 100.0 Na- 133 117 131 178 112 129 Feldspar Quartz 156 138 156 199 115 132 Microcline 127 105 139 159 95 97 Mean Li Minerals 146 141 132 231 104 54 Grain Size Muscovite 105 95 127 93 80 57 by Ta-Nb- Frequency 29 148 129 26 101 26 (µm) Minerals Garnet 60 76 80 78 55 109 Biotite 28 31 24 35 24 44 Apatite 27 45 41 85 42 28 Other 25 27 52 25 26 27

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The QEMSCANTM results for the samples ground to -212 µm are shown in Table 13.5. As can be seen from Figure 13.3, the results of the bulk modal abundance of the -10 mesh samples and the K80 -212 µm samples are in good agreement.

Table 13.5 – Modal Analysis of K80 -212 µm Samples Survey Nemaska Exploration Project 12240-001/MI5011-MAY10 MW 212 ME 212 SW 212 SE 212 PP 212 WP 212 Sample µm µm µm µm µm µm Fraction -212 µm -212 µm -212 µm -212 µm -212 µm -212 µm Mass Size Distribution (%) 100.0 100.0 100.0 100.0 100.0 100.0 Particle Size 67 74 80 67 51 69 Sample Sample Sample Sample Sample Sample Na-Feldspar 25.5 24.9 19.2 27.4 24.6 38.9 Quartz 38.0 36.6 38.8 23.8 28.7 28.1 Microcline 14.0 8.5 13.6 18.2 13.9 28.6 Li Minerals 17.7 26.9 22.6 27.8 28.3 2.2 Muscovite 4.0 2.5 3.1 2.5 4.2 1.9 Mineral Mass Ta-Nb- 0.1 0.0 0.0 0.0 0.0 0.0 (%) Minerals Garnet 0.4 0.2 1.6 0.2 0.0 0.1 Biotite 0.0 0.0 0.0 0.0 0.0 0.1 Apatite 0.0 0.2 0.1 0.0 0.0 0.0 Other 0.2 0.1 0.8 0.1 0.3 0.1 Total 100.0 100.0 100.0 100.0 100.0 100.0 Na-Feldspar 54 62 58 56 43 58 Quartz 75 82 76 64 53 70 Microcline 44 50 59 54 36 55 Li Minerals 48 58 53 59 53 28 Mean Grain Muscovite 22 23 21 18 19 17 Size by Ta-Nb- Frequency 37 10 19 0 6 5 (µm) Minerals Garnet 38 37 24 13 5 6 Biotite 7 15 6 5 7 14 Apatite 9 99 16 9 11 9 Other 11 7 22 6 8 7

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Figure 13.3 – Comparison of Mineral Distributions Between -10 Mesh and K80 -212 µm Samples

Table 13.6 shows the normalized mass of Li minerals in the K80 -212 µm samples. Free and liberated Li minerals range from about 85% for sample MW to about 98% for sample SE.

Table 13.6 – Normalized Mass of Li Minerals in K80 -212 µm Samples Mineral Name ME 212 µm MW 212 µm PP 212 µm SE 212 µm SW 212 µm WP 212 µm Free Li Min 83.3 80.8 70.0 77.0 66.9 6.5 Lib Li Min 12.3 5.4 19.9 20.9 20.9 51.7 Li Min: Micr 0.1 0.1 4.5 0.0 0.2 9.1 Li Min: Musc 0.0 0.0 0.0 0.0 0.1 0.0 Li Min: Na- 1.9 0.6 1.8 0.3 1.7 3.9 Felds Li Min: Qtz 1.2 0.6 1.9 1.3 3.2 7.0 Li Min: Ta-Nb 0.0 0.0 0.0 0.0 0.0 0.0 Minerals Li Min: Micr. 1.3 11.4 1.7 0.4 6.6 21.1 Musc. Qtz. Flds Li Min : Other 0.0 0.0 0.2 0.0 0.0 0.0 Li Min 0.0 1.1 0.1 0.0 0.3 0.6 Complex Total 100.0 100.0 100.0 100.0 100.0 100.0

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Table 13.7 shows the modal analysis of the master composite sample by particle size range. The Li minerals tend to be more concentrated in the coarser fractions. Table 13.8 shows the liberation of Li Minerals by size fraction in the master composite sample. As expected, the Li mineral liberation increases with decreasing particle size. The liberation does not increase significantly below -212 µm.

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Table 13.7 – Bulk Modal Analysis of Master Composite Sample Survey Nemaska Exploration Project 12240-001/MI5019-JUN10 Sample Composite Fraction Combined +425 µm -425 /+212 µm -212/+106 µm -106/+38 µm -38 µm Mass Size Distribution (%) 19.8 34.0 17.4 14.4 14.3 Particle Size 54 425 193 92 44 14 Sample Sample Fraction Sample Fraction Sample Fraction Sample Fraction Sample Fraction Li Minerals 22.4 6.3 32.0 8.6 25.2 3.4 19.3 2.4 16.8 1.6 11.4 Quartz 31.4 6.1 30.6 10.1 29.7 5.7 32.7 4.7 32.9 4.8 33.2 Na-Feldspar 25.3 3.3 16.9 8.7 25.5 5.1 29.4 4.3 29.6 3.9 27.4 Microcline 15.9 2.7 13.8 4.8 14.1 2.7 15.3 2.5 17.3 3.3 22.7 Muscovite 3.4 0.6 3.1 1.4 4.2 0.5 2.6 0.4 2.5 0.6 4.0 Mineral Mass Biotite 0.0 0.0 0.1 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 (%) Garnet 1.1 0.6 3.2 0.3 0.9 0.1 0.4 0.1 0.4 0.0 0.3 Ta-Nb- 0.1 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Minerals Apatite 0.2 0.0 0.1 0.1 0.2 0.0 0.1 0.0 0.1 0.0 0.2 Other 0.2 0.0 0.1 0.0 0.1 0.0 0.2 0.0 0.3 0.1 0.7 Total 100.0 19.8 100.0 34.0 100.0 17.4 100.0 14.4 100.0 100.0 100.0 Li Minerals 243 126 75 38 14 Quartz 247 152 86 44 16 Na-Feldspar 191 132 83 40 13 Microcline 244 140 84 42 12 Mean Grain Muscovite 102 64 44 26 9 Size by Biotite 26 13 12 12 5 Frequency (µm) Garnet 209 92 68 30 10 Ta-Nb- 16 8 38 28 6 Minerals Apatite 29 67 25 21 9 Other 28 23 30 24 11

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Table 13.8 – Normalized Liberation Mass of Li Minerals for Master Composite Sample Mineral Name Combined +425 µm -425 /+212 µm-212/+106 µm -106/+38 µm -38 µm Free Li Min 69.4 55.9 67.4 81.0 81.4 90.9 Lib Li Min 17.0 23.0 18.0 11.5 14.2 3.9 Li Min: Micr 0.5 0.5 0.6 0.3 0.3 1.1 Li Min: Na-Felds 1.6 1.9 1.8 1.1 1.0 1.8 Li Min: Qtz 4.8 5.8 6.4 2.7 1.7 0.9 Li Min: Musc 0.1 0.0 0.0 0.3 0.1 0.0 Li Min : Micr. 6.4 12.4 5.7 3.0 1.2 1.3 Musc. Qtz. Flds Li Min : Other 0.0 0.0 0.0 0.1 0.1 0.0 Li Min Complex 0.2 0.5 0.1 0.0 0.1 0.1 Total 100.0 100.0 100.0 100.0 100.0 100.0 Figure 13.4 shows the mineral release curves of Li minerals, microcline, and muscovite, for the master composite sample. Mineral release curves are used to predict the amount of liberated mineral of interest at varying size distribution. This is an indicator of optimum grind targets in order to achieve the most liberation for the least amount of grind energy. The liberation of Li minerals increases by only 2-3% below 212 µm.

Figure 13.4 – Mineral Release Curves of Li Minerals, Microcline, and Muscovite for Master Composite Sample

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Figure 13.5 illustrated the grade-recovery curves for the master composite sample at various particle size ranges. As expected, grade and recovery increases with decreasing particle size. Approximately 86% of Li minerals recovery can be achieved at a grade of 3.4% Li.

Figure 13.5 – Grade vs. Recovery Curves for Li Minerals of Master Composite Sample

c) Variability Crushing and Grinding Test Work A total of 18 variability samples were submitted to SGS Lakefield for Bond ball mill and Bond rod mill grindability tests, as well as Bond abrasion tests. One (1) outcrop sample was submitted for the Bond low-energy impact test. The sample location information and grindability results are summarized in Table 13.9 and the grindability test statistics are provided in Table 13.10.

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Table 13.9 – Grindability Test Summary Head Sample Drill Interval (m) Work Indices (kWh/t) AI Grade Li O1 Name Hole From To 2 CWI RWI BWI (g) (%) Sample #1 WHA-10-78 130.0 140.0 1.5 10.3 11.0 0.663 Sample #2 WHA-10-78 145.0 155.0 1.6 12.4 12.9 0.737 Sample #3 WHA-10-78 160.0 170.0 2.5 12.1 15.4 0.689 Sample #4 WHA-10-79 115.0 125.0 1.8 12.1 12.0 0.661 Sample #5 WHA-10-79 130.0 140.0 1.5 12.2 14.1 0.510 Sample #6 WHA-10-79 150.0 160.0 1.4 11.4 12.7 0.73 Sample #7 WHA-10-81 178.0 188.0 1.6 12.6 12.4 0.622 Sample #8 WHA-10-81 210.0 220.0 2.5 12.4 13.2 0.641 Sample #9 WHA-10-81 240.0 250.0 1.3 10.5 11.6 0.65 Sample #10 WHA-11-097 338.0 348.0 2.0 11.4 13.2 0.608 Sample #11 WHA-10-89 67.0 77.0 2.5 12.1 14.1 0.704 Sample #12 WHA-10-89 85.0 95.0 2.3 12.0 14.3 0.621 Sample #13 WHA-11-090 51.0 60.9 1.9 10.3 12.7 0.507 Sample #14 WHA-11-091 61.0 71.0 2.0 11.8 12.2 0.607 Sample #15 WHA-11-095 185.0 195.0 2.6 12.0 14.6 0.550 Sample #16 WHA-11-092 68.0 78.3 2.2 12.3 13.4 0.706 Sample #17 WHA-11-093 205.7 215.0 2.1 11.2 13.9 0.658 Sample #18 WHA-11-094 134.0 144.0 2.0 11.4 13.3 0.572 Outcrop 14.1

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Table 13.10 – Grindability Test Statistics

Head Grade RWI BWI AI Statistics Li2O (%) (kWh/t) (kWh/t) (g) Average 2.0 11.7 13.2 0.635 Std. Dev. 0.4 0.7 1.1 0.068 Rel. Std. Dev. 45.2 6 8 11 Minimum 1.3 10.3 11.0 0.507 10th Percentile 1.5 10.5 11.9 0.538 25th Percentile 1.6 11.4 12.5 0.607 Median 2.0 12.0 13.2 0.645 75th Percentile 2.3 12.2 14.0 0.683 90th Percentile 2.5 12.4 14.4 0.713 Maximum 2.6 12.6 15.4 0.737 According to the SGS Lakefield database, the samples tested were considered to be in the soft to moderately soft range of hardness with respect to the Bond Rod mill Work Index (“RWI”) and in the soft to medium range with respect to Bond Ball mill Work Index (“BWI”). The outcrop sample was characterized as being hard in terms of Bond Crushing Work Index (“CWI”). The samples were also characterized as being abrasive. The variability in hardness among the samples tested was relatively low. For equipment sizing purposes, BBA chose the 75th Percentile values for RWI, BWI, and AI.

All of the variability samples were also assayed by the XRD Whole Rock Analysis method (WRA) and the results are presented in Table 13.11.

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Table 13.11 – WRA of Variability Samples Sample Grade (%)

Name SiO2 Al2O3 Fe2O3 MgO CaO Na2O K2O TiO2 P2O5 MnO Cr2O3 V2O5 LOI SUM Sample #1 74.4 15.9 0.45 0.04 0.22 4.39 2.17 ˂0.01 0.30 0.14 0.02 ˂0.01 0.85 98.8 Sample #2 75.0 15.8 0.61 0.03 0.21 3.20 2.74 0.01 0.20 0.11 0.03 ˂0.01 0.91 98.8 Sample #3 75.6 15.7 0.79 0.08 0.26 1.68 2.54 ˂0.01 0.12 0.07 0.03 ˂0.01 0.80 97.7 Sample #4 74.3 16.0. 0.49 0.04 0.28 3.36 2.36 ˂0.01 0.25 0.09 0.02 ˂0.01 0.84 98.1 Sample #5 74.5 15.5 0.61 0.06 0.20 2.13 4.18 ˂0.01 0.15 0.08 0.03 ˂0.01 0.75 98.2 Sample #6 75.4 15.0 0.64 0.08 0.21 2.35 3.72 ˂0.01 0.12 0.06 0.02 ˂0.01 0.70 98.3 Sample #7 72.6 16.2 0.87 0.21 0.62 3.02 3.61 0.03 0.16 0.13 0.02 ˂0.01 0.77 98.3 Sample #8 74.6 16.5 0.63 0.05 0.22 2.84 1.93 ˂0.01 0.23 0.16 0.03 ˂0.01 0.69 98.2 Sample #9 75.0 15.4 0.70 0.07 0.31 4.04 2.20 ˂0.01 0.09 0.11 0.03 ˂0.01 0.57 98.5 Sample #10 74.0 15.8 0.80 0.07 0.25 2.95 2.45 ˂0.01 0.08 0.11 0.03 ˂0.01 0.54 97.1 Sample #11 75.6 15.8 0.78 0.05 0.21 2.28 2.09 ˂0.01 0.05 0.13 0.03 ˂0.01 0.61 97.7 Sample #12 75.0 16.4 0.76 0.07 0.22 3.12 1.72 ˂0.01 0.04 0.12 0.03 ˂0.01 0.51 97.9 Sample #13 74.3 16.4 0.61 0.05 0.25 3.85 1.98 ˂0.01 0.10 0.12 0.02 ˂0.01 0.42 98.1 Sample #14 73.8 16.0 0.54 0.06 0.22 4.0 2.66 ˂0.01 0.08 0.12 0.03 ˂0.01 0.50 98.0 Sample #15 75.9 16.2 0.82 0.06 0.20 2.40 1.83 ˂0.01 0.05 0.10 0.03 ˂0.01 0.37 97.9 Sample #16 75.0 16.0 0.66 0.07 0.24 3.08 1.96 ˂0.01 0.07 0.13 0.03 ˂0.01 0.44 97.6 Sample #17 74.3 16.6 0.69 0.08 0.25 3.83 1.55 0.01 0.08 0.14 0.02 ˂0.01 0.50 98.0 Sample #18 73.6 16.4 0.94 0.18 0.47 3.99 1.33 0.03 0.10 0.13 0.03 ˂0.01 0.46 97.6 Average 74.6 16.0 0.69 0.08 0.27 3.14 2.39

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d) Dense Media Test Work In the July 2011 NI 43-101 Report for the Updated Mineral Resources Estimate, it was determined that Dense Media Separation using a heavy medium with a specific gravity (“SG”) of 2.7 would not be a suitable method for pre-concentration and would only be suitable as a means of upgrading the mineralized material after grinding to 10 mesh. Further test work was proposed and subsequently, four (4) series of heavy liquid tests were carried out. These tests are summarized in this section. More detailed results are provided in the SGS Lakefield report of October 20, 2011 (project 12486-001).

i) First Series of Heavy Liquid Testing – HL1

The first series of heavy liquid tests were performed on two (2) fractions: -12.7 mm + 4 mm and 4 mm + 0.84 mm. The results are summarized in Figure 13.6.

Figure 13.6 – Heavy Liquid Test Results – HL1

From these results it was concluded that: – For the coarser size fraction, it is possible to reject about 50% of the mass to the float product with a lithium loss of about 12%. The sink product would contain 87.9% of the Li with a grade of 3.23% Li2O. At a mass rejection of about 20%, the Li loss would be less than 5% with a lithium grade of about 1%. – For the finer fraction, it is possible to reject 62.8% of the mass to the float product with a Li loss of about 10%. The sink product would contain 87.9% of the Li with a grade of 1.87%. Through interpolation, it is predicted that by adjusting the density to about 2.65 g/cm3, it may be

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possible to reject 35% of the mass with a lithium loss of less than 5% at a concentrate grade of 1.13% Li2O. – Further detailed analysis of the results indicated a possibility to produce an acceptable concentrate grade for Hydromet processing. The feed size fraction should be between ¼" and 35 mesh (6.35 mm and 0.5 mm). First, the media density will be adjusted to 2.5 g/cm3 to reject petalite to the float product. In the next stage, the media density would be raised to 2.65 g/cm3 to reject silicate gangue minerals. The sink product would then be further 3 processed at 2.9 g/cm . The sink product is expected to contain 6.0% Li2O with a global spodumene recovery of 40%. The float product will contain some non-liberated spodumene and will be combined with the -0.5 mm fraction and further processed by flotation. This formed the basis for the second series of heavy liquid testing. ii) Second Series of Heavy Liquid Testing – HL2

The test flow sheet for the second series of HLS tests is illustrated in Figure 13.7. As indicated in the previous section, the feed material was primary crushed to 6.35 mm. The crushed product was then processed sequentially at heavy liquid densities of 2.5 g/cm3 to remove petalite and to 2.65 g/cm3 to reject the gangue minerals. The sink product was further crushed to 3.35 mm and further processing at an SG of 3.0 g/cm3. The test results indicated that: – Processing the -6.35 mm + 0.5 mm size fraction at an SG of 2.5 g/cm3 can reject about 30% of the mass with a lithium loss of about 3.3%; – About 0.46% of the mass reported to the 2.5 g/cm3 float product and contained mainly petalite; – About 30% of the material reported to the 2.65 g/cm3 float product with a lithium loss of 3.35%;

– A final dense media circuit concentrate grade of 6.0% Li2O was produced with a lithium recovery of 37%; – All the high grade rejected material was to be combined with the -0.5 mm undersize material for further processing by flotation. The flotation feed contained 59.1% of the total lithium and represented 58.5% of the original mass. The flotation feed grade was 1.66% Li2O, which was similar to the original feed grade.

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Figure 13.7 – Test Flow sheet for HL2

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iii) Third Series of Heavy Liquid Testing – HL3

The test flow sheet for the third series of HL tests was fairly similar to the second. The major differences were the feed was crushed to 3/8” (9.5 mm) instead of ¼” (6.35 mm), implementation of a third crushing stage to 2 mm and repassing the final middlings through the heavy liquid at 2.65 g/cm3 to remove gangue middlings. The test results indicated that: – The lithium recovery to DMS concentrate can be increased from 37% to about 42.5%; – About 38% of the mass can be rejected as gangue before further processing; – The flotation feed would represent 48.6% of the original feed mass containing 45.7% of the total lithium at a grade of 1.57% Li2O. iv) Fourth Series of Heavy Liquid Testing – HL4

In the fourth series of tests it was decided to reduce the number of grinding stages to two (2), as shown in Figure 13.8. The test results indicated that: – The lithium recovery to DMS concentrate can be increased to about 50%; – About 40% of the mass can be rejected as gangue; – The flotation feed would represent 45% of the original feed mass containing 43.6% of the total lithium at a grade of 1.6% Li2O.

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Figure 13.8 – Test Flow sheet for HL4

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e) Bench Scale Flotation Test Work Bench scale flotation tests were carried out on the head sample and on the combined heavy liquid middlings and undersize fraction (-0.5 mm) from the dense media test work.

All flotation tests were conducted in a Denver 12 or a 10-kg flotation cell. All grinding was performed either in a 12-kg Titan laboratory mill or in a 10-kg mill. In all tests, steel rods were used as grinding media. After the stage grinding at a closing screen size of 300 microns, a 50-mm Mozley cyclone was used to remove the primary slimes. Scrubbing was performed in a Denver flotation cell at pH 11 and the presence of lignin sulfonate for 10 minutes. Most of the bench scale flotation tests were performed batch-wise with no recirculation of streams. However, two (2) locked cycle tests were conducted. The locked-cycle flow sheets tested were developed from the initial batch-wise bench scale flotation tests. Some key findings of the flotation test program are summarized below.

The results of desliming are illustrated in Figure 13.9.

Figure 13.9 – Correlation between Slime Mass Fraction and Lithium Loss

There is a strong linear correlation between lithium loss and mass fraction of slimes. On average, about 4.24% of the composite sample reported to the cyclone overflow, which corresponded to a lithium loss of 2.9%. The cut size (K80) of the overflow was less than 10 µm.

Since the mica grade in the sample was relatively high at 8.5%, mica flotation was carried out prior to spodumene flotation. The main variables investigated in the

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flotation test work were spodumene and mica collector types and reagent dosages. The most important reagents used were: • Armac T and Armac C collectors for mica and feldspar flotation; • Aero 3030C collector for muscovite flotation; • Kersone to promote coarse mica flotation; • LR19, LR19B, and FA-2 collectors for spodumene flotation; • Marasperse D618, which is a calcium lignin sulphonate reagent, for promoting the dispersion of fine particles and improving desliming; • Sulphuric acid to reduce pulp pH before amine conditioning (Armac C or T) in mica flotation; • Soda ash for pH control in spodumene flotation. In the mica flotation tests, Armac C was found to be a stronger mica collector, resulting in a slightly higher K2O distribution in the mica concentrate. On average, the mica concentrate contained about 3 to 5% of the lithium. A portion of the lithium is in the mica lattice and is not due to spodumene deportment to the mica concentrate. Consequently, there is a strong linear correlation between lithium loss to the concentrate and mica concentrate mass pull.

The effect of eliminating mica flotation was investigated and it was concluded that eliminating mica flotation can impose negative effects on spodumene flotation, making it difficult to reach the desired grade.

The effects of different collectors were evaluated and the results are provided in Figure 13.10. Collector D30LR showed relatively poor performance in terms of maximum concentrate grade. Collector LR19 generated good concentrate grade, but the recovery was slightly lower in comparison to the other collectors. It was recommended that future confirmatory test work should focus on LR19, LR19B, and FA-2.

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Figure 13.10 – Effect of Different Collectors on Spodumene Flotation

The effect of primary grind size was investigated and the results are shown in Figure 13.11. These tests were performed by stage grinding 10 kg charges in a 10-kg rod mill to 100% passing 380 µm, 300 µm, and 210 µm. The grade-recovery relationships were similar for the coarser grinds and slightly better at the finer grind. The final concentrate grades were higher for the two (2) finer grade tests.

As mentioned above, two (2) locked-cycle flotation tests were carried out. The flow sheet for the first locked-cycle test is illustrated in Figure 13.11. The -3.35 mm feed charge was stage-ground to 300 µm, followed by a desliming stage. The combined mica rougher and cleaner tails were scrubbed, followed by a second stage of desliming and then high-density conditioning at an ambient temperature (22˚C). After rougher flotation, the rougher tails were conditioned at high density with additional collector. The rougher and rougher scavenger concentrates were combined and cleaned in four (4) consecutive cleaning stages. The final concentrate from each cycle was passed through a wet high-intensity magnetic separator. LR19B was used as the collector and D618 was used as the dispersant in the scrubbing stage. The scavenger tails and magnetic product from the wet high- intensity magnetic separator were considered as final tails. The conclusions from the first locked-cycle test were as follows:

• The final concentrate grade was 6.41% Li2O with a recovery of 84.2%. The scavenger tails graded 0.09% Li2O with a lithium deportment of 6.81%. The lithium loss to the magnetic stream was 0.58%. Magnetic separation increased the concentrate grade from 6.35 to 6.41% Li2O and reduced the iron from 1.9% to 1.47% Fe2O3; February 2013 QPF-009-12/B

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• The highest Li losses were to the scavenger tails;

• The K80 of the rougher tail was 230 µm and the cyclone overflow was 100% less than 20 µm. In the second locked-cycle test, LR19B was replaced with LR19, and ferric sulphate was added to the rougher tail to activate coarse spodumene grains.

The major difference in performance was the continuous decrease of the spodumene concentrate grade. The concentrate grade in the first cycle was 6.29% Li2O and it gradually decreased to 5.17% Li2O in the final cycle. It was believed that ferric iron had a negative effect on spodumene flotation and it activated feldspar and quartz particles. This test indicated that the depression of fine gangue minerals is crucial for spodumene flotation.

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Figure 13.11 – Locked-Cycle Test Flow sheet – LCT1

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13.1.4 Pilot Scale Dense Media Separation and Flotation Test Work a) DMS Pilot Testing i) Test Work Samples

Dense Media Separation (“DMS”) pilot testing was carried out on two (2) separate samples from the Whabouchi deposit. The first being a 25 tonnes blasted outcrop sample and the second was a 5 tonnes mine representative sample consisting of a blend of outcrop and drill core materials.

Testing was carried out in several stages consisting of crushing, scrubbing, screening, DMS, magnetic separation, and dewatering. The larger blasted outcrop sample was processed first to assess the effectiveness of various processing strategies. The processing route for the smaller mine representative sample was based on the results of the larger blasted sample and was considered to be the most likely flow sheet to be used in actual production.

ii) Blasted Sample Testing

The blasted sample was processed in a total of eight (8) DMS stages to assess the effectiveness of various upgrading stages, both by DMS and magnetic separation. In addition, the impact of separation at three (3) different feed top sizes was also assessed.

iii) Mine Representative Sample Testing

Based on test data during processing of the blasted sample, the mine representative sample was processed using a simplified flow sheet and was considered to approximate the final future DMS plant flow sheet. The test flow sheet scheme is illustrated in Figure 13.12.

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Figure 13.12 – DMS Processing Flow sheet – Mine Representative Sample

A detailed mass balance is provided in Table 13.12.

Table 13.12 – Detailed Mass Balance – Mine Representative Sample Stream Stream ID Stream Wt Wt Grade Li Distribution # Type (t) (%) (% Li2O) (%) 1 DMS Feed Feed 4.42 100.0 1.61 100.0 2 DMS #1 Floats Tailings 1.32 29.8 0.31 5.7 3 DMS #1 Sinks Intermediate 2.04 46.1 2.75 79.0 4 DMS #1 Fines PP Feed 0.89 20.2 1.02 12.8 5 DMS #1 Slimes Tailings 0.17 3.9 1.02 2.5 6 DMS #2 Floats Intermediate 1.67 37.8 2.16 50.9 7 DMS #2 Sinks Concentrate 0.29 6.6 6.22 25.7 8 DMS #2 Fines PP Feed 0.06 1.3 2.32 1.9 9 DMS #2 Slimes Tailings 0.01 0.3 2.32 0.4 10 DMS #2 Nonmag 0.28 6.4 6.28 25.2 11 DMS #2 Mag 0.01 0.2 4.21 0.5 12 DMS #3 Floats Intermediate 0.81 18.3 1.49 17.0 13 DMS #3 Sinks Concentrate 0.21 4.7 6.59 19.1 14 DMS #3 Fines PP Feed 0.60 13.6 1.60 13.6 15 DMS #3 Slimes Tailings 0.05 1.2 1.60 1.2 16 DMS #3 Nonmag 0.20 4.5 6.65 18.6 17 DMS #3 Mag 0.01 0.2 4.93 0.5 18 DMS #4 Floats Tailings 0.20 4.6 0.12 0.3 19 DMS #4 Sinks PP Feed 0.56 12.6 1.97 15.4

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Stream Stream ID Stream Wt Wt Grade Li Distribution # Type (t) (%) (% Li2O) (%) 20 DMS #4 Fines PP Feed 0.04 1.0 1.82 1.1 21 DMS #4 Slimes Tailings 0.01 0.1 1.82 0.2 Since four (4) stages of DMS may add capital cost to the processing plant, the mass balance was recalculated assuming only two (2) DMS stages. Summary mass balances for the two (2) scenarios are presented in Table 13.13.

Table 13.13 – Mine Representative Sample Summary Mass Balances – 2 and 4-Stage Scenarios 4-Stages Wt Wt Grade Li Distribution Combined Products (t) (%) (% Li2O) (%) Combined Tailings 1.77 40.0 0.42 10.4 Combined Middlings 2.15 48.7 1.48 44.9 Combined Concentrates 0.50 11.3 6.37 44.8 Feed 4.42 100.0 1.61 100.0

2-Stages Combined Products Wt Wt Grade Li Distribution (t) (%) (% Li2O) (%) Combined Tailings 1.50 34.0 0.41 8.6 Combined Middlings 2.62 59.4 1.78 65.7 Combined Concentrates 0.29 6.6 6.22 25.7 Feed 4.42 100.0 1.61 100.0 With the elimination of the final two (2) stages in the flow sheet, the production of final DMS concentrate would decrease from 11% (by weight) at a grade of 6.4% Li2O to 7% by weight at a grade of 6.2% Li2O. The lithium losses to the tailings would decrease by 2% and the grade of the middlings (i.e., flotation pilot plant feed) would increase from 1.5% to 1.7% Li2O. The lithium recovery to DMS concentrate would decrease from 45% to about 26% and the lithium deportment to the flotation circuit would increase from about 45% to 66%.

iv) Flotation Pilot Testing

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first in order to assess the effectiveness of various processing strategies. The processing route for the DMS middlings product obtained by processing the mine representative sample was based in part on the results of the larger blasted sample, the concentrate specification required by the potential purchaser of the concentrate, and the need to minimize capital cost. The trial results for this processing route were used to develop the design criteria for the processing plant for the PEA Study.

The main reason for producing a coarse DMS product was to obtain a combined DMS and flotation particle size distribution that would be acceptable to the potential concentrate buyer. Based on the particle size distributions obtained from the pilot DMS testing and preliminary flotation tests, it was determined that a two-stage DMS operation would provide a combined concentrate that would be sufficiently coarse to meet the specifications. Therefore, it was decided that the mine representative DMS middlings product obtained from a two-stage circuit was tested in the flotation pilot plant.

Preliminary pilot plant flotation trials confirmed that mica pre-flotation was well as wet high intensity magnetic separation of the flotation concentrate would be required to meet concentrate grade specifications.

Although it was determined that LR19 and Armac C are the collectors of choice for mica pre flotation and spodumene flotation, it is recommended that a cost-benefit analysis of substitute reagents be carried out prior to detailed engineering. b) Spodumene Recovery Summary

Based on DMS and flotation pilot testing at a head grade of 1.61% Li2O, the following was concluded:

• 25.5% of the total lithium was recovered in the DMS concentrate, while 65.7% of the lithium reported to the flotation circuit; • The flotation Li recovery was 80%, for an overall lithium recovery of 78%;

• The DMS concentrate grade was 6.2% Li2O, while the flotation concentrate grade was 5.9% Li2O. The combined concentrate grade was 6.0% Li2O; • The DMS concentrate represented 31.5% of the total concentrate and contained 32.7% of the total recovered lithium, while the flotation concentrate represented 68.5% of the total concentrate and contained 67.3% of the total recovered lithium.

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c) Sedimentation and Filtration Testing Sedimentation and filtration tests were carried out at SGS Lakefield in order to size the slimes tailings thickener and to select and size the appropriate equipment for filtering flotation tails and flotation concentrate. Based on the filtration tests, it was concluded that a pressure filter would be required to filter the flotation tailings. The flotation tailings will be combined with the DMS tailings and dry stacked. The combined moisture content would be about 10%. The flotation concentrate would be filtered by belt filter and combined with the DMS concentrate. The combined moisture content would be 5-6%. Further confirmatory tests were carried out by an equipment vendor. These results were used for the belt filter sizing and demonstrated that the use of steam was not required to produce a combined DMS and flotation concentrate at about 3% moisture to prevent freezing during transportation in the winter months. d) Process Flow sheet The proposed Process Flow sheet for this PEA Study is presented in Figure 13.13. The crushing circuit consists of one (1) jaw crusher along with primary and secondary cone crushers. The crushed product is sent to the DMS circuit which consists of two-stage of DMS. The feed size to the circuit is 9.5 mm. All -0.5 mm material is screened and later recombined with the middlings float product from the second DMS stage and sent to the grinding circuit. The grinding circuit consists of a ball mill and rod mill in series. The rod mill P80 is 200 µm. The rod mill product is sent to the flotation circuit, which comprises a conditioning and desliming stage, mica pre-flotation stage, and spodumene flotation stage.

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Figure 13.13 – Proposed Process Flow Sheet

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13.2 Hydrometallurgical Process

This section of the Report summarizes the pyrometallurgical and hydrometallurgical testing conducted on the Nemaska Lithium Whabouchi spodumene concentrate. This summary is based on reports provided by Feeco, SGS Lakefield and Ameridia.

The test work’s objectives were to develop a process flow sheet for the production of market-grade lithium hydroxide and lithium carbonate.

13.2.1 Introduction

SGS Lakefield was mandated by Nemaska Lithium to conduct a laboratory program for the development of the hydrometallurgical flow sheet. The flow sheet was piloted in three (3) phases: • Phase 1: Concentrate leach and all purification steps (primary and secondary impurity removal and ion exchange); • Phase 2: Lithium hydroxide membrane electrolysis and crystallization; • Phase 3: Lithium carbonate precipitation. Prior to the hydrometallurgical test work at SGS Lakefield, the Whabouchi spodumene concentrate was roasted at Feeco’s pilot plant facilities in Green Bay, Wisconsin, U.S.A.

In addition to SGS Lakefield’s electrolysis test work performed in Phase 2, electrodialysis bench tests have been completed at Ameridia’s laboratory located in New Jersey, U.S.A.

13.2.2 Pyrometallurgical Test Work

The spodumene roasting and acid roasting test work was conducted by Feeco in the fall of 2011.

a) Sample Preparation Fine and coarse concentrates from the original concentration flow sheet were shipped from SGS Lakefield to Feeco’s laboratory. The coarse material was generated in the dense media separation process while the fine material was produced through flotation.

Two (2) blends of concentrates were prepared using the as-received coarse material and the dried fine material. About 1.7 tonnes of a first blend was prepared, containing 75% fines and 25% coarse. A second blend (1.5 tonnes) was prepared, containing 50% fines and 50% coarse.

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b) Spodumene Conversion (Roasting) Roasting converts the spodumene’s crystalline structure from alpha to beta. The conversion occurs at a temperature of about 1,025 ºC. The beta phase is reactive with sulfuric acid and produces lithium sulphate, which is amenable to leaching.

The test work was carried out over a three (3) day period using a 30” diameter by 20’ feet long parallel flow gas-fired rotary kiln. The feed rate was maintained at 50 kg/h and a thermocouple was inserted in the discharge end hood to monitor the product’s temperature. c) Acid Roasting A paddle mixer was used to blend the beta spodumene with 93% sulfuric acid. The acidified material was returned to the kiln after the kiln had cooled to about 200 ºC. The kiln feed rate was between 135 and 180 kg/h. The retention time for the acid roasting was about 32 minutes. d) Pyrometallurgical Test Work Additional Information The following observations were made by Nemaska Lithium’s representative who witnessed Feeco’s test program. This information was not published by Feeco. • Only 82% extraction was achieved in the acid roasting step compared to the expected 95%. • Sulfuric acid was added in a 30% stoichiometric excess but this value was not supported by a chemical analysis. • The kiln temperature of 200 ºC was thought to be sufficient for the acid roasting step as it was expected that the exothermic reaction would raise the material’s temperature to 250 ºC. • Coarse alpha concentrate was easily converted to beta spodumene but required a longer residence time in the kiln. • Acid roasting was not affected by the material size. • The presence of iron in the concentrate affected the conversion temperature. A higher temperature could be reached without causing the formation of glass beads when the iron was removed from the concentrate. 13.2.3 Hydrometallurgical Test Work – Phase 1

The Phase 1 test program was carried out in November 2011 by SGS Lakefield. Phase 1 consisted of concentrate leaching, primary impurity removal, secondary impurity removal and ion exchange.

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a) Sample Preparation Approximately 2,600 kg of acid roasted beta spodumene (solid lithium sulphate) were processed during the first phase of the pilot plant.

The two (2) blends tested during the pyrometallurgy test work were used to feed the pilot plant (75/25 and 50/50). Table 13.14 shows the chemical analysis of the samples.

Table 13.14 – Phase 1 Pilot Plant Feed Analysis Sample Li Si Al Fe Na S Cr Zn Mn Mg Ca K (%) (%) (%) (%) (%) (%) (g/t) (g/t) (g/t) (g/t) (g/t) (g/t) 75/25 2.24 25.0 10.5 1.04 0.39 6.09 167 134 1962 1186 3431 3653 50/50 2.29 24.4 10.4 0.96 0.36 6.06 163 103 1755 905 2311 3376 b) Concentrate Leach and Primary Impurity Removal The objectives of the concentrate leach and the Primary Impurity Removal (“PIR”) were to dissolve the lithium sulphate from the acid roasted beta spodumene and remove the major impurities. Approximately 1,200 kg of 75/25 and 1,400 kg of 50/50 were processed in 85 hours.

In the concentrate leach step, solids were mixed with site water in a 50:50 ratio and agitated for 30 minutes. Lithium and contaminants (Fe, Al, Si, Mn and Mg) were leached and the final pH of the slurry was around 1.7. The lithium sulphate contained in the acid roasted beta spodumene was 100% leached.

In the PIR, the pH of the slurry was elevated to approximately 5.6 by adding hydrated lime. The major impurities (Fe, Al and Si) were precipitated as insoluble metal hydroxides. Air was sparged into the PIR tanks to maintain the oxidative potential of the slurry. The retention time in the PIR was 60 minutes.

The slurry coming from PIR was filtered on pan filters and the filtrate proceeded to Secondary Impurity Removal (“SIR”). Lithium extraction reached 79% for the 75/25 blend and 86% for the 50/50 blend.

Vacuum filtration testing was carried out on direct PIR discharge and thickened underflow. At 0.7 bar vacuum the direct discharge sample filter cake moisture was 26.7%. c) Secondary Impurity Removal The objectives of the secondary impurity removal were to precipitate Ca, Mg and Mn impurities from the PIR filtrate.

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The pH of the solution was increased to 10 by adding sodium hydroxide (NaOH) in the first two (2) tanks. Sodium carbonate (Na2CO3) was added in the third tank to convert all the remaining divalent impurities to insoluble carbonates. The total residence time for SIR was 240 minutes including the surge capacity.

Impurity levels at the discharge averaged 1 mg/L Mn, 14 mg/L Mg and 241 mg/L Ca in the pilot plant runs. Concentrations as low as 2 mg/L Mg and 200 mg/L Ca were attained by optimizing key parameters such as retention time. The overall lithium recovery was 99.1%. d) Ion Exchange The objective of the Ion Exchange (“IX”) circuit is to further reduce the calcium and magnesium tenors from the SIR discharge to 10 mg/L each.

The IX circuit consisted of three (3) columns packed with a cationic resin, which is selective towards divalent and trivalent metal ions.

The process consisted in a lead/lag/regeneration operation. At any time, two (2) columns would be removing Ca and Mg while the third one would be resin regeneration mode.

One cycle consists of the following steps: loading, feed wash, acid strip, acid wash, regeneration and regeneration wash.

The magnesium tenors varied between 0.07 and 0.2 mg/L and a 99.0% removal efficiency was achieved. The calcium tenors varied between 2.4 and 5.7 mg/L and a 97.6% removal efficiency was achieved. Lithium losses in the IX circuit were minimal at 2.7%.

13.2.4 Hydrometallurgical Test Work – Phase 2

The second phase of the test program was carried out in December 2011 by SGS Lakefield. Phase 2 consisted of membrane electrolysis test work. a) LiOH Membrane Electrolysis The objective of the electrolysis process is to produce a lithium hydroxide solution from a high purity lithium sulphate solution. The pilot plant was carried out in a three-compartment membrane electrolysis cell.

The central compartment was separated from the cathodic compartment by a cationic membrane and from the anodic compartment by an anionic membrane.

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Under the influence of an electric field (a 400-440 amps current was applied to the cell), lithium ions from the central compartment are transported through the cationic membrane. In parallel, the sulphate ions are moved to the anodic compartment. The anodic reaction generates protons and sulfuric acid is produced.

The feed solution composition is presented in Table 13.15.

Table 13.15 – Membrane Electrolysis Feed Solution Composition Tenor of Solution Components Sample Li Na K Ca Mg Fe Zn ID (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) 15,700 3,980 107 3.8 0.2 < 0.2 < 0.7 Ag Al As Ba Be Bi Cd (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) < 0.5 < 0.8 < 3 0.03 < 0.002 < 1 < 0.3 Co Cr Cu Mn Mo Ni P IX (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) Product < 0.3 < 0.2 < 0.1 < 0.04 < 0.6 < 1 < 5 Carboy 1 Pb Sb Se Sn Sr Ti Tl (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) < 2 < 1 < 3 < 2 0.61 < 0.1 < 3

U V W Y SO4 Cl (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) < 1 < 0.07 < 2 < 0.02 120,000 5 The pilot plant was run for two (2) five (5) days campaigns and samples were taken every four (4) to six (6) hours. Lithium hydroxide at a 14.6 g/L lithium concentration and a 20-30 g/L sulfuric acid solution were produced.

Overall the pilot plant was operated for 228 hours and 930.5 kWh of electrical energy was consumed. The current efficiency was 43.5% for the first pilot plant campaign and 34.9% for the second.

b) LiOH-H2O Crystallization The objective of the crystallization process is to produce high quality solid lithium hydroxide monohydrate from the lithium hydroxide solution generated through membrane electrolysis. Atmospheric and vacuum evaporation were tested in three (3) steps.

The water evaporation rate was between 6.0 and 6.8 mL/min during the atmospheric evaporation tests and the product ranged between 18.2 and 19.9% Li.

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The crystals were filtered, washed with distilled water and dried in an inert gas environment.

Vacuum evaporation was tested at two (2) different conditions: at 78-82 kPa (vacuum) and 62-70 ºC and also at 62-70 kPa (vacuum) and 79-81 ºC. The evaporation rate was approximately 7.7 mL/min and the final purity of the product was similar to the one that underwent atmospheric evaporation (17.9% Li).

13.2.5 Hydrometallurgical Test Work – Phase 3

Phase 3 of test program was carried out in March 2012 by SGS Lakefield and consisted of the production of lithium carbonate. a) Lithium Carbonate Production The objective of the test work is to prove that lithium hydroxide conversion to lithium carbonate can be an effective method of producing high purity lithium carbonate.

Lithium hydroxide solution produced from the membrane electrolysis pilot plant campaign was used as a feed for the lithium carbonate production. The feed had an average lithium grade of 14.7 g/L with calcium and magnesium concentrations of 3.5 mg/L and < 0.07 mg/L respectively.

The lithium carbonate production process consists of two stages, the first is referred to as Lithium hydroxide Carbonization (“LC”) and the second is called lithium bicarbonate decomposition (“DC”).

During carbonization, carbon dioxide gas is reacted with lithium hydroxide at room temperature to form lithium carbonate. Once all lithium hydroxide is carbonized, an excess of carbon dioxide converts lithium carbonate to lithium bicarbonate.

During decomposition, the solution is heated to near boiling. Lithium bicarbonate formed in the first stage is decomposed to insoluble lithium carbonate and carbon dioxide.

The lithium carbonate production pilot plant was run for three (3) days, 24 hours per day. At the end of the three (3) day pilot plant, 12.5 kg of lithium carbonate were produced (79% from the LC step and 21% from the DC step).

13.2.6 Electrodialysis Test Work

Electrodialysis using Bipolar Membranes (“EDBM”) was tested as an alternative method of producing lithium hydroxide. Ameridia was asked to conduct EDBM testing in June 2012.

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The objective of the test work was to confirm the suitability of the EDBM technology for Nemaska’s solution, measure current efficiency as well as resulting base and acid concentrations.

An EDBM stack consists of three (3) compartments: one (1) for the salt stream (Li2SO4), one (1) for the base recovery (LiOH) and one (1) for the acid recovery (H2SO4). When an electric field is applied, the cations migrate from the salt, through the cation membrane and into the base loop. Anions migrate through the anionic membrane and into the acid loop. The bipolar membrane splits water molecules. The H+ and OH- ions move to their respective compartment and allow the formation of the acidic and basic solutions.

Based on the results, EDBM appears to be a good technology for a large scale LiOH production facility.

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14.0 MINERAL RESOURCE ESTIMATES

This section is an extract of the 2011 Technical Report.

An initial mineral resource estimate has been previously reported on the Whabouchi Property by Nemaska in early summer 2010. Please refer to Section 6.0 History for the previous mineral resource estimate on the Project.

This section reports the results of an updated mineral resource estimate for the Whabouchi Project. SGS Geostat completed the mineral resource update using the digital database supplied by Nemaska which includes channel data from trenches and drill hole data completed by Nemaska since 2009. The historical data of the Project was not used in the present mineral resource estimate as the drill core is no longer available, which makes the analytical data impossible to validate.

The database used to produce the mineral resource estimate is derived from a total of 460 channels and diamond drill holes and contains the collar, survey, lithology, and analytical results information. A significant amount of channels does not have analytical data but include only lithological information which was considered during the modeling of the mineralized envelopes. The database contains drill hole data up to hole WHA-11- 117 but the analytical data from the last drill holes of the 2011 exploration program (WHA-11-115 to 117) was not included in the database because results were not available at the time of the mineral resource estimate. Refer to Table 12.3 for a summary of the records in the database used for the updated mineral resources estimate.

The mineral resource estimate is derived from a computerized resource block model. The construction of the block model starts with the modeling of 3-D wireframe envelopes of the mineralization using channel and drill hole Li2O analytical and lithological data for the updated mineral resource estimate, once the modeling is complete, the analytical data contained within the wireframe envelopes is normalized to generate fixed length analytical composites. The composite data is used to interpolate the grade of blocks regularly spaced on a defined grid that fills the 3-D wireframe envelopes. An optimized pit shell model using the pit optimization software Whittle was produced using the completed block model. The interpolated blocks located below the bedrock/overburden interface and within the optimized pit shell comprise the mineral resources. The blocks are then classified based on confidence level using proximity to composites, composite grade variance and mineralized solids geometry. The 3-D wireframe modeling, block model, Whittle optimized pit shell and mineral resource estimate were completed by SGS Geostat based on information provided by Nemaska.

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14.1 Exploratory Data Analysis

Exploratory data analysis for lithium was completed on original analytical data and composite data contained within the modeled mineralized envelopes.

14.1.1 Analytical Data

There are a total of 8,882 assay intervals in the database used for the current mineral resource estimate. Most of the drill hole intervals defining the mineralized envelopes have been sampled continuously. The few gaps with no analytical data located within the mineralized intervals were considered having zero grade for the purpose of the mineral resource estimate. These gaps generally correspond to local xenolites of adjacent lithologies floating inside the pegmatite intrusions. Table 14.2 shows the range of Li2O values from the analytical data.

Table 14.1 – Range of Li2O Analytical Data for Mineral Resource Estimation Li O 2 (%) Count 8,882 Mean 1.12 Std. Dev. 0.91 Min 0 Median 1.06 Max 4.59 The channel samples collected at Whabouchi are mostly located at the highest topographic area of the Property where the outcrop exposure is best. The channel’s azimuth ranges from N020° to N210° with an average of N149° which is generally perpendicular to the orientation of the pegmatite intrusions. The channels average 9.7 m in length and the sampling interval is typically one metre.

The core holes drilled on the Project are generally oriented N330°, perpendicular to the general orientation of the pegmatite intrusions, and have a weak to moderate deviation toward the east. Their spacing is typically 50 m with tighter 25 m spacing between sections 200 mE and 1,050 mE. The drill holes dips range from 43° to 75° with an average of 50° and the drill hole intercepts range from approximately 70% of true width to near true width of the mineralization.

14.1.2 Composite Data

Block model grade interpolation is conducted on composited analytical data. A 2-m composite length has been selected based on the N-S thickness of the 5 m by 3 m by 5 m block size defined for the resource block model. The minimum length of composite kept

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for the interpolation process is 1 m. Compositing is conducted at the start of the bedrock- overburden contact in the case of drill holes. No capping was applied on the analytical composite data. Table 14.2 shows the statistics of the analytical composites used for the interpolation of the resource block model and Figure 14.1 shows the related histogram for Li2O.

Figure 14.2 and Figure 14.3 display the spatial distribution of the composites in plan and longitudinal view respectively (hole collars are shown as blue circles and sample composites are shown as black diamonds).

Table 14.2 – Statistics for the 2-metre Composites for Li2O Li O 2 (%) Count 1,879 Min 0 Median 1.62 Max 3.72 Mean 1.62 Std. Dev. 0.586

Figure 14.1 – Histogram of 2-metre Composites for Li2O

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Figure 14.2 – Plan View Showing the Spatial Distribution of the Composites

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Figure 14.3 – Longitudinal View Showing the Distribution of the Composites (Looking North)

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14.1.3 Specific Gravity

Section 11.4 summarizes the SG determination in details. The results of the SG measurements conducted by SGS Geostat and Nemaska in 2010 and 2011 on selected mineralized core samples returned an average SG value of 2.70/m3 for the mineralized pegmatite. This value was used for the calculation of the tonnages from the volumetric estimates of the resource block model.

14.2 Geological Interpretation

SGS Geostat conducted the interpretation and modeling of the 3-D wireframe envelopes of the mineralization based on the channel and drill hole data in collaboration with Nemaska personnel. The modeling was first completed on sections to define mineralized prisms using the lithologies and analytical data for lithium. A minimum grade of 0.3% Li2O over a minimum drill hole interval length of 3 m was generally used as guideline to define the width of mineralized prisms, corresponding to the N-S width of the individual blocks. The final 3-D wireframe model was constructed by meshing the defined mineralized prisms based on the geological interpretation. Local smaller 3-D wireframe envelopes of significant size xenolith material located inside the large envelopes were also modeled.

A bedrock-overburden interface 3-D surface has been generated by triangulating the lower intercept of the overburden-coded lithology from the drill hole dataset. Figure 14.4 and Figure 14.5 show the final 3-D wireframe envelopes in plan and longitudinal view respectively. The different colors of the envelopes do not represent any specific parameters and are there to help the visual differentiation of the different envelopes.

Figure 14.4 – Final 3-D Wireframe Envelopes in Plan View

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Figure 14.5 – Final 3-D Wireframe Envelopes in Longitudinal View (Looking North)

14.3 Resource Block Modeling

A block size of 5 m (E-W) by 3 m (N-S) by 5 m (vertical) was selected for the resource block model of the Project based on drill hole spacing, width and general geometry of mineralization. The 5 m vertical dimension corresponds to an approximation for the bench height of a potential small open pit mining operation. The 5 m E-W dimension corresponds to about a quarter to a fifth of the minimum spacing between the drill holes and accounts for the variable geometry of the mineralization in that direction. The 3 m N-S dimension accounts for the average minimum width of the mineralization modeled at Whabouchi. The resource block model contains 177,997 blocks located below the overburden/bedrock surface for a total of 13,206,808 m3. Blocks located at the interface with the bedrock/overburden surface have been calculated with block fraction. Table 14.3 summarizes the parameters of the block model limits.

Table 14.3 – Resource Block Model Parameters Coordinates Block Number Direction (Local Grid) Size (m) of Blocks Min (m) Max (m) East-West (x) 5 301 0 1,500 North-South (y) 3 151 -350 100 Elevation (z) 5 121 -250 350

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14.4 Grade Interpolation Methodology

The grade interpolation for the Whabouchi resource block model was completed using the Inverse Distance to the power square (“ID2”) methodology. The interpolation process was conducted using three (3) successive passes with relaxed search conditions from one (1) pass to the next until all blocks are interpolated. In the first pass, the search ellipsoid distance was 75 m (long axis) by 75 m (intermediate axis) by 25 m (short axis) with an orientation of 0° azimuth, 75° dip and 0° spin which represents the general geometry of the pegmatites in the deposit. Search conditions were defined with a minimum of seven (7) composites and a maximum of 30 composites, with a minimum of three (3) holes required to estimate the block. The first pass resulted in the interpolation of 71% of the block model. For the second pass, the search distance was twice the search distance of the first pass and composites selection criteria were kept the same as for the first pass. A total of 95% of the blocks was interpolated during the second pass. Finally, the search distance of the third pass was increased to 500 m (long axis) by 500 m (intermediate axis) by 100 m (short axis) and again the same composites selection criteria were applied. The purpose of the last interpolation pass was to interpolate the remaining unestimated blocks mostly located at the edges of the block model. Figure 14.6 illustrates the three (3) search ellipsoids used for the different interpolation passes. Figure 14.7 and Figure 14.8 show the results of the block model interpolation in plan and longitudinal view respectively.

Figure 14.6 – View of the Search Ellipsoids Used for the Different Interpolation Passes

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Figure 14.7 – Block Model Interpolation Results in Plan View

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Figure 14.8 – Block Model Interpolation Results in Longitudinal View (Looking North)

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14.5 Mineral Resource Classification

The mineral resources at Whabouchi are classified into Measured, Indicated and Inferred categories. The mineral resource classification follows the CIM requirements and guidelines and is based on the density of analytical information, the grade variability and

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spatial continuity of mineralization. The mineral resources were classified in two (2) successive stages: automated classification followed by manual editing of final classification results.

The first classification stage is conducted by applying an automated classification process which selects around each block a minimum number of composite from a minimum number of holes located within a search ellipsoid of a given size and orientation. For the Measured resource category, the search ellipsoid is 35 m (strike) by 35 m (dip) by 5 m with a minimum of seven (7) composites in at least four different drill holes. For the Indicated category, the search ellipsoid is twice the size of the Measured category ellipsoid using the same composites selection criteria. The second classification stage involves the delineation of coherent zones for the Measured and Indicated categories based on the results of the automated classification. The objective is to homogenize or “smooth” the results of the automated process by removing the “Swiss cheese” or “spotted dog” patterns typical of the automated process results. The second stage is conducted by defining 3-D solids on a bench by bench basis for the Measured and Indicated categories. Figure 14.9 and Figure 14.10 show the block model classification in plan and longitudinal views respectively (categories: Measured – red, Indicated – blue, and Inferred – grey).

Figure 14.9 – Block Model Final Classification in Plan View

0 100 200 N metre

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Figure 14.10 – Block Model Final Classification in Longitudinal View (Looking North)

0 100 200 metre

14.6 Mineral Resource Estimation

The mineral resources of Whabouchi are reported using an open-pit mining perspective. Due to the significant depth extent of the resource block model, it is considered that not all the interpolated blocks could meet the requirement of reasonable prospect of economic extraction. In order to define the mineral resources of the Project, SGS Geostat completed an optimization of an open pit using the Whittle software. The final mineral resources include all the resource blocks located within the optimized pit shell and below the overburden/bedrock interface.

Table 14.4 summarizes the parameters used in the pit optimization process. The pit optimization process returned an in-pit Li2O grade cut-off of 0.36%. The base case cut- off for the updated mineral resource estimate has been set to 0.4% Li2O.

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Table 14.4 – Parameters Used for the Whittle Pit Optimization Parameters Value Unit Sales Revenue

Concentrate Price (6%) Li2O) 310 C$/tonne Operating Costs Mining Mineralized Material 2.76 C$/t milled Mining Overburden 2.76 C$/t milled Mining Waste 2.76 C$/t milled Crushing and Processing 13.00 C$/t milled General and Administration 1.60 C$/t milled Metallurgy and Royalty Concentration Recovery 80 % NSR Royalty 2 % Geotechnical Parameter Pit Slope 50 degrees

The final in-pit mineral resource estimate at the base case cut-off of 0.4% Li2O totals 11,294,000 tonnes, with an average grade of 1.58% Li2O in the Measured category, 13,785,000 tonnes, with an average grade of 1.50% Li2O and in the Indicated category, with an additional 4,401,000 tonnes, with an average grade of 1.51% Li2O the Inferred category. The updated mineral resource estimation for Whabouchi deposit is tabulated in Table 14.5 for the base case cut-off of 0.4% Li2O. Two (2) others cut-offs at 1.0% and 1.5% Li2O are also shown for reference. Figure 14.11 is a longitudinal section showing the optimized pit outline with the final in-pit resource block model.

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Table 14.5 – Whabouchi Deposit Updated Mineral Resource Estimate Cut-off Grade Li O Resources Li O Grade 2 Tonnes* 2 (%) Categories (%) Measured 11,294,000 1.58 Indicated 13,785,000 1.50 0.4% Measured + (Base Case) 25,079,000 1.54 Indicated Inferred 4,401,000 1.50 Measured 11,086,000 1.59 Indicated 12,903,000 1.55 1.0% Measured + 23,989,000 1.57 Indicated Inferred 4,245,000 1.53 Measured 7,299,000 1.72 Indicated 7,172,000 1.75 1.5% Measured + 14,471,000 1.73 Indicated Inferred 2,402,000 1.68 Note: The mineral resource estimate has been calculated using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definitions Standards for mineral resources in concordance with National Instrument 43-101 – Standards of Disclosure for Mineral Projects. Mineral resources which are not mineral reserves do not have demonstrated economic viability. Inferred mineral resources are exclusive of the Measured and Indicated resources. Bulk density of 2.70 t/m3 is used. Effective date June 6, 2011. * Rounded to the nearest thousand.

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Figure 14.11 – Long Section Showing Optimized Pit Outline and Final In-Pit Resource Block Model (Looking North Grid)

0 100 200

Metres

14.7 Block Model Validation

A validation of the mineral resource Li2O grade was conducted as part of the verification process. The validation includes: 1) a visual comparison of the color-coded block values versus the composites data in the vicinity of the interpolated blocks, and 2) a comparison of the grade average and standard deviation parameters for the original assay data located within the mineralized envelop, the composite data and the block model data. Table 14.6 summarizes the comparative statistics of the assay, composite and block model datasets.

Table 14.6 – Comparative Statistics for Assays, Composites and Block Model Datasets 2-m Assays* Block Model Composites Count 5,865 3,039 177,997 Mean 1.58 1.54 1.52 Std. dev. 0.77 0.66 0.29 Min 0 0 0.07 Median 1.59 1.57 1.53 Max 4.59 3.97 3.33 *Assays included in the mineralized intervals only

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14.8 Comments about the Mineral Resource Estimate

There are no known factors or issues related to permitting, legal, mineral title, taxation, socio-economic or political settings that could materially affect the mineral resource estimate.

The price of Li2O concentrate of US$310 per tonne selected for the pit optimization is based on the nine (9) months average selling price for Li2O concentrate disclosed by the Li2O concentrate producer Talison Lithium Ltd in their MD&A dated May 2011. SGS Geostat considers the selected price to be reasonable for the purpose of the current mineral resource estimate.

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15.0 MINERAL RESERVE ESTIMATE

Since this Report is a PEA Report, the resources have to be classified as in-pit measured and indicated resources and it is described in Section 16.0.

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16.0 MINING METHODS

16.1 Block Model Validation

The mining engineering work required for the study, such as the pit optimization, engineered pit design, mine planning, and in-depth economic analysis is based on the resource block model prepared by Geostat. The model has the name Whabouchi 2011_BM_ID2_final_8juin2011.csv. For the purpose of this study, the model was transferred from a Comma Separated Value File (CSV) into the MineSight™ mining software. Geostat also provided BBA with digitized topographical and bedrock mapping data.

The block model was provided in a local coordinate system based upon the section location and orientation used in geological interpretation.

The block dimensions and the block model details are as follows: • X-coordinate (abscissa-axis) = 5 metres; • Y-coordinate (ordinate-axis) = 3 metres; • Z-coordinate (vertical axis) = 5 metres. The following data was provided by Geostat in the model: • ix, iy, iz (increments in model); • Fixed Density (Rock = 3.06 t/m3, ROM = 2.70 t/m3); • Classification (3 = Measured, 2 = Indicated, 1 = Inferred);

• Li2O % (Lithium Oxide %); • Percent_Env (only for the bedrock-overburden interface, fraction of block considered mineralization).

Additional variables were introduced into the BBA MineSight model, in order to ascertain additional block model economic statistics. The additional variables are summarized in Section 16.3.2 below.

16.2 Pit Optimization

In order to develop an optimal engineered pit design for Nemaska’s Whabouchi Project, an optimized pit shell was prepared using the Lerchs-Grossman 3D algorithm in MineSight™ (“LG 3D”). The LG 3D pit optimizer algorithm is a true pit optimizer, based upon dynamic programming of the graph theory. The pit optimizer calculates the net value of all blocks in the model (i.e. profit minus loss) and searches for the ultimate pit shell that delineates the volume of extraction, which maximizes the revenue. In order

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to maximize the revenue, it takes into consideration the following: mining costs, processing costs, processing recoveries, weight recovery values, and the overall pit slope.

As outlined in this section, the pit optimization has used only those blocks classified as either a Measured or Indicated resource in order to generate revenue.

16.2.1 Pit Optimization Parameters

Additional variables were coded into the block model in order to ascertain block model economics.

The main economic equations that govern the outcome of the LG 3D algorithm are as stated: • Volume of block (“Vol.”) (m3) = 5 m × 3 m × 5 m • Head Grade in Li (“HGLI”) = 〖%Li〗_2 O × 0.465 (atomic mass) • Li Recovery (“LIREC”) = (HGLI/100) × (79/100) • Spodumene Produced (“SPODP”) = (LIREC × 100) /2.81 (Tonnes) • Spodumene Sold (“SPODS”) = (SPODP) × (1-0.02) (Tonnes) • Block Value (“BVAL”) = {Vol. × Density × (TOPO %) /100)} × (400 SPODS-122 SPODP) * ($) *Where value of concentrate FOB mine = $400/t Spodumene Concentrate, less Freight cost mine to refinery = $122/t con.

Additional pit optimization parameters include: • Mining cost (mineralized material/waste) = $2.50/t mined • Total operating cost per tonne of mineralized material = $24.00/ t milled (Processing, G&A) • Pit Slope = 48° 16.2.2 Cut-off Grade Calculation

A milling cut-off grade (“CoG”) is used to classify the material inside the pit limits as mineralization or waste. Due to the fact that the material falls within the pit, the breakeven cut-off grade is the grade required to cover the costs for processing and general and administration costs (“G&A”).

After multiple simulations on in-pit resources estimation, the mill cut-off grade for the Nemaska Lithium “Whabouchi” Pit was established at 0.4% Li2O.

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16.2.3 Pit Optimization Results

The technical parameters described in Section 16.2.1were used in the LG 3-D algorithm to generate the optimum pit shell for Nemaska Lithium’s Whabouchi pit. Figure 16.1 shows a 2D view of the pit optimization.

Figure 16.1 – 2D LG Pit Shell

16.3 Engineered Pit Design

The detailed pit design for Nemaska Lithium’s Whabouchi Project was carried out using the optimal LG 3D pit shell as a guide. The designed pit design includes mining features required for an operational open-pit mine, such as minimum mining widths, safety berms, bench face angles, inter-ramp angles, and benching arrangement. Bench heights of ten (10) m have been used, and double benching is used throughout the final designed pit.

The geotechnical parameters used for the detailed pit design were prepared by external consultants, Journeaux Assoc., and are presented in the following section.

16.3.1 Pit Slope Parameters

As previously mentioned in Section 16.3 Engineered Pit Design, the geotechnical parameters for the pit design were provided by Journeaux Assoc. in a report entitled “Report on Pit Slope Design”, dated February 16, 2012.

The majority of the recommended pit slope design parameters are listed in Table 16.1. The original recommendations suggest bench heights of 10 m, with a bench face angle of 75° and inter-ramp angle of 56°, resulting in 4-m berms.

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Table 16.1 – Recommended Slope Angles for Whabouchi Project Bench Berm Vertical Bench Wall BFA (°)1 IRA (°)1 Height (m) Width (m) Separation (m) North Wall 10 4 75 10 56 South Wall 10 4 75 10 56 East End Wall 10 4 75 10 56 West End Wall 10 4 75 10 56 1 “BFA” = Bench Face Angle and “IRA”= Inter-Ramp Angle. According to Journeaux Assoc., the slopes are considered stable due to very competent bedrock and to the fact that the joint systems are deeply steeping. The conclusions and recommendations can be found in the report from Journeaux Assoc. entitled Report on Pit Slope Design, Whabouchi Project, Nemaska Lithium Report No. L111474, March 23, 2012.

16.3.2 Additional Design Aspects

The in-pit haulage roads for the PEA Study of Nemaska Lithium are 22 m wide in order to accommodate two-way traffic for the 46-tonne trucks. A single-lane, 16 m-wide ramp is used for the lower levels of the mine. All in-pit ramps have been restricted to a maximum gradient of 10%. The ramp exit has been designed to allow easy access to the waste rock piles situated to the North (UTM) of the pit. Both the 2D and 3D representations can be seen in Figure 16.2 and Figure 16.3, respectively.

The engineered pit design dimensions are: • Length: 1,250 m; • Width: 320 m; • Depth: 190 m. Three (3) cross-section views (Eastings) are also Xdemonstrated from Figure 16.4 to Figure 16.6.

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Figure 16.2 – 2D Pit Design

N N

Figure 16.3 – 3D Pit Design and LG Pit Shell with Resources (≥ 0.4 % Li2O)

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Figure 16.4 – Cross-Section View Pit Design, East 437.5 m

Figure 16.5 – Cross-Section View Pit Design, East 752.5 m

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Figure 16.6 – Cross-Section View Pit Design, East 972.5 m

16.4 Dilution and Mineralized Material Loss

Using the provided block model and the mining polygon method, BBA performed calculations in order to determine the dilution and mining recovery for the engineered pit design. The resulting dilution was 4.5% at 0.34% Li2O, and the mineralized material loss was 4.5%.

The mining polygon method for dilution estimation simulates the possible mining configuration by digitizing a series of mining polygons around the mineralized material blocks that would be delivered to the mill as mill feed. The blocks that are being delivered to the mill are those that are classified as either Measured or Indicated resource class material, with a grade of 0.4% Li2O or greater. Therefore, these are the blocks that the mining polygon method triggers. The method follows a set of guidelines to ensure that the work on the dilution estimate is consistent and systematic throughout all of the chosen benches for the estimation.

The guidelines for digitizing the polygons in order to estimate the expected dilution are as follows: • Mining dilution simulation should only include Measured and Indicated resource classes; • Mining polygons must only be taken for mineralized material blocks having a certain pre-determined cut-off grade; February 2013 QPF-009-12/B

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• When mining widths were greater than 4 metres (i.e., ≥ 4 m in width), the polygon perimeters were digitized with an offset of 0.3 m inside the mineralized material /waste contact of the block; • When mining widths were less than 4 metres (i.e., ≤4 m in width), the polygon perimeters were digitized with an offset of 0.5 m outside the mineralized material /waste contact of the block; • Minimum blocks to be mined in a single cut should be two blocks in length; • Mining cuts were taken only within the pit design. Figure 16.7 shows an example of the polygon method used in order to estimate dilution in the model. Additional waste is added and a portion of mineralized material is removed, which contributes to the final diluted mineralized material amount. The dilution content is calculated as the ratio of the waste tonnes along the perimeter of the mining block divided by the tonnes of mineralized material located inside the mining block.

Figure 16.7 – Example of Mining Dilution Polygon

+0.5m -0.3m

The polygon estimates were performed on four (4) different benches equally spaced in the pit design: z = 297.5 m, z = 257.5 m, z = 212.5 m, z = 167.5 m. The plan views for these respective benches are shown in Figure 16.8 through Figure 16.11. An isometric view of the selected benches, shown with the engineered pit design, can be seen in Figure 16.12.

Figure 16.8 – Plan View of Mining Polygon, Bench 297.5 m

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Figure 16.9 – Plan View of Mining Polygon, Bench 257.5 m

Figure 16.10 – Plan View of Mining Polygon, Bench 212.5 m

Figure 16.11 – Plan View of Mining Polygon, Bench 167.5 m

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Figure 16.12 – Isometric View of 4 Benches Selected to Perform Mining Polygon Method

After completion of the polygons, they were extruded by one (1) block height (5 m) in order to generate the solids for estimating the mining dilution, dilution grade and mining loss.

With all the mineralized material, waste and grade data extracted from the computation of the solids, a weighted average was taken from the benches. The compilation of results can be seen in Table 16.2.

Table 16.2 – Summary of Dilution Results Mineralized Dilution Bench material Dilution Grade Li O Loss 2 (%) (%) (%) 297.50 5.00 5.29 0 257.50 5.22 3.46 0.41 212.50 4.36 3.89 0.39 167.50 4.02 4.42 0.44 Average 4.50 4.50 0.34 16.5 In-Pit Resources

The Resources for the engineered pit design amount to 19.639 Mt of Measured and Indicated Resources at an average grade of 1.49% Li2O using a cut-off grade of 0.4% Li2O after 4.5% dilution @ 0.34% Li2O and an mineralized material loss of 4.5%.The expected mine life is approximately 19 years, based on a production rate of 1.095 Mt of ROM per year. The stripping amounts to 59.4 Mt, resulting in an overall stripping ratio of 3.02 tonnes waste per tonne mineralized material. The resources contained in the engineered pit design are given in Table 16.3 by resource categories.

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Table 16.3 – Final In-Pit Resources Nemaska PEA Study (PEA) (4.5% Dilution @ 0.34%Li2O, 4.5% Mineralized Material Loss) Total Resources Estimate – (Cog 0.4% Li2O) Mineralized Li O Category 2 Material (kt) (%) Measured 10,197 1.530 Indicated 9,442 1.455 Total 19,639 1.49 Waste (kt) Inferred 377 Rock 56,646 OB 2,356 Total Stripping 59,379 Total SR 3.02 16.6 Mine Planning

The yearly mining schedule has been developed based on a mill feed of 1,095,000 tpy, or 3,000 tpd. The life of the mine of this Project is approximately 19 years, based on the 19,639 M tonnes of Measured and Indicated Resources from the engineered pit design for Nemaska’s Whabouchi Project.

A starter pit was created in order to ensure the extraction of higher grade run-of-mine (“ROM”), along with the minimization of the stripping ratio in the first three (3) years of the life of mine. The starter pit that was used in the phasing of the mine plan can be seen in the figure below. To achieve these goals during the pre-production period, 1.14 Mt of waste material must be stripped from the engineered pit design, and 0.43 Mt of overburden must be stripped.

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Figure 16.13 – Starter Pit Shown Against Engineered Pit Design

A weighted average ramp-up of 77% is applied to the run of mine and stripping amounts in Year 1. Two (2) months’ worth of production at a mill capacity of 30% is also used in preproduction (“PP”), which totals 54,000 tonnes of ROM.

The mining sequence presented in Table 16.4 below shows the details of the 19 years of the mine life, which was used for the financial analysis. The mine plan drawings can be found in the following Figure 16.14 through Figure 16.23.

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Table 16.4 – Nemaska Lithium Final Mine Plan Nemaska Lithium PEA Study Mine Plan

Dilution= 4.5% at 0.34% Li2O, Mineralized Material Loss=4.5% Mill Minerali Mill Stockpile Reclaim Total Period Stockpile Reclaim OB Rock Inferred Strip Moved zed Grade Grade Grade Stripping Material

(Year) (Mt) (Li2O %) (Mt) (Li2O %) (Mt) (Li2O %) (Mt) (Mt) (Mt) Ratio (Mt) (Mt) YPP 0.054 1.460 0.051 1.460 0.43 1.14 0.00 -- 0.54 1.67 Y1 0.840 1.528 0.26 1.25 0.00 1.79 1.50 2.34 Y2 1.095 1.514 0.21 1.90 0.00 1.93 2.12 3.21 Y3 1.095 1.542 0.23 2.09 0.00 2.12 2.32 3.41 Y4 1.095 1.454 0.02 2.57 0.01 2.38 2.60 3.70 Y5 1.095 1.426 0.04 2.95 0.00 2.72 2.98 4.08 Y6 1.095 1.483 0.45 2.70 0.00 2.88 3.15 4.25 Y7 1.095 1.542 0.39 3.63 0.00 3.67 4.02 5.11 Y8 1.095 1.576 0.23 4.62 0.00 4.43 4.85 5.95 Y9 1.095 1.533 0.11 4.73 0.01 4.43 4.85 5.94 Y10 1.095 1.437 5.87 0.01 5.37 5.88 6.97 Y11 1.095 1.452 4.85 0.04 4.46 4.89 5.98 Y12 1.095 1.454 4.54 0.04 4.18 4.58 5.67 Y13 1.095 1.468 4.26 0.04 3.93 4.30 5.39 Y14 1.095 1.444 3.70 0.09 3.46 3.79 4.89 Y15 1.095 1.483 2.56 0.06 2.40 2.62 3.72 Y16 1.095 1.495 1.72 0.04 1.61 1.77 2.86 Y17 1.095 1.530 0.88 0.00 0.80 0.88 1.98 Y18 1.095 1.533 0.66 0.00 0.61 0.67 1.76 Y19 0.130 1.502 0.051 1.460 0.04 0.48 0.04 0.12 Total 19.64 1.49 0.05 1.46 0.05 1.46 2.36 56.65 0.38 3.02 59.38 79.02

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Figure 16.14 – Plan View Mine Plan Pre-Production

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Figure 16.15 – Plan View Mine Plan Year 2

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Figure 16.16 – Plan View Mine Plan Year 4

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Figure 16.17 – Plan View Mine Plan Year 7

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Figure 16.18 – Plan View Mine Plan Year 9

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Figure 16.19 – Plan View Mine Plan Year 11

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Figure 16.20 – Plan View Mine Plan Year 13

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Figure 16.21 – Plan View Mine Plan Year 15

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Figure 16.22 – Plan View Mine Plan Year 17

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Figure 16.23 – Plan View Mine Plan Year 18

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16.7 Waste Rock and Tailings Management

The Waste Rock Pile is located to the north of the Nemaska engineered pit. The Waste Rock Pile is divided into two (2) phases: Phase (A) is the first phase to be filled, and is located on the west side of the total footprint; Phase (B) is the second phase to be filled, and is located on the east side of the footprint. The reason for the division of the Waste Rock Pile into two (2) phases is due to the fact that the Route du Nord is running through the center of the combined (overall footprint). The relocation is delayed until the completion of Phase (A). Phase (A) expires around Year 12, once a capacity of 13.2 Mm3 (bank volume) has been reached. Afterwards, Phase (B) is used until the end of the Life of Mine (“LOM”). The capacity of Phase (B) is 14.5 Mm3 (bank volume). This yields a total Waste Rock Pile capacity of 27.7 Mm3 (bank).

The Project has selected to use a co-disposal methodology for the tailings produced at the concentrator and the waste rock from the mine. Co-disposal is the mixing of fine and coarse mine waste to produce a single waste stream. Mixing the fine and coarse waste reduces the empty void space primarily associated with coarse waste streams, while simultaneously increasing the strength of the fines. Tailings produced at the concentrator have moisture content around 10%.

The deposition strategy to blend the coarse and fine waste feeds is based on mixing at the edge of the waste rock pile by placing both tailings and waste rock near the crest of the active dump, then pushing both over the face with a dozer. The design parameters used for the waste/tailings stockpile do not require that a fixed ratio be maintained. Waste and tailings trucks can arrive randomly. Compaction will be achieved by traffic of equipment on the pile and self-weight.

Thus, the tailings are being transported from the concentrator to the rock pile and mixed at the crest with a dozer. The dimensions of the CAT 772 (or equivalent), recommended for the hauling fleet, were used for the design criteria of the loading area at the concentrator. Loading of the tailings is being done by a suspended conveyor with a single discharge point and with a design capacity to load the CAT 772 (or equivalent) in ten (10) minutes.

Based on the engineered pit design, the amount of in-pit waste rock is approximately 18.6 Mm3 (in-pit), leaving sufficient space for up to 9.1 Mm3 for the coarse tailings material from the mill.

The design criteria used for both phases of the Waste Rock Pile are governed by geotechnical specifications. The geotechnical specifications are as follows: • Bench Face Angle (“BFA”): 30°; • Inter-Ramp Angle (“IRA”): 26.6°;

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• Maximum Height of Pile: 70 m; • Swell Factor: 30%; • Bench Height: 10 m; • Ramp Width: 22 m; • Ramp Grade: 10%; • Waste Rock Pile split into two (2) phases: Phase (A), Phase (B), see Figure 16.24. 16.8 Mine Operations

Mining operations will be conducted 24 hours day, seven (7) days week and 365 days per year. The operations at Nemaska will use conventional mining methods, which include a drilling and blasting sequence, followed by hauling and loading.

16.8.1 Drilling

Production drilling will be accomplished using a fleet of diesel drilling rigs. The mineralized material and waste zones will be drilled with 6 ½ inch diameter holes, 5 m spacing, and 5 m burden.

A re-drill factor of 5% has been included to account for productivity lost to collapsed holes or lost drill steels. The number of drills operating at any given time is dependent on the annual production rate and varies over the course of the mine life. The operating drill requirements are to have one drill operating throughout the life of mine. However, the fleet has allocated two (2) drills (one on stand-by), in order to account for any unforeseen downtime.

Holes will be drilled to a total depth of 11.5 m including 1.5 m of sub-drilling. A stemming height of 4 m will be used to maximize the effectiveness of the explosive column.

With all of these factors, the tonnage capacity per shift is respectively 11,267 tonnes of mineralized material and 12,770 tonnes of waste.

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Figure 16.24 – Waste Rock Piles (Phases A and B)

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Table 16.5, shown below, demonstrates the detailed drilling parameters, as well as assumptions made about the mineralized material and waste material.

Table 16.5 – Drilling Parameters and Assumptions Parameter Mineralized Waste Units material Hole Diameter 6 1/2 6 1/2 inches Hole Diameter 165.0 165.0 mm Bench Height 10.0 10.0 m Subdrill Length 1.5 1.5 m Insitu Bulk Density 2.70 3.06 t/m3

Hole Spacing 5.0 5.0 m Burden 5.0 5.0 m Rock Mass per Hole 675 765 tonnes/hole

Penetration Rate 25 25 m/hr Shift Drill Time 8.06 8.06 hr Metres/Shift 201.56 201.56 m Redrill 5% 5% % Holes/Shift 16.69 16.69 holes

Drilling Capability 11,267 12,770 Tonnes/shift 16.8.2 Blasting

Blasting will be provided under a contract with an explosive company that will supply blasting materials and technology, and ensure the delivery and storage of explosive products. Blasting will be accomplished using 100% emulsion type explosive production with an average density (in the hole) of 1.25 g/cm3.

Based on the drilling patterns listed above, the powder factor is estimated to be 0.297 kg/tonne in mineralized material and 0.262 kg/tonne in waste. The explosives will be trucked and stored on-site by the explosives supplier in storage facilities built for this purpose. The explosives contractor will also be responsible for providing a down-the-hole service.

Table 16.6 below shows the blasting parameters used that contribute to the production calculations.

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Table 16.6 – Blasting Parameters Parameter Mineralized Waste Units Material Hole Diameter 165.0 165.0 mm Bench Height 10.0 10.0 m Subdrill Length 1.5 1.5 m Stemming Length 4.0 4.0 m Loaded Length 7.5 7.5 m m3 explosive/ linear Volume/m 0.0214 0.0214 metre drilled Rock Mass Per Hole 675 765 tonnes/hole Bulk Emulsion Usage 100.0% 100.0% % Density 1.25 1.25 gm/cc Kg / Hole 200.5 200.5 kg Explosive Density 1.25 1.25 gm/cc Powder Factor 0.297 0.262 kg/tonne The cost for the blasting contractors is included in the summary of mine operating costs. Costs for a blasting contractor supervisor, Mobile Manufacturing Unit (“MMU”) operator otherwise known as the bulk emulsion truck operator, and mechanic are all summarized in the section on operating costs.

Emulsion-type explosives are preferred for their higher resilience to water. Environmentally, they are also sounder and produce 2% less residual ammonia than other explosive materials. In the size range chosen for the charge diameter of these emulsive products, these explosives will allow for better overall performance and handling.

In order to obtain good fragmentation, electronic detonators will be used. This type of detonator is a less costly alternative. The total cost per tonne (including explosives manufacturing, transport, down-the-hole service, and related labour fees) has been estimated at $0.341 per tonne of mineralized material and $0.319 per tonne of waste.

16.8.3 Loading and Hauling

Production will be accomplished using a fleet of 46-tonne capacity haul trucks and hydraulic shovels with a 6 m3 bucket capacity. This fleet combination should allow for 4-pass loading of trucks hauling mineralized material and waste, and 5-pass loading for trucks carrying overburden.

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Haul truck loading time is estimated at 2.7 minutes for mineralized material and waste and 3.3 minutes for overburden.

16.9 Fleet Requirements

The requirements for the primary mining equipment include a fleet of drills; haul trucks, shovels, and one (1) supporting wheel loader, which are based on the haul distances, equipment availability, utilization, and overall productivity data. Availability profiles for major equipment have been estimated using vendor benchmarks, similar projects in BBA’s database, as well as internal operational experience.

The truck fleet consists of units having a payload capacity of 46 tonnes (CAT 772 or equivalent). A maximum of seven (7) trucks will be necessary to support the mine productivity level of 1.095 Mtpy of ROM in Year 8, along with the waste removal schedule. In addition, tailings re-handling requires an operating fleet of one (1) truck to satisfy the variable amounts of tailings re-handling throughout the life of mine. This one (1) truck is summarized separately from the total operating cash costs, since it is not a cost directly associated with the mine planned tonnage. (The description for tailings is summarized further in the section on Waste Rock and Tailings Management). Operating truck requirements were determined using the appropriate operating time parameters, fill factors, haulage distances and cycle times, and tonnes to be moved by material type.

The truck fleet size was calculated using a mechanical availability of 88% in the earlier years, transitioning to 83% in the later years. The equipment utilization factor is 95%.The gross operating hours (“GOH”) used for the operating truck calculation, are based on a 2 x 12-hour shift work schedule and upon the variable mechanical availability. The GOH varies from 10 hours to 10.6 hours over the life of mine. The truck fleet size on a yearly basis was then smoothed out to better represent actual operation.

The proposed hydraulic shovel fleet consists of 6 m3 front-end configured units (CAT 390D or equivalent) for loading the blasted ROM, waste rock, and overburden. The tailings re-handling does not require a loader or excavator, since the truck will be filled directly from the conveyor connected to the concentrator.

Annual shovel productivity was determined using the appropriate operating time parameters, fill factors, material properties, and bucket capacities. The number of shovels will reach a maximum of two (2) units in Year 6. In addition, one (1) 4-m3 wheel loader (CAT988H or equivalent) will assist in the loading of ROM in cases where the main loading equipment is temporarily unavailable or when temporary additional loading production is required. The wheel loader will also assist in other general duties such as stockpile management. The fast response time and high mobility of the wheel loader will enable particular bench face blending requirements.

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As mentioned in Section 16.3.1 on the drilling parameters, one (1) operating drill is required to satisfy production. In addition, a second drill has been added for stand-by use. The drill is a DTH Cubex QXR 920 (or equivalent), which will be drilling 6½" diameter holes.

For a complete list of all primary, secondary and auxiliary equipment, see Table 16.7.

16.10 Mine Manpower Requirements

The manpower requirements for the mine are divided into two (2) categories: hourly operations and staff personnel. Hourly operations personnel can be further subdivided into the major equipment and support equipment operators and the maintenance personnel. Most of the operators for the major mine equipment are based on a four (4) crew rotation schedule.

The number of hourly personnel (mine operations and mobile equipment maintenance) reaches a peak of 73 in Year 10. The average ratio of hourly maintenance personnel to hourly operators is 0.48.

The staff mine personnel list includes such employees as mine superintendent, engineers, planners, foremen, and geologists. The number of salaried employees reaches a maximum of 27 in Year 5 and subsequently decreases to 17 in Year 15 for the short remainder of the mine life.

A full list of the personnel over the life of the mine can be seen in the chart shown below.

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Table 16.7 – Complete Mining Fleet Equipment Type PP Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10 Yr 11 Yr 12 Yr 13 Yr 14 Yr 15 Yr 16 Yr 17 Yr 18 Yr 19 Hydraulic Excavator (CAT390D- 6-m³) 1 1 1 1 1 2 2 2 2 2 2 2 2 2 1 1 1 1 0 Wheel Loader (CAT988-H) 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Haul Truck (CAT772-51 ton-46 tonnes) 2 3 4 5 5 6 6 7 8 8 8 8 8 8 8 7 6 5 4 2 Drill DTH (Cubex QXR 920 – 6”) 1 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 1 Total Primary Equipment 4 7 8 9 9 10 11 12 13 13 13 13 13 13 13 11 10 9 8 4 Wheel Dozer 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 0 0 0 0 0 Track Dozer (D7) 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 0 0 0 Motor Grader (Caterpillar 14 M) 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Water Truck 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Total Field and Shop Fleet 8 11 12 13 13 14 15 16 17 17 17 17 17 17 17 14 13 11 10 6 Fuel/ Lube Truck 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Service Truck ( 250 HP 22,000 GVW) 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Tire Changer (attachment for 988-H) 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 0 0 0 Pick Up Truck (4 x 4 crew cab Chevrolet 2500) 2 4 4 4 4 4 4 4 4 4 4 4 4 4 4 2 2 1 1 1 Pick Up Truck (4 x 4 single cab Chevrolet 2500) 2 4 4 4 4 4 4 4 4 4 4 4 4 4 4 2 2 1 1 1 Light Plant (1,000 W diesel generator) 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 0 0 Mobile Pump (125 HP diesel) 1 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 1 1 1 Total Auxiliary Equipment 9 14 14 14 14 14 14 14 14 14 14 14 14 14 14 10 10 6 5 5 Total Mine Equipment 17 25 26 27 27 28 29 30 31 31 31 31 31 31 31 24 23 17 15 11

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Figure 16.25 – Personnel Trend over LOM

120 Mining Personnel Trend

100

80 Personnel

60 Total Personnel Mining Salaried Personnel Total of

Hourly Personnel Total

Number 40

0 PP Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19 Year

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17.0 RECOVERY METHODS

Section 13 of this Report described the metallurgical test work and how the results were used to derive the preliminary Process Flow sheet (“PF”) and mass balance. The process design has been split in two (2) locations. The Concentrator Plant will be located at Whabouchi and the Hydrometallurgical Plant will be located at Valleyfield near the CEZinc facilities. The Concentrator design is presented in Section 17.1 whereas the Hydrometallurgical Plant design is presented in Section 17.2.

General processing plant descriptions of various areas are provided. This information serves as input information for the development of the capital and operating cost estimates presented in Section 21.

17.1 Concentrator Plant

The PF and mass balance constitute the basis for determining the recovery method of spodumene from the Whabouchi mineralized material to obtain a spodumene concentrate which be treated at Valleyfield to produce lithium hydroxide and lithium carbonate of suitable grade to produce battery grade lithium compounds. In this section of the Report, a process design basis is established and is used to further develop the PF, mass and water balance and to develop the process design criteria, including the selection and sizing of major process equipment.

17.1.1 Process Design Criteria

The process design is based on the processing plant capacity of 1.1 MTPY. The test results, given in Section 13, were used to develop the PF and preliminary mass and water balances which form the base criteria for the process design. In addition, the required size and quantities of major processing equipment were calculated based on the aforementioned criteria, briefly summarized in Table 17.1.

Table 17.1 – Process Design Basis Parameters Comment Unit Value All flow rates are expressed as nominal value Production Parameters Annual Mineralized Material Processed Dry Solid tpy 1,095,000 Annual Production DMS Concentrate Dry Solid tpy 72,880

DMS Concentrate Grade % Li2O 6.2 Flotation Concentrate Dry Solid tpy 157,225

Flotation Concentrate Grade % Li2O 5.9

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Parameters Comment Unit Value DMS Concentrate as % of Total Concentrate % 31.7 Flotation Concentrate as % of Total % 68.3 Concentrate. Concentrate Moisture % 5.1 Total Concentrate Production Rate Dry solid tpd 630.4 Total Concentrate Production Rate Dry solid tph 26.3 DMS Circuit Li Recovery* % 93.9 DMS Circuit Concentrate Weight Yield % 6.7 Li Deportment to DMS Concentrate % 25.7 Li Deportment to DMS Middlings and Fine to % 68.3 Flotation DMS Middlings and Fines to Flotation Dry Solid tpy 699,099

Flotation Feed Grade % Li2O 1.7 Overall Flotation Circuit Li Recovery % 77.4 Flotation Circuit Weight Yield as % of Total % 14.4 Feed Li Deportment to Flotation Concentrate % 52.8 Overall Li Recovery % 78.5 Overall Weight Yield % 21.0 Overall Concentrate Grade % 6.0 * - Recovery is defined as Li in DMS concentrate + Li in middling and fine divided by total Li input

Operating Schedule Plant Operating Days per Week day 7 Plant Operating Hours per Day hour 24 Equipment Availability % 90 Plant Availability % 90 Crushing Plant Operating Days per Week day 7 Crushing Plant Operating Hours per Day hour 24 Crushing Equipment Availability % 75 Crushing Plant Availability % 75

Plant Design Basis

Mineralized Material Feed Grade % Li2O 1.6 Annual Mineralized Material Feed Rate to the Mill Dry Solid tpy 1,095,000 Mineralized Material Bulk Density g/cm3 1.7

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Parameters Comment Unit Value DMS Concentrate Bulk Density g/cm3 1.9 Flotation Concentrate Bulk Density g/cm3 1.7 DMS Tailings Bulk Density g/cm3 1.9 Flotation Circuit Tailings Bulk Density g/cm3 1.6

Mill Solid Annual Feed rate Dry Solid tpy 1,095,000 Daily Mill Throughput (Annual Average) Dry Solid mtpd 3,000 Hourly Throughput (Annual Average) Dry Solid mtph 125 Daily Mill Throughput (Design) Dry Solid mtpd 3,333 Hourly Throughput (Design) Dry Solid mtph 139

Crusher Solid Annual Feed rate Dry Solid tpy 1,095,000 Daily Crusher Throughput (Annual Average) Dry Solid mtpd 3,000 Hourly Throughput (Annual Average) Dry Solid mtph 125 Daily Crusher Throughput (Design) Dry Solid mtpd 5,000 Hourly Throughput (Design) Dry Solid mtph 208 Mineralized Material Moisture % 3.0

Grinding Circuit Feed to Rod Mill Dry Solid tph 48 Feed to Rod Mill Dry Solid tpy 417,196

F80 um 6,500

P80 um 375

Feed to Ball Mill (Fresh Feed) Dry Solid tph 80 Feed to Ball Mill (Fresh Feed) Dry Solid tpy 627,389

F80 um 348

P80 um 200

Flotation Circuit Flotation Feed Dry Solid tpy 627,389 Flotation Feed Dry Solid tph 80

Head Grade % Li2O 1.7

F80 um 200 Tailings - Total Dry Solid tpy 865,480 Tailings - Coarse DMS Solid Flow Rate Dry Solid tpy 325,591

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Parameters Comment Unit Value % of Total Tails % 37.6 Tailings - Flotation Circuit Dry Solid tpy 539,889 % of Total Tails % 62.4 17.1.1 Process Flow sheet and Mass and Water Balance

The PF, including mass and water balances, is shown in Figure 17.1 and Figure 17.3. The mass and water balances are given as design values and take into consideration the plant utilization. These balances were derived from the block diagram presented in Section 13. Both Table 17.1 and Figure 17.1 were established using the pilot plant test results that was described in Section 13 and for which the head grade was 1.61% Li2O. Those were used in this study for the equipment and layout design.

In Figure 17.2, the mass balance has been adjusted to the life of mine average head grade of 1.49% Li2O. This mass balance was calculated by assuming a constant Li recovery at each of the major process steps. This assumption was based on previous test work results from experience. The annual concentrate production is 213,600 tpy. Those mass balance results were used as inputs for the financial analysis in Section 22.

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Figure 17.1 – PF and Mass Balance

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Figure 17.2 – PF and Mass Balance at 1.49 % Li2O

Crushing DMS Stage 1

Tails

Grinding

DMS Stage 2 Scrubbing & 2nd Desliming

Spodumene Rougher 1st Desliming Flotation

Mica Rougher Mica Scavenger Spodumene 1st Flotation Flotation Cleaner Flotation

Spodumene 2nd Mica Cleaner Cleaner Flotation Flotation

LIMS

DMS Concentrate Final Concentrate Flotation Concentrate

Flotation Flotation DMS Final Flotation Feed DMS Tails Final Tails Feed Concentrate Concentrate concentrate Tails Solid (tph) 125 37 80 17 8 24 63 101 (% Li Lithium grade 2O) 1.49 0.30 1.59 5.92 6.20 6.01 0.45 0.40 Lithium Distribution (%)1006 685326791521

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Figure 17.3 – Water Balance

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17.1.2 General Process Description and Plant Design

General process and plant design criteria for the concentrator are based on the following: • The location of the jaw crusher concentrator, crushed ore stockpile, major conveyors and thickener are shown on the site plan developed in this study. The site plan is presented in Section 18 and shown in Figure 17.4, Figure 17.5, and also in Appendix A. • Crushing is performed with a 150 kW jaw crusher followed by secondary and tertiary cone crushers. • The crushed ore is upgraded in a two-stage dense media circuit to produce a coarse spodumene concentrate and a middlings product that is ground and treated by flotation. • The DMS middlings product is ground in a rod mill and then combined with screened 0.5 mm material from the DMS circuit and further ground in a ball mill to P80 200 microns. This ground product forms the feed for the flotation circuit. • The flotation circuit consists of desliming and mica pre-flotation followed by further desliming and spodumene flotation. All slimes, mica concentrate, and spodumene rougher flotation tails are combined and filtered in a pressure filter to a moisture content of about 13% and subsequently combined with DMS tailings and transported by truck to the mine waste stockpile. The spodumene flotation concentrate is filtered by belt filter to about 5-6% moisture and combined with the DMS concentrate before being sent to the concentrate storage.

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Figure 17.4 – General Site Plan (Scale of 1/7,500)

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Figure 17.5 – General Site Plant (Scale of 1/1,250)

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17.1.3 Primary Crushing

The run-of-mine, which has an F80 of 600 mm, is fed from the mineralized material stockpile to the static grizzly and hopper by a front-end loader. The grizzly rejects material larger than 800 mm. The hopper discharges in a vibrating grizzly feeder.

The reciprocating grizzly feeder separates the feed into two (2) streams; a fine material stream, which is sent directly to the crushed ore belt conveyor, and a coarse stream, which is fed to the primary crusher. The primary crusher is a jaw-type crusher, measuring 813 mm x 1,220 mm with an installed power of 150 kW, and receives material with an F80 of 700 mm. It has an availability of 60% with a product P80 of 160 mm. This product is discharged onto the crushed ore belt conveyor.

The crushed ore belt conveyor feeds material to the scalping screen feed hopper with a 7.6 m3 capacity. The crushed mineralized material is discharged to the scalping screen feed conveyor and then fed to the 8 x 20 double deck scalping screen (5.8 m2 surface area). The screen is a double-deck machine. The top deck has a 45 mm aperture, while the bottom deck has a 10 mm aperture. The fines are sent directly to the fine ore belt conveyor.

The oversize and mid-size material separated in the screen are conveyed to the secondary cone crusher and tertiary cone crusher respectively. The secondary crusher and tertiary crushers are both cone-type crushers with installed power of 220 kW and 295 kW respectively.

The discharge of the secondary and tertiary cone crushers is transferred to the screen belt conveyor which reports back to the scalping screen feed hopper mentioned above.

Dust is controlled in the crushing area using a dedicated dry dust collector and collecting duct work, which discharges collected material onto the fine ore transfer belt conveyor.

An air compressor is installed in the crushing plant. It supplies the service and instrumentation air for the crushing area and the fine ore bin.

All equipment of the crushing area, with the exception of the belt conveyors, is mounted on trailers and is mobile.

17.1.4 Two-Stage DMS Circuit

The fine mineralized material from the crushing circuit is collected and stored in a (510 m3) fine ore bin. Material from the fine ore bin is conveyed by bucket elevator and conveyor belt to the DMS feed preparation screen with a 0.5 mm aperture. The undersize material from the preparation screen is pumped to the DMS fine product dewatering cyclone, which feeds the grinding circuit prior to flotation. February 2013 QPF-009-12/B

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The oversize material from the DMS preparation screen flows by gravity to the first stage oversized product collector tank, where it is combined with the dense media from the circulating media primary distributor. The combined slurry feeds the first stage DMS cyclones by gravity. The first DMS stage serves to remove petalite in the cyclone overflow. The dense media SG in the first stage is 2.65.

The underflow from the cyclones flows by gravity to the DMS separation screen, where the coarse concentrate product is separated from the dense media and washed. The washed coarse product is then sent to the second DMS stage by bucket elevator. The dense media from the front end of the separation screen is collected in the first stage circulating media tank. The diluted media from the discharge end of the separation screen is collected and sent to the diluted media tank for dense media recovery by magnetic separation and returned to the DMS circuit.

The cyclone overflow product is separated and washed in the same manner as the underflow. The coarse washed overflow product is sent to the DMS tailings hopper.

The washed underflow product from the first stage is collected in the second DMS stage cyclone feed hopper, where it is combined with dense media from the second stage media distributor. The dense media SG in the second DMS stage is controlled at 3.0. The second stage overflow and underflow products are separated and washed in the same manner as in the first stage. The coarse washed underflow product is sent to the concentrate conveyor belt, where it is combined with the flotation concentrate and stockpiled for subsequent transportation to the refinery. The coarse washed overflow product contains a sufficient amount of lithium, and therefore, is sent to the grinding circuit for further lithium recovery by flotation.

A portion of the dense media from the second stage circulating media distributor is diverted to a Multotec tube densifier to ensure that the circulating media does not become too diluted over time. The densified media is returned to the second stage circulating media tank, where a portion is sent to the first stage circulating media tank. The media densities in the first and second stage media circulating tanks are adjusted by the addition of water. The media used is ferrosilicon.

17.1.5 Grinding and Classification Circuit

The coarse washed overflow product from the second DMS stage is fed to a one-hour capacity rod mill feed bin. The rod mill feed is fed to a 450 HP (335 kW) rod mill with an F80 and P80 of 6,500 µm and 375 µm, respectively. The rod mill product discharges into the rod mill pump box which feeds a classification cyclone. The underflow of the DMS fine product dewatering cyclone also reports by gravity to the classification cyclone. The classification cyclone underflow is fed by gravity to a 500 HP ball mill with a P80 of 345 µm and a P80 of 200 µm. The ball mill product discharges into the rod mill pump

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box, which in turn feeds the classification cyclone. The classification cyclone overflow reports to the first desliming cyclone pump box.

17.1.6 Desliming and Flotation Circuits

The grinding circuit classification cyclone overflow is pumped to the first stage desliming cyclones, which have a cut size of about 45 µm. The cyclone overflow is sent to the slimes tailings thickener. The cyclone underflow flows by gravity to the mica flotation conditioning tank where the mica collector, namely Armac C, and fuel oil are added. The conditioning tank feeds a bank of five (5) mica rougher flotation cells. The mica rougher tails flow by gravity to a bank of seven (7) mica scavenger flotation cells. Both the mica rougher concentrate and mica scavenger concentrate are pumped to the mica cleaner flotation conditioning tank where additional collector is added. The mica cleaner flotation conditioning tank feeds a bank of seven (7) mica cleaner cells. The mica cleaner concentrate is pumped to the spodumene rougher tailings pump box. Both the mica scavenger and cleaner tails are dewatered through of a cluster of five (5) 250 mm cyclones and sent to the spodumene flotation circuit.

The dewatered scavenger and cleaner tails from the mica flotation circuit are scrubbed in an attrition scrubber and then de-slimed in the second stage desliming cyclones. The overflow from second stage desliming is sent to the slimes tailings thickener. The thickened underflow is pumped to flotation tails holding tank for subsequent filtration. The underflow from the second stage desliming cyclones flows by gravity to a high density (65% solids) conditioning tank where spodumene collector (LR19) is added. The conditioned slurry is pumped to a bank of five (5) spodumene rougher flotation cells. The rougher tails are pumped to a dewatering cyclone and then to the flotation tails holding tank for subsequent filtration in a pressure filter. The filtered tails are conveyed to the tailings hopper along with DMS tails. The combined tails are then transported by truck to the waste stockpile.

The rougher spodumene concentrate is pumped to a bank of four (4) cleaner flotation cells, where additional collector is added. The pH is controlled by the addition of sodium carbonate. The first cleaner tails are dewatered and returned to the attrition scrubber. The first cleaner concentrate is pumped to a second cleaning stage. The second cleaning stage tails are also dewatered and returned to the attrition scrubber, while the second stage cleaner concentrate is pumped to a wet LIMS magnetic separator to remove iron. The final concentrate is then filtered by belt filter and stockpiled along with DMS concentrate for subsequent transportation to the refinery.

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17.1.7 General Concentrator Plant Services a) Compressed Air Two (2) 1,500 scfm air compressors (one operating and the other on standby) will provide compressed air for the tailings filter press, instrument air, and other plant services. A desiccant dryer of equivalent capacity is also included. An air blower will provide air for the mica and spodumene flotation circuit. b) Freshwater and Process Water Fresh water will be sourced from a well. Fresh water will be used to provide makeup water for the process water tank and to meet clean water requirements for the concentrator. Clean water will be stored in a separate clean water tank, which will also include filtered process water. The clean water will be used for gland seals, reagent preparation (flocculants and dispersants) and firewater. c) Fire Protection Approximately 760,000 litres (757.1 m3) of clean water from the clean water tank is reserved for firefighting. The clean water tank is equipped with a fire water circulation pump to ensure fire water is readily available.

17.2 Hydrometallurgical Processing Plant

Based on the metallurgical testing results described in Section 13, the hydrometallurgical plant process design criteria, process flow sheets, equipment list as well as plant layouts were prepared for a plant feed rate of 213,558 t/y of spodumene concentrate. The following section summarizes the process developed for the Project.

The spodumene concentrate will be transported by train approximately 700 km to the hydrometallurgical plant located in Valleyfield, Quebec. The hydrometallurgical plant processing activities include decrepitation, acid roasting, leaching, purification, electrodialysis, crystallization, precipitation, drying and packaging of the different products.

17.2.1 Plant Design Criteria

The hydrometallurgical process plant is scheduled to operate 365 days per year, seven (7) day per week and 24 hours per day. The plant availability is estimated at 93%.

The hydrometallurgical process plant is designed for a feed capacity of about 26.2 t/h of spodumene concentrate, resulting in a lithium carbonate powder average production rate of 10,000 tonnes per year and a lithium hydroxide monohydrate crystals average production rate of 20,734 tonnes per year. The overall lithium recovery of the hydrometallurgical circuit is 88.6%.

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The hydrometallurgical plant feed consists of a blend of dense media separation concentrate and flotation concentrate. The concentrates characteristics are included in the design criteria summary presented in Table 17.2.

Table 17.2 – Hydrometallurgical Design Criteria - Summary Parameters Unit Value Concentrate average processing rate tonne per year (t/y) 213,558 Concentrate average processing rate tonne per day (t/d) 585.1 Concentrate composition % DMS / % Flotation 31.7 / 68.3

Concentrate grade (total) % Li/Li2O 2.8 / 6.0

From DMS % Li2O 6.2

From Flotation % Li2O 5.9 Concentrate size distribution

From DMS P100 (microns) 1,000

From Flotation P100 (microns) 300 Concentrate emergency stockpile capacity tonnes (t) 10,000 Hydrometallurgical plant operating time hours per day (h/d) 24 Hydrometallurgical plant average feed rate tonnes per hour (t/h) 26.2 Overall hydrometallurgical plant availability % 93.0 Lithium hydroxide monohydrate average tonnes per year (t/y) 20,734 production Lithium hydroxide monohydrate product % 0.1 moisture 2 Lithium hydroxide monohydrate product grade % LiOH-H2O 99.700 Lithium carbonate average production tonnes per year (t/y) 10,000 Lithium carbonate product moisture % 0.1

Lithium carbonate product grade % Li2CO3 99.987 Overall lithium recovery % 88.6 17.2.2 Flow Sheets and Process Description a) Simplified Flow Sheet A simplified flow sheet is presented in Figure 17.6 and summarizes the hydrometallurgical plant process. The equipment list is based on the process flow diagrams and the equipment is sized based on the design criteria and mass balance.

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Figure 17.6 – Lithium Hydrometallurgical Plant Simplified Flow sheet

b) Process Description i) Concentrate Reception and Conversion

The spodumene concentrate is transported from the mine and concentrator site to the hydrometallurgical process plant in 90-tonne railcars (100 short tons). Deliveries are expected approximately every three (3) days (eight (8) railcars per delivery).

The concentrate is discharged from one (1) railcar at a time into a receiving hopper. A reclaim conveyor transports the concentrate from the hopper to a bucket elevator.

The concentrate discharges into one (1) of two (2) bucket elevators that convey the concentrate vertically into the kiln feed silo. A diverter chute allows the concentrate from either bucket elevator to fall onto the emergency February 2013 QPF-009-12/B

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stockpile. A wheel loader is used to reclaim the concentrate from the emergency stockpile. A second reclaim hopper and conveyor are used to convey the material from the emergency stockpile back to the bucket elevators. Only one (1) bucket elevator will be in operation at any given time. The second bucket elevator is used as a stand-by unit.

The kiln feed silo live capacity represents eight (8) hours of continuous hydrometallurgical plant feed.

Two (2) belt feeders feed the concentrate from the silo to the kiln feed belt conveyor. This belt conveyor is equipped with a belt scale to monitor and control concentrate feed to the hydrometallurgical plant.

The concentrate falls into a screw conveyor that will distribute the concentrate in an even layer into the spodumene conversion kiln.

The natural gas fired rotary kiln heats the spodumene to approximately 1,050°C. The high temperature converts the spodumene concentrate from the alpha crystalline structure to the beta crystalline structure. Unlike alpha- spodumene, beta-spodumene is amenable to acid roasting and water leaching.

A flash cooler uses ambient air to cool the beta-spodumene to approximately 200°C.

The acid roaster is a continuous paddle-type mixer. The mixer is constructed of Inconel®, a type of superalloy resistant to high temperatures and corrosion.

Sulphuric acid is sprayed onto the beta-spodumene in the roaster with a stoichiometric excess of 30%. The resulting reaction produces solid lithium sulphate and aluminum silicates.

ii) Concentrate Leach and Primary Impurity Removal

The solid lithium sulphate and the gangue material fall into the concentrate leach tank via a chute. Wash water from a downstream belt filter, which may contain lithium sulphate, as well as a weak lithium sulphate solution from the downstream electrodialysis process are also fed to the tank in order to produce a 50 % (w/w) feed slurry. Lithium sulphate, being soluble in water under these conditions, will dissolve, along with any other sulphates produced during the roasting step (iron sulphate, aluminum sulphate, sodium sulphate etc.). The slurry discharges from the tank by overflow into the first Primary Impurity Removal (“PIR”) tank.

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There are three (3) PIR tanks arranged in series and overflowing one (1) into the other.

An hydrated lime slurry is fed to the PIR tanks to precipitate out impurities such as iron and aluminum in the form of insoluble hydroxides.

The discharge of the PIR tanks is pumped to the downstream belt filter. The PIR belt filter dewaters and washes the PIR residues in order to recover as much lithium sulphate-rich filtrate as possible. Five (5) washing steps are performed.

The belt filter cake, which consists mainly of aluminum silicate, is conveyed outside the plant building where it is stockpiled in anticipation of future marketable possibilities.

iii) Secondary Impurity Removal

In the Secondary Impurity Removal (“SIR”) tanks, the pH is further increased to precipitate even more dissolved metals as solid hydroxides and carbonates. The PIR belt filter filtrate is pumped to the first SIR tank along with lithium hydroxide solutions recovered from the downstream ion exchange and electrodialysis processes. Lithium hydroxide reacts with dissolved manganese and zinc ions which will precipitate as insoluble hydroxides.

A sodium carbonate solution is fed to the third SIR tank. The sodium carbonate will react with dissolved calcium and magnesium ions, precipitating as carbonates which are also insoluble.

The discharge slurry is pumped to a centrifuge to remove the solid impurities from the lithium solution. The solid residue is directed to the tailings. The filtrate is pumped to the SIR product tank. The solution is pumped to a self- cleaning fabric filter where the solids smaller than five (5) microns, and any solids that have precipitated after the centrifuge, are removed.

The filtrate is stored in the ion exchange feed tank before it is pumped to the next cleaning process.

iv) Ion Exchange

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contaminants, it is brought offline to undergo the cleaning/stripping/ regeneration processes. Meanwhile, the lag column becomes the lead column while the third column which was on stand-by becomes the lag column.

Diluted HCl will be used as the column stripping agent while diluted LiOH recovered from the downstream LiOH-H2O crystallization process is used as the regeneration agent.

Clean lithium sulphate solution discharging from the columns is stored in the LiOH feed tank. Waste solutions from the column stripping step are collected in the IX residue tank and pumped to the tailings tank for disposal. Waste solutions from the column regeneration step are collected in the regeneration solution tank and metered back to the first SIR tank.

v) LiOH Membrane Electrodialysis

Membrane electrodialysis is used to convert lithium sulphate to lithium hydroxide.

The electrodialysis cell consists of an anode, a cathode and a series of cationic, anionic and bipolar membranes in between. The lithium-sulphate rich solution is fed to the compartments between the anionic and cationic membranes. An electric current is applied across the cell causing positive ions to cross the cationic membrane and negative ions the anionic membrane. Meanwhile, a weak sulphuric acid solution is fed to the compartment between the anionic membrane and the bipolar membrane and a weak lithium hydroxide solution is fed to the space between the cationic membrane and the bipolar membrane.

In total, it is estimated that 80 cells are required to convert the lithium sulphate solution to lithium hydroxide. The electrodialysis cells and ancillary equipment will draw approximately of 25.1 MW/h of electrical power.

A portion of the LiOH solution is directed toward the lithium carbonate production circuit) to produce the required 10,000 t/y of lithium carbonate. The remaining solution is pumped to the LiOH-H2O crystallization circuit.

The by-products of the electrodialysis process are a weak solution of sulphuric acid and a spent solution of lithium sulphate. The sulphuric acid is neutralized as solid gypsum and the lithium sulphate solution is recycled to the concentrate leach tank.

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vi) LiOH-H2O Crystallization

The crystallization circuit produces lithium hydroxide monohydrate crystals from the lithium hydroxide solution generated through electrodialysis.

The crystallization circuit consists of a two-stage mechanical vapor recompression falling film evaporator followed by a single-effect steam- driven forced circulation crystallizer.

The LiOH solution feed is pre-heated and enters the evaporation circuit where a concentrated LiOH solution is produced. Evaporated water is recovered as distilled water and recycled to the plant.

The concentrated LiOH solution feeds the crystallizer. This unit removes water from the incoming feed, which causes the precipitation of LiOH-H2O crystals. The remaining liquor is purged from the system. This purged solution replaces the NaOH solution that would normally be purchased to be used in the IX circuit for resin regeneration and in the SIR circuit for pH control. The excess purge is directed toward the lithium carbonate production circuit. The evaporated water is recovered as distilled water and recycled to the plant.

The LiOH-H2O crystal grade is expected to be around 16.4 % Li (w/w). The crystals are separated from the mother liquor in a centrifuge.

vii) LiOH-H2O Drying and Packaging

The LiOH-H2O crystals coming out the centrifuge are dried before the final packaging step. A flash drying system using indirect heating with natural gas will be used.

A robot operated system is used to package the crystals into one (1) tonne bags. Around 61 bags per day will be produced.

viii) Li2CO3 Precipitation

The feed to the lithium carbonate production circuit comes from two (2) sources: lithium hydroxide from the electrodialysis circuit and sodium-rich lithium hydroxide solution from the crystallization circuit purge.

The Li2CO3 production process is carried out in two (2) steps. The first step is called the Lithium hydroxide Carbonization (“LC”) or it may be referred to as the lithium carbonate precipitation step. The second step is called the lithium bicarbonate decomposition step (“DC”).

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The LC step is carried out in a single tank. Carbon dioxide is injected at the bottom of the tank to react with lithium hydroxide, producing lithium carbonate.

The slurry is pumped to a decanter. The decanter overflow containing a lithium bicarbonate by-product is fed to the second step. The decanter underflow contains the precipitated lithium carbonate and is pumped to a belt filter feed tank.

In the DC step, the lithium bicarbonate solution is heated to 95ºC using reactors installed in series. The tanks are jacketed and heated with steam.

The slurry from the DC step is pumped to a decanter. The decanter overflow is directed towards the water treatment system. The decanter underflow joins the LC decanter underflow in the belt filter feed tank.

A vacuum belt filter is used to dewater the lithium carbonate slurry. The belt filter cake falls by gravity to the flash dryer. The belt filter filtrate is returned to the decomposition step to convert any unreacted lithium bicarbonate into lithium carbonate.

ix) Li2CO3 Drying, Pulverization and Packaging

A flash drying system using indirect heating with natural gas dries the lithium carbonate final product.

The dried product is transferred to a jet mill. The jet mill uses high pressure air to pulverize the lithium carbonate to a P80 of about five (5) microns.

A robot operated system is used to package the Li2CO3 into 25 kg bags. Around 1,178 bags per day will be produced.

x) Gypsum Precipitation and Disposal

The weak sulfuric acid solution coming from the electrodialysis circuit is neutralized with hydrated lime to produce gypsum. The gypsum slurry is pumped to a thickener.

The thickener overflow is directed towards the water treatment system while the underflow is pumped to the gypsum tailings pond. A dedicated pond is planned to be used for the gypsum tailings because of future gypsum marketable possibilities. Water recovered from the pond is recycled to the plant’s water treatment system.

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xi) Tailings Disposal

Tailings from the SIR circuit and most of the plant sumps are collected in the SIR tailings tank. Lime is added to neutralize the tailings.

The tank’s content is pumped to a thickener. The thickener overflow is directed towards the water treatment system while the underflow is pumped to the SIR tailings pond. Water recovered from the pond is recycled to the plant’s water treatment system.

xii) Reagents – Hydrated Lime Quicklime is purchased in bulk and stored on site in a 200 tonnes capacity silo. A slaker is used to prepare hydrated lime. Lime is used to control the SIR tailings pH, precipitate gypsum and precipitate iron in the PIR tanks. – Soda Ash Soda ash is purchased in one (1) tonne bags and delivered to the mixing tank using a screw feeder followed by an eductor. Approximately 1.3 bags of soda ash are required daily. Soda ash is used to precipitate calcium and magnesium in the SIR circuit. – Hydrochloric Acid Hydrochloric acid is delivered in tanker trucks and unloaded into a single storage tank. The acid is purchased undiluted at a concentration of 32 % (w/w). Hydrochloric acid is used in the IX circuit for stripping the resin. – Sulfuric Acid Sulfuric acid is also delivered in tanker trucks and unloaded into a single heated tank located outside the building. Sulfuric acid is purchased at a 93 % concentration and is used at this strength. The only requirement for sulfuric acid is for the acid roasting step.

xiii) Services – Cooling Water A 1,050 m3/h capacity tower is required to provide cooling water to the electrodialysis circuit and crystallization circuits.

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– Steam A 1,400 HP boiler provides steam at 30 psig and 134 ºC. A total of about 21,500 kg/h of steam is required for the crystallization circuit and the decomposition reactors. – Compressed Air Two (2) high pressure air compressors (one in operation and one stand-by) provide instrument and plant air for all users throughout the plant. Only the jet mill requires a dedicated air compressor. The main instrument air users are the various dust collection systems and the instruments. Plant air is required in the PIR tanks and for the utility stations. – Natural Gas Natural gas is used to fire the kiln burner and both dryer burners. No gas storage is planned on site. Natural gas will be supplied by Gaz Metro from a connection to their existing network. Overall gas consumption is about 3,200 m3/h.

xiv) Water Management – Process Water Process water comes from the PIR belt filter washes two (2) to five (5). Process water can be used for reagents preparation (lime and soda ash), for washing the SIR centrifuge cake and backwashing the SIR self-cleaning filter. The excess process water is pumped to the water treatment system. – Distilled Water Distilled water consists of the condensate from the crystallization circuit. The distilled water is at a higher temperature than all the other sources of water (higher than 50 ºC).

Distilled water users are the PIR belt wash water make-up, the boiler make- up and the lithium carbonate belt filter cake and belt wash water. A make-up from the purified water tank is required to meet the distilled water demand. – Purified Water Many sources of water and process solutions are collected in the water treatment feed tank. The gypsum, SIR tailings and decomposition thickeners overflows are collected in this tank along with the IX residue, the spent Li2SO4 from electrodialysis and the Li2CO3 cake wash water.

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This water is first treated using a potable water treatment system. The potable water is then fed to a purified water treatment system that consists of reverse osmosis. To meet the purified water requirements, potable water from the city of Valleyfield is used as make-up water and is fed to the purified water treatment system.

Purified water is used in the acid roaster scrubber, the IX circuit, the electrodialysis circuit and the belt filters vacuum pumps.

The concentrate from the reverse osmosis system is directed to the city of Valleyfield’s sewage system.

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18.0 PROJECT INFRASTRUCTURE

The mine and the concentrator infrastructures will be located at Whabouchi and are described in Sections 18.1 to 18.9. The hydrometallurgical infrastructures will be located at Valleyfield and are described in Section 18.10.

18.1 Whabouchi General Site Plan

Two different scale plans have been issued for the Project: Drawing No.°3073002°- 0010000-41-D20-0001, to a scale of 1 / 7,500, presents the overall area where the mining Project is to be undertaken and Drawing No.°3073002°-0010000 41-D20-0002, to a scale of 1/1,250, showing the processing plant and related infrastructures more precisely, including the office and mining garage area plus the related services for water supply and sewage disposal and the surrounding hauling and service roads. Both drawings are presented in Appendix A.

They contain the topographic information from LIDAR laser mapping technology provided for the Project by Nemaska Lithium.

The site is accessible by the Route du Nord, which also leads to the existing base camp that the Owner intends to use for both construction and operation purposes.

The camp site is about 12 km from the site toward the west and the Nemiscau airport is another 7 km farther away.

All mining activities will be concentrated in a 75 hectares area; area delimited by the Route du Nord to the north, the UTM 5,724,500 mN, NAD83 Zone 18 to the south, the Des Montagnes Lake to the west and the Spodumene Lake to the east.

18.1.1 The Route du Nord Public Road

The Route du Nord is actually located just outside a 500 metres safe distance from the open pit.

This road is scheduled to be relocated north of the Phase II waste rock pile, as shown on drawing No°3073002°-001000-41-D20-0001. This relocation will require approximately 2.1 kilometres of new road to be built.

18.1.2 Mine Service and Hauling Roads

The site hauling roads were designed for Cat 772 with an 82,100 kg target operating weight. The quantities required for building those roads were established based on draft profiles for each of the three (3) sections of mining road. Those profiles were determined

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based on the roads not exceeding a maximum slope of 8 %. To suit the Cat 772 dimensions, those mining roads have been designed to be 17 metres wide.

The quantities of fill required to construct the service roads were established the same way using a standard AASHO-H20 highway live load. All service roads were designed to be ten (10) metres wide.

18.1.3 Service Buildings

The service buildings are to include a gate house, an administration office, and a mine management and engineering office, all located along the site access road.

The gate house will be used to monitor pedestrian and all vehicle access to the site. A control gate will be operated by a security guard.

The administration and the mine management offices are proposed to be connected together for easier personnel circulation and access to all three (3) conference rooms.

The first sector will mostly be reserved for administration purposes and will house such services as the computer servers and technicians, the environmental department, the first aid clinic and finally, the general manager’s and the human resources offices. The administration office will include an 18-person conference room.

The second office building will be the main gateway for the mine personnel, with changing and lunch room facilities. The mine superintendent and foreman offices will also be located within this complex, as well as two (2) adjacent conference rooms that could be used for staff training and coordination meetings. The mine garage area is easily accessible from this complex within a 25 metres distance. Finally, the same office unit will house the technical department, including the engineering and surveyor personnel, as well as the geology experts.

18.1.4 Maintenance Garage and Warehouse

As discussed previously, the garage building is to be located at an easy walking distance from the service buildings and personnel parking lot.

The garage will be used for servicing the mine equipment in addition to other mobile equipment. The garage will be equipped with overhead cranes, tools, air compressors, lubricants, and will house the following facilities: • Two (2) maintenance bays for the mine trucks and other large mobile equipment; • One (1) large equipment wash bay; • One (1) light-vehicle maintenance bay.

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As for the 270 m2 adjacent warehouse, it will have shelving that is arranged to maximize storage space. The warehouse will have access to the equipment maintenance area through an internal door.

Apart from the maintenance staff changing-room and kitchen, the garage complex will also house three (3) offices for the garage management personnel, i.e., the superintendent, the foreman and the maintenance clerk.

18.1.5 Fresh Water Supply

It is foreseen that fresh and fire water will be supplied to the site from wells since the process and domestic water demand is relatively low, basically under 6 m3 per hour.

It is otherwise expected that the process itself will need to be fed about 20 m3/h for start- up purposes. This is the basic flow rate that is being retained for the fresh water supply source.

18.1.6 Sewage Treatment

At the site, sewage treatment facilities will be built for all three (3) centers of operation: the administration buildings, the maintenance garage and the concentrator.

Without shower facilities on site, as it had been directed by Nemaska Lithium at the beginning of the study, the unit water demand of 75 litres per person per day has been used to establish the amount of sewage to be treated daily.

Once the manpower requirement was established, it was found that the largest producer of waste water would be the administration/mine offices, but with a limited quantity of 3.75 m3 per day.

A maximum quantity of 9.25 m3 per day is expected to be generated by the entire site.

As the geotechnical investigation report L-11-1452 from Journeaux Assoc. dated December 2011 demonstrated, some locations on the site may be suitable for seepage field operations. The preliminary design of all sewage disposal installations has been based on that simple technology. It is expected that the construction of three (3) septic tanks and seepage fields would be less expensive than a 9.25 m3 per day modular treatment plant.

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18.2 Control System

18.2.1 Process Network

An automation Ethernet backbone at 100 Mbps, in a ring type topology, shall link all the main automation equipment, such as SCADA, Historian, HMI, PLC (processor only), and the main electrical substation.

One (1) fibre optic ring shall be implemented to link all main plant equipment. Logical rings shall also be implemented to separate the supervisory network of the I/O networks. Logical segmentation shall be done by using VLAN tagging.

Automation services are: • PLC inter-communication; • PLC/Remote I/O communication; • SCADA communication; • Field Device communication. 18.2.2 Process Control System

Each sector will have its own Process Control System (“PCS”) with its own remote I/Os. Eight (8) main processors shall control the following sectors: Crushing, DMS, Grinding, Concentrate, Water Control and Utility, Tailings, and Plant Emergency Power Control. Other PLCs will be supplied with some of the main equipment (crushers, filters, etc.).

The central SCADA system will be able to control and supervise the remote equipment, but during communication outages, the equipment will be controlled locally.

18.2.3 Cabling Strategy between PLC Panels and Instruments

The cabling strategy used between the PLC panels and instruments assumes that the instruments are wired to Remote I/O or PLC. The average distance from instruments to Remote I/O used in the estimate is 30 metres.

18.2.4 SCADA

The SCADA is based on client/server technology and will be comprised of two (2) SCADA servers for redundancy, four (4) main SCADA operators and six (6) local operation clients throughout the plant.

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18.2.5 SCADA and PLC Power Sources

In case of plant power loss, a diesel generator will provide emergency power to different electric loads throughout the plant. PCS, switches, main servers, phone system, and security systems will be fed by a UPS. These UPS systems will be powered from the emergency power.

UPS used for the Control System will be monitored using SNMP Ethernet protocol.

18.2.6 Redundancy

For the automation network, the ring shall offer a second route in case of a communication outage on one (1) segment.

18.2.7 Process Analog Instruments

Process analog instruments will, whenever possible, support HART protocol and they will be wired to analog inputs/outputs of the process controller by means of traditional 4-20 mA loops.

18.3 Communication System (Local and External)

18.3.1 Telecommunication Guidelines

The telecommunication system will be based on Ethernet links throughout the plant buildings and administrative buildings.

A single-mode fibre optic backbone will be deployed to accommodate both automation and corporate services on the same cable. For remote sites, a Wimax link will transport automation and corporate services. For short cable runs, a CAT6 cable will be used. The CAT6 cable will be armored when installed in an instrumentation cable tray.

18.3.2 Telecommunication Systems

A mobile radio system will be provided for the construction and operation phase covering the mine site, the construction site and the unloading area. This system is provided by a telecom tower and a telecommunications shelter hosting all the communication equipment. An IP phone system will be provided at the beginning of the construction phase.

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• Access control system (gate, door); • Mobile radio system. 18.3.3 Telecommunication Services

The site will be connected to an Internet Service Provider (“ISP”) via a microwave link to the Relais routier Nemiscau (12 km west of the main facility), expected to be available by June 2014. The microwave link will be supplied and maintained by the ISP. For the PEA Study, the bandwidth costs have been evaluated with 5 Mbps for the first year and then ten (10) Mbps for the next two (2) years on a 3-year contract basis.

A backup system will use a cellular modem. The current cellular coverage allows the usage of the 4G technology.

The IP PBX phone system will be connected to an Internet Telephone Service Provider (“ITSP”).

18.3.4 Telecommunications Distribution

During the construction phase, all communication services, such as Internet and phone, will be distributed via Wi-Fi, Wimax and Microwave point-to-point radios to reach all buildings. All mine trucks and pick-ups will be equipped with a Wimax/WiFi antenna that shall also act as a WiFi local access point.

A fibre optic link will be extended from the process plant to the administrative offices, to the mine garage and to the main tower using the 25 kV distribution lines or underground trenches.

18.3.5 Corporate Network

A corporate Ethernet backbone at 1 Gbps in a star type topology will be deployed for the process and security video broadcast distribution, the IP phone system and the corporate network applications.

All the major network equipment will be located in dedicated server rooms located in the administrative office, the telecom shelter and in the control room (or main substation). For the feasibility study, the administrative server room itself will reside in the administration offices.

Corporate services are: • Wired/wireless phones and phone system; • Process and security camera system; • Access control system (gate, door);

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• Fire detection. In order to transport all services on a single wireless link, VLANs will be reserved for each service. a) Camera System A camera system, with recorder and a viewer, will be installed in the main gate office. Aside from the gate cameras, seven (7) cameras will be installed in the plant for process control purposes. One (1) viewing station will be installed in each control room for process control purposes.

18.4 Heating, Ventilation and Air Conditioning

The heating, ventilation and air conditioning (HVAC) will be provided for all buildings based on the required working temperatures. The heating and air conditioning design conditions are based on data provided by Environment Canada for the site. • January 2.5% design dry bulb temperature, -37 °C; • January 1% design dry bulb temperature, -40 °C; • July 2.5% design dry bulb temperature, 27 °C; • July 2.5% design wet bulb temperature, 20 °C; • Annual total degree days below 18 °C, 7,500 degree days. 18.5 Fuel Storage Facilities

The fuel storage facility is designed to have 90,000 litres of storage capacity and is built on a concrete pad. The system will include the following: • Two (2) 50,000 litres capacity double-wall diesel fuel tanks; • Two (2) diesel fuel distribution points for heavy and support mining equipment; • One (1) diesel fuel distribution point for small vehicles; • A complete command center for the transfer pump system and distribution station, including luminaires. 18.6 Water Supply and Fire Protection

Fresh and fire water will be supplied to the buildings from a new well. A pumping station, with two (2) electric pumps, will draw water from the well to a 760,000 litres capacity reservoir used for fire protection services located at the concentrator. The fire water pumping station will include one (1) electrical pump complete with jockey pump and one (1) emergency diesel powered pump.

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Design provides for a plant-wide fire protection system, including all electrical rooms and other high risk areas. Final design will confirm fire protection in further detail, in order to conform to local regulations along with the Insurer’s requirements. The fire water will be distributed through an independent underground and heat traced piping system. Fire hydrants will be located on-site along access roads, allowing for easy access in the event of an emergency.

At the fire water entrance of each building, a post indicating valve and Siamese connection will be provided. Fire cabinets, complete with 1½” fire hose, will be located at every level and throughout the building at 30 m intervals. Fire cabinets, complete with 2½” fire hose connections, will be located in all stairwells at every level, including roof level. For extra hazard in heated or unheated areas, wet or dry sprinkler systems will be provided. Special pre-action, double interlock detection systems will be provided for electrical rooms.

The fire alarm system will consist of a panel located in the guardhouse, with detectors and manual stations installed to cover all of the different areas. Alarm signals will be automatically transmitted to the security station. There will be emergency meeting places to be designated in the event of a fire alarm, along with evacuation routes and procedures. Fire extinguishers will be included in fire cabinets and will be required in certain areas such as offices, laboratory, warehouse, lunch rooms, and fuel stations. Hot work permits will be required for maintenance on rubber lined equipment.

18.7 Power Supply and Distribution

18.7.1 Power Line, Main Substation and Electrical Distribution

The total power demand of the Project was determined to be approximately 7.5 MW, based on the estimated connected load, running load and running power. Table 18.1 shows the power demand breakdown by sector.

Table 18.1 – Estimated Total Project Power Demand Power Demand Area (MW) Crushing 0.85 Processing 3.42 Infrastructure 2.41 Network Loss (2%) 0.13 Power Demand Subtotal 6.82 Security Factor (10%) 0.68 Total Power Demand 7.50 Note: The power demand was calculated using an average efficiency factor, load factor and diversity factor.

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The Whabouchi Project will get its power from the Hydro-Québec Albanel substation. From that point, a new 25 kV overhead power line will be built over approximately 20 km and will terminate at a new outdoor substation. At this substation, the incoming voltage will be stepped down to a distribution voltage of 4.16 kV through one 10/13 MVA transformer. The transformer will feed a 5-kV switchgear.

The 5-kV switchgear is located indoors in the pre-fabricated electrical room and provides power to: • The Administration/Garage area through an overhead 5-kV line; • The primary, secondary and tertiary crushing package; • A MV MCC; • A capacitor bank to improve the power factor to 95%; • Three (3) 2-MVA unit substations that generate the 600 V. There is no electrical distribution to and within the mine, as all mining equipment, including pumps, will be powered from diesel motors.

The rod and ball mills have the biggest motors on this Project, at 500 HP each, and are driven by a 5-kV variable frequency drive equipped with an active front end to minimize the generation of harmonics onto the electrical system, for which the electrical supply utility has very strict limits.

18.7.2 Emergency Generators

Emergency power will be provided at 600 V by a 1 MW diesel powered generator. The following process equipment will operate on emergency power: • Auxiliary services of bigger motors of the plant; • Tank agitators; • Thickener rake; • Sump pumps; • Partial heating/lighting; • Communication and control equipment. 18.8 Effluent Water Treatment

At the time of this Report, the only effluent considered for water treatment in the Project is the effluent water produced during truck wash at the mine garage. The wash effluent is collected and treated along with the effluent for sanitary treatment.

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18.9 Camp Accommodations

A construction camp and a permanent camp are required for the Project. These two (2) camps will be built at the location of the Compagnie de Construction et Développement Cri Ltée (“CCDC”) site. The CCDC Ltée will be mandated for the two (2) camps and all required services.

Initially, the permanent camp will be used as a construction camp. An additional construction camp (to accommodate the expected peak workforce) will be built as an extension to the permanent camp. In total, 125 rooms in the permanent camp and 90 rooms in the additional construction camp will be built for a total of 215 rooms.

The common areas of the permanent camp, such as the kitchen and the recreational areas will be used by the residents of the construction camp. Thus, the kitchen capacity in the permanent camp is able to accommodate the additional load of the peak construction workforce. Outdoor pathways will be in place to connect the permanent camp with the construction camp. The permanent camp will be in place in April 2014 and will be the first to be used. The additional construction camp will be built for September 2014 and will be used until the end of the construction and commissioning period.

18.9.1 Site Location

The CCDC site designated for the construction camp and the permanent camp is located at the Relais routier Nemiscau (km 291 of Route du Nord), approximately 12 km west of the Whabouchi Project site. Employees will be transported daily by bus.

18.9.2 Permanent Camp a) Permanent Camp Services The CCDC Ltée will be mandated for the following services: • The construction of the camp; • The maintenance of the camp; • Housekeeping and janitorial service; • Camp management and security. b) Permanent Camp Dormitory The permanent camp dormitory complex is designed as wings to a one-storey configuration connected to a central hub connecting to the services core building.

The dormitory complex includes: • Rooms to accommodate reduced-mobility staff;

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• Thirty (30) standard rooms, approximately 120 square feet, with private bathrooms; • Ninety-five (95) standard rooms, approximately 100 square feet, with shared bathrooms; – In shared facilities: sinks, toilets and showers - in each dormitory there would be six (6) sinks, six (6) toilets (or urinals) and six (6) showers per fifteen (15) rooms. c) Permanent Camp Services Core Building The services core building will provide dining and recreational facilities for the residents of the camp. The service core building includes: • Kitchen, preparation and storage; • Services and dining area; • Warehouse; • Lobby, mustering area, washrooms; • Locker and mud room. The room will be used for personal items storage during employee rotation and will be also used as a mud room; • Administration area, security office, medical facility and mail room; • Exercise and weight rooms; • Game and activities room; • Commissary, TV and lounge areas; • Central laundry room. 18.9.3 Supplemental Construction Camp

The supplemental construction camp dormitory includes:

• Approximately 90 standard rooms (around 100 square feet) with shared bathrooms; • A shared facilities/bathrooms equipped with six (6) sinks, six (6) toilets (or urinals) and six (6) showers per 15 rooms; • A laundry area. 18.10 Valleyfield Plant Infrastructure

This section summarizes infrastructure such as power line, site roads, concentrate rail load out, site buildings, tailings storage facilities and site services that are required to complement the processing of spodumene concentrate at the Hydrometallurgical Processing Plant.

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All topographic information for the location of infrastructure was gathered from readily available data from the City of Valleyfield, including a geotechnical drilling investigation of surrounding land. It is understood that a LIDAR based topographic map will be available for the Feasibility Study phase of the project.

Detailed geotechnical investigations will need to be performed in order to optimize civil design criteria related to the foundations of the hydrometallurgical process plant.

An overall general site layout and access plan is shown on Figure 18.1.

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Figure 18.1 – Overall General Site Layout and Access Plan

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18.10.1 Power Line

The Hydrometallurgical Processing Plant will be fed through a 120 kV overhead electrical power line supplied and installed by Hydro-Québec from the existing distribution point at the Langlois substation.

18.10.2 Site Roads

The Hydrometallurgical Processing Plant is accessible year-round without any interruption by the city’s road network. The access road is located directly to the east of the property, off Des Érables Boulevard. A second access road is located in the south- western corner of the site, off Gérard-Cadieux Boulevard. A single guard house will be used to monitor both access points.

Site and service roads will be eight (8) m wide and will provide access to: • Process facility from both city roads; • Aluminum silicate pile; • Shipping warehouse; • Warehouse. 18.10.3 Concentrate Transportation and Load-Out Facility

Concentrate will be shipped by railcar from Chibougamau to the hydrometallurgical processing plant. An existing CN railway runs along the western side of the site. Railcars will be diverted from the railway at the north-western corner of the site onto a new 600 m railway. This new railway will run in parallel to the SW limit of the site. The new railway will be doubled over a 250 m distance to facilitate railcar handling.

It is expected that deliveries of eight (8) 100-short ton railcars will be made every day. Railcars will be emptied from the bottom one (1) at a time over the concentrate receiving hopper. The hopper is equipped with a feeder that directs the concentrate onto an unloading conveyor.

18.10.4 Tailings Storage Facilities

A preliminary assessment of tailings disposal requirements to store and manage the tailings and process water was prepared for the hydrometallurgical processing plant.

Tailings storage design has been performed for the three (3) distinct plant tailings. The concentrate leach and PIR tailings will be stockpiled dry. The gypsum slurry produced by neutralizing the waste from the electrodialysis process will be stored in a dedicated pond. The remaining tailings will be stored in a second pond.

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a) Aluminum Silicate Pile Concentrate leach and PIR tailings (aluminum silicates) are discarded from the Hydrometallurgical Processing Plant as a cake from a belt filter. They will be stockpiled on a pad outside. Pile sizing has been based on an annual production of 219,000 t/y (dry) of tailings with a 27 % moisture content. Design criteria were based on an assumed final density of 1.76 t/m3 and a single year of storage. Since aluminum silicate is a marketable product, it has been assumed that most of the aluminum silicate produced by the plant will be sold. Therefore, the total designed stockpile capacity is 170,000 m3. Provisions for a 200 by 120 m pad have been made in the north western corner of the site. b) Gypsum Storage Pond Gypsum is produced as a by-product of the electrodialysis process. Gypsum slurry is discarded from the Hydrometallurgical Processing Plant at a rate of 80,600 t/y (dry) with 50% moisture content.

Because gypsum is a marketable product, gypsum storage pond sizing was based on a one (1) year storage capacity. Design criteria were based on an assumed final depositional density of 1.39 t/m3 and a total tailings placement requirement of about 86,700 m3. The water volume pumped into the tailings ponds is expected to be 240 m3 per day. The tailings will retain about 50% of the pumped water while 50% i.e. 120 m3 per day will be available for reclaim.

Provisions for a 100 by 100 m pond have been made in the south eastern corner of the site. c) SIR Tailings Pond The SIR Tailings Pond will receive the balance of the tailings generated by the Hydrometallurgical Processing Plant at a rate of 2,200 t/y (dry) with 50% moisture content. The pond is sized to receive 20 years of tailings.

Design criteria were based on an assumed final depositional density of 1.47 t/m3 and a total tailings placement requirement of about 44,000 m3. The water volume pumped into the tailings pond is expected to be 6.5 m3 per day. The tailings will retain about 50% of the pumped water while 50% i.e. 3.3 m3 per day will be available for reclaim.

Provisions for a 75 m by 75 m pond have been made in the south eastern corner of the site, north of the gypsum storage pond.

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18.10.5 Site Buildings

The Hydrometallurgical Processing Plant building will house the decrepitation, roasting, leaching, purification, electrodialysis, crystallization, precipitation, drying and packaging circuits. It will be 86 m wide by 122 m long and 30 m high at its highest point. The building will also house: offices, a metallurgical lab, control room, electrical rooms and change rooms.

In addition, the site will include the following buildings: • Shipping warehouse; • Warehouse; • Guardhouse. 18.10.6 Site Services

Provision has been made for the following site services: • The plant will be connected to the City of Valleyfield’s existing potable water distribution network and sanitary waste water treatment facility; • The plant will be connected to the existing Gaz Metro natural gas distribution network in order to provide fuel to the kiln, dryers and boiler; • Reclaim water system from the SIR tailings pond and gypsum tailings pond; • Water treatment for recovered process water; • The power requirements of the plant will be supplied by a 120 kV power line. The total power demand is estimated at 32.1 MW of which 31.1 MW is required for the process; • Based on the power requirements, one (1) oil type 37.5 MVA transformer was selected; • An emergency power system consisting of one (1) MW diesel generator will provide a standby source of power to feed essential services (emergency and exit lighting, fire protection equipment, etc.) as well as critical process loads (slurry tank agitators preventing settling down of material, thickener lifting devices, etc.) in the event of power loss from the grid; • Allowances have been included in the estimate for automation, fire alarm, communication system and security system; • Allowances for plant mobile equipment (light vehicles, loaders, railcar movers and lifts) is included.

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19.0 MARKET STUDIES AND CONTRACTS

This Section summarizes the key information from these studies about lithium markets two (2) market studies. These studies were prepared, one (1) by SignumBOX and the other by Roskill Consulting Group Ltd. (“Roskill”), independent and experienced consultants. SignumBOX was mandated to perform a market study to evaluate potential target markets for the lithium concentrate. Roskill was mandated to evaluate the battery grade lithium hydroxide market.

19.1 Lithium Hydroxide and Lithium Carbonate

In this Section, two (2) market study reports prepared for Nemaska are summarized. One called “Battery grade lithium hydroxide Market Study”, as of September 2012, by Roskill Consulting Group Limited and the other prepared by SignumBOX Inteligencia de Mercados, entitled “Lithium Minerals Market”, as of March 2012.

19.1.1 Introduction

Lithium is the lightest and more reactive of the alkali metals and reacts with water, oxygen, carbonate dioxide and nitrogen at room temperature. It is a soft white-grey metal and giving its high reactivity in nature, it is always present as a compound. Its natural characteristics make it a suitable product for many applications in several industries. Some of the properties that allow lithium to be used in many different applications are: the highest specific heat capacity amongst solids, a high electro chemical potential, a low atomic mass and a low density.

Lithium occurs in pegmatites, brines, clay and ocean water, but only the one found in pegmatites and brines is commercially viable at current lithium prices.

19.1.2 Demand

The following sub-sections will cover the different aspects of the demand for lithium hydroxide and lithium carbonate and their outlook. a) Lithium Applications Lithium is used in a vast range of applications that can be classified in two (2) main markets:

Technical application markets: These markets demand lithium minerals, with very specific requirements allowing for direct use of the ore, such as low iron content. The main markets for technical grade lithium minerals are glass and ceramic. As the spodumene mineral found in the Whabouchi deposit is of chemical grade focus

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in this section will be on the different applications related to the chemical application markets.

Chemical application markets: These markets require different common lithium chemicals, among them lithium carbonate, lithium hydroxide, lithium bromide, lithium chlorite, butyl lithium and lithium metal. These lithium compounds are obtained through the processing of spodumene mineral or from brines.

Since the business case of Nemaska, and of this Report, is the production of lithium hydroxide and lithium carbonate, this section will summarize the main applications relating to these two (2) lithium compounds in particular. b) Lithium Consumption 2011 As shown on Figure 19.1 below “Total lithium consumption by application”, the battery market represents the largest use of lithium with 29% of the overall consumption in 2011. After batteries, the combined frits and glass market represents the second largest application with a 28% market share, while lubricating greases is the third largest application with 14% of the total consumption.

Figure 19.1 – Total Lithium Consumption by Application

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c) Lithium Carbonate and Lithium Hydroxide The two (2) main lithium compounds used by the different industries are lithium carbonate, representing 48%, and lithium hydroxide, representing 20% of the overall demand. Currently lithium hydroxide is mainly used in the lubricating greases as shown in Figure 19.2.

Figure 19.2 - Lithium Demand By Compound (2011)

d) Lithium Consumption Forecast: 2011- 2025 As can be seen from Table 19.1 below both lithium carbonate and lithium hydroxide demand are forecasted to increase significantly compared to other lithium compounds used in other industrial applications. Lithium hydroxide demand in particular is predicted to grow from the actual 27,533 tonnes level in 2011 to over 183,303 tonnes by 2025, the demand almost doubling every five (5) years from today.

Table 19.1 – Lithium Demand By Compound – Forecast 2011-2025 Compound – Tonnes 2011 2015 2020 2025 LCE Lithium Carbonate 66,736 95,068 149,743 253,739 Lithium Hydroxide 27,533 49,889 99,297 183,303

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Compound – Tonnes 2011 2015 2020 2025 LCE Lithium Concentrate 19,229 25,683 31,393 36,757 Lithium Metal 7,755 9,763 13,193 18,256 Butil-Lithium 7,293 9,445 13,232 18,617 Lithium Chloride 7,616 9,516 12,782 16,922 Other Lithium Compounds 3,893 5,368 8,103 12,525 Total Demand 140,056 204,732 327,743 540,119 Source: SignumBOX estimates This forecasted increase in demand for lithium hydroxide is mainly due to the increased use of lithium hydroxide in rechargeable batteries, especially batteries for the automotive application. Despite the fact that lithium carbonate has been widely used in batteries for portable devices, lithium hydroxide apparently suits better in cathodes such as lithium iron phosphate (LiFePO4), which are being developed for electric cars. There are different methods of producing LiFePO4 some of them require exclusively lithium hydroxide or lithium carbonate while other methods works equal with either. In the case of the method that works better with lithium hydroxide, LiFePO4 solution is prepared for lithium hydroxide monohydrate LiOH- H2O. It is estimated that by 2025, 50% of the lithium requirements for hybrid electric cars would be in the form of lithium hydroxide. It is also expected that the use of lithium hydroxide in batteries for portable devices will increase from 5% in 2011 to about 20% in 2025.

Table 19.2 “Lithium consumption by application – forecast 2011 – 2025” clearly shows that the major increase in demand is expected to come from the battery industry, mainly for the portable devices secondary (rechargeable batteries) and the Hybrid (“HEV”) and electric vehicles (“EV”).

Table 19.2 – Lithium Consumption By Application – Forecast 2011 - 2025 Application– Tonnes LCE 2011 2015 2020 2025 Batteries – Portable Devices 27,416 44,865 71,009 105,236 Secondary Batteries – Portable Devices 3,000 3,647 4,654 5,940 Primary Batteries – HEV/EV 3,359 18,223 62,412 181,628 Batteries – 2WEVs 3,607 7,401 14,455 23,274 Batteries – Grid 500 2,500 5,000 7,500 Frits 20,000 24,308 30,581 36,320 Glass 17,000 20,662 25,993 30,872 Lubricating Greases 18,000 24,249 34,641 44,211

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Application– Tonnes LCE 2011 2015 2020 2025 Air Conditioning 5,400 6,439 8,100 9,621 Continuous Casting Powders 6,000 7,154 9,000 10,689 Medical 4,000 4,502 5,194 5,876 Aluminium 5,000 5,933 7,464 8,865 Polymers 4,000 4,793 6,029 7,161 Others 12,000 14,308 18,000 21,379 Total Consumption 129,282 188,983 302,532 498,571 Source: SignumBOX estimates This situation implies that batteries will represent the largest use of lithium hydroxide in 2025; increasing from 3% in 2012 to 69% in 2025 and those batteries will continue to be the main application for lithium carbonate.

Figure 19.3 to Figure 19.6 below, show the evolution of the demand for lithium hydroxide and lithium carbonate from the 2011 level to the forecasted 2025 expected demand.

Figure 19.3 – Lithium Hydroxide Demand By Application – 2011

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Figure 19.4 – Lithium Hydroxide Demand By Application – 2025

Figure 19.5 – Lithium Carbonate Demand By Application – 2011

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Figure 19.6 – Lithium Carbonate Demand By Application – 2025

e) Lithium Consumption Forecast: Sensitivity Analysis A sensitivity analysis has been done in order to assess different demand and price scenarios, using the following criterias: • The possibility of a deeper economic slowdown in 2012. This situation would imply a slowdown of the demand of lithium, mainly in the industries related with ceramics and glasses, automotive application, and batteries; • The penetration rate of hybrid and electric cars in the market; which depends mainly on how the cost of the battery will evolve, which represents currently the biggest cost of an electric car. Three (3) scenarios, reflecting economic growth and electric vehicle penetration rate have been considered. Table 19.3 show these three scenarios.

Table 19.3 – Possible Scenarios For Future Lithium Demand Scenario Economic Growth EV Penetration Rate Optimistic Base High Realistic Base Base Pessimistic Low Low Source: SignumBOX estimates In terms of the growth of total lithium demand, the more pessimistic scenario results in a Compound Annual Growth Rate of 8.8% (CAGR) and the more optimist scenario in a CAGR of 11.8% for the next 21 years as is shown in Figure 19.7.

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Figure 19.7 – Lithium Demand – Sensibility Analysis (tonnes LCE)

Lithium hydroxide is much more sensitive to market conditions than the other lithium compounds, because of its use as a cathode material in batteries for hybrid and electric cars. As can be seen in Table 19.4, lithium hydroxide growth would range between 12.6% and 16.8% per year. It is important to highlight that these estimations do not consider future economic recessions as well as product substitution. They only consider different scenarios for the economic growth in 2012 and different scenarios for HEV and EV’s penetration rate. It is not yet known with certainty which will be the most used lithium chemical compound in cathode material for batteries for hybrid and electric cars. It has been assumed that the vast majority of the manganese-spinel and lithium iron phosphates battery types would require lithium hydroxide as raw material while the vast majority of lithium polymer and other types of cathode materials would require lithium carbonate as raw material.

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Table 19.4 – Lithium Demand Forecast – Sebsitivity Analysis Total

Demand CAGR Scenario 2011 2015 2020 2025 2011 - 2025 Optimistic 140,056 209,765 363,040 667,191 11.8 % Realistic 140,056 204,732 327,743 540,119 10.1 % Pessimistic 140,056 180,925 293,628 455,677 8.8 %

Lithium

Carbonate CAGR Scenario 2011 2015 2020 2025 2011 - 2025 Optimistic 66,736 98,340 166,169 312,876 11.7% Realistic 66,736 95,068 149,743 253,739 10.0% Pessimistic 66,736 79,470 133,190 213,413 8.7%

Lithium

Hydroxide CAGR Scenario 2011 2015 2020 2025 2011 - 2025 Optimistic 27,533 51,321 115,860 242,929 16.8% Realistic 27,533 49,889 99,297 183,303 14.5% Pessimistic 27,533 43,879 84,777 145,688 12.6% Source: SignumBOX estimates

19.1.3 Supply a) Lithium Reserves and Resources: Description

There are two (2) main sources of supply for lithium. The main one is currently continental brines and the second one is hard rock, mainly pegmatite but also from petalite and lepidolite.

Figure 19.8 shows the actual distribution of supply between these two (2) main sources in 2011.

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Figure 19.8 – Lithium Chemicals By Source – 2011

In the SignumBOX report, the lithium minerals for direct use, representing 13% of the current supply are not taken into consideration, as it is mainly ore of technical grade. b) Lithium Reserves and Resources: Estimation There are different estimations of the amount of reserves and resources around the world. The US Geological Survey (“USGS”) updates every year its estimation of lithium reserves and resources and according to the latest publication, lithium reserves are 12.7 million tonnes (Li) while resources are estimated at 25.5 million tonnes (Li).

The Table 19.5 shows the estimations of reserves and resources established by both USGS and Mr. Keith Evans, a geologist by profession who has been involved in the lithium business since 1970. Mr. Evans has written many articles regarding lithium reserves and resources and participated in several international lithium conferences. c) Lithium Chemicals Supply Currently the chemicals production capacity around the world is estimated at about 145,000 t Lithium Carbonate Equivalent (“LCE”). Approximately 41.5% of the supply obtained from pegmatites which is used in the form of lithium concentrate in direct applications (13% of total lithium supply) and in the form of lithium chemicals (28% of total lithium supply), as shown in the above Figure 19.8.

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Table 19.5 – Lithium Reserves and Resources (Tonnes Li) USGS Keith Evans Country / Reserves and Reserves Resources Source Resources Chile 7,500,000 > 7,500,000 7,100,000 Argentina 850,000 2,600,000 2,550,000 China 3,500,000 5,400,000 3,350,000 Australia 970,000 1,800,000 1,539,800 Brazil 64,000 1,000,000 85,000 Canada 180,000 3,600,000 255,600 US 38,000 4,000,000 6,620,000 Zimbabwe 23,000 - 56,700 Bolivia 0 9,000,000 8,900,000 Portugal 10,0000 - 4,264,000 Total 13,135,000 > 34,000,000 34,721,100 Sources: USGS 2012 and Keith Evans

d) Worldwide Supply of Lithium Total lithium production capacity is currently estimated around 198,500 t expressed as LCE. According to the Companies’ information, it is forecasted that a production capacity from current lithium producers is to increase by 129,000 t in the next four (4) years reaching about 330,000 t in 2020. According to SignumBOX, it is estimated that production capacity from Newcomers in 2014-2015, will bring an additional 50,000 t of lithium carbonate equivalent in the overall production capacity. Therefore, SignumBOX estimate that by 2020 the production capacity would reach about 500,000 t of which 330,000 t will come from current producers and 170,000 t from new projects. e) Production Capacity from Newcomers Figure 19.9 shows the potential production capacity from Newcomers.

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Figure 19.9 – Potential Production Capacity From Newcomers

19.1.4 Prices

Current lithium carbonate prices are in the range of US$ 6,000- US$ 6,500 per tonne, whilecurrent lithium hydroxide prices are in the range of US$ 8,500- US$ 8,900 per tonne. Recently, the main producers of lithium compounds have announced increases of 10% to 15% of their main products. The main reason behind these increases is the higher cost of raw materials required to produce lithium carbonate, in particular soda ash. With the estimates of further production capacity and demand, it is expected that in 2025 outside of China, lithium carbonate prices would reach US$ 8,000 and lithium hydroxide prices would reach US$ 13,100.

19.1.5 Conclusions

The main conclusions of the two (2) reports received by Nemaska are: • There are a limited number of producers of lithium hydroxide presently around the World, all of them using a traditional process of transforming lithium carbonate into lithium hydroxide; • Demand for lithium hydroxide of battery grade is expected to grow at the rhythm of about 30% per annum over the next 15 years; • Lithium hydroxide expected growth demand is mainly related to secondary batteries use over the next years;

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• Overall lithium hydroxide demand is expected to grow at a rate between 12.6 % to 16.8 % depending on the rate of penetration of the electric and hybrid vehicles. 19.2 Contract

Nemaska has signed on October 3, 2012, an off-take and collaboration agreement (“Agreement”) with Phostech Lithium Inc. (“Phostech”), a Clariant AG group company. Per the Agreement, Phostech agreed to evaluate and purchase from Nemaska the output of the lithium hydroxide monohydrate to be produced by Nemaska’s Phase I plant. Nemaska and Phostech also agreed to collaborate in order to determinate the economic and technical feasibility of tailoring the lithium hydroxide fabrication process according to Phostech’s specifications. Pending financing and budgetary considerations, Nemaska Lithium intends to build a Phase I, 500 tonnes a year lithium hydroxide plant using its patent pending chemical transformation process. Phostech recently opened its facility in Candiac, Quebec that produces the Life Power® P2 C-LiFePO4, a safe, high-power cathode material for rechargeable lithium-ion batteries used in hybrid and electric vehicles, solar and wind energy storage and power tools.

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20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

All the information presented in this section concerns the mine site, except Section 20.7.

An Environmental and Social Impact Assessment (“ESIA”), including an Environmental and Social Baseline Study (“ESBS”), for the Company’s Whabouchi Project is ongoing at the mine site.

The ESBS will present information on the physical, biological and social environment for the mine study area, which has been defined to cover an area of approximately 300 km2.

20.1 Environmental Studies

Baseline environmental studies at the Whabouchi site began in August 2010 with field surveys for water quality, sediment quality, benthic invertebrates, and fish. During 2011 and through 2012, additional data were collected, focusing on fish, surface water quality, bathymetry, hydrology, ground water quality, soil quality, air quality, noise, large mammals, small mammals, bats, birds, amphibians, and reptiles. Figure 20.1 shows a list of completed, ongoing, and planned work for the environmental baseline.

Table 20.1 – State of Studies for the Whabouchi Natural Environment Baseline (July 2012) Environmental Ongoing or Planned Work Undertaken to Date Component Work Establishment of a GIS for the Project and Continuing updating of Cartography incorporation of existing databases (topographic GIS with data from base, lidar, etc.), desktop work, and field data. field surveys. Collection of data from existing weather stations, One (1) year of including Nemiscau A (21 km) and La Grande meteorology data to be Climate Rivière A (240 km) amongst others. Interpolated developed using either climatic design data obtained from Environment Global Analysis Data. Canada. Extensive series of hard rock geology drilling and analyses by Nemaska Lithium described elsewhere in this Report. Particle size analysis of surficial deposit samples Continuing chemical analyses of samples of Geology from trenching. rocks, soils and surficial Samples of soil and surficial deposits collected, in deposits. part during exploratory trenching (21 trenches dug) for chemical analyses. Geomorphology interpretation and field validation.

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Environmental Ongoing or Planned Work Undertaken to Date Component Work Twelve (12) drill holes converted into observation wells for groundwater quality. Drawdown tests. Continuing analyses of Hydrogeology Eleven (11) permeability tests. groundwater samples. Collecting and chemical analyses of groundwater samples. Field check of non-mapped watercourses Determination of watershed boundaries Hydrology Field flow measurements. Calculation of theoretical flows from CEHQ data. Collection of bathymetric data for three (3) areas

and production of bathymetric maps. Water and Samples taken and analysed. sediment quality Air quality filters to be Air quality Air quality monitoring station installed. analysed. Noise data to be Noise Noise monitoring station installed. analysed. Aquatic biology Collection of data on habitats. – habitats Aquatic biology – benthic Analysis of benthic invertebrates at six (6) stations. invertebrates Experimental fishing at all local lakes and streams (by net and electrofishing). Aquatic biology

– fish Hundred (100) fishes preserved for analyses. Forty (40) fishes from six (6) stations were analysed chemically. Terrestrial Detailed botanical work including vegetation biology - mapping and identification of important vegetation communities and species present. Terrestrial Field work determining the boundaries of wetlands biology – in the field. wetlands Terrestrial Amphibian and reptile survey conducted including biology – call recording Analysis of recorded amphibians and Incidental observations while conducting other field calls ongoing. reptiles work.

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Environmental Ongoing or Planned Work Undertaken to Date Component Work Breeding bird survey. Terrestrial Analysis of recorded biology – birds Incidental observations while conducting other field calls ongoing. work. Aerial survey for large mammals in winter. Track counts in snow for large and medium Difficult specimen sent mammals in winter. to government biologist Terrestrial for confirmation of biology – Trapping survey for small mammals. identification. mammals Ultrasound recordings for bats. Analysis of recorded Incidental observations while conducting other field calls ongoing. work. Species of Information requests to agencies such as the

special status CDPNQ and MDDEP and field surveys. Species of Preliminary species list established and presence

traditional use. confirmed during field work. 20.1.1 Study Area

Two (2) study areas have been defined for the environmental studies (Figure 20.1). The local study area includes all of the areas likely to be directly physically impacted by the mine development (pit, buildings, and roads). The regional study area is a larger area extending out 10 km from the pit. This larger study area encompasses areas likely to have increased sound levels or dust deposition resulting from activities at the site as well as areas that may be visually impacted.

The following sections present a portrait of the natural environment at the site. It is as yet incomplete, since a number of field studies and analyses are still to be completed. Also, subjects which have been covered in other sections are not presented here (i.e. geology, physiography, etc.).

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Figure 20.1 – Regional and Local Study Area

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20.1.2 Physical Environment a) Surface Water and Groundwater Environment Both the local and regional study areas are within the Rupert River drainage. The Rupert River has a watershed of 43,400 km2 and runs into the southern part of James Bay. All of the small watersheds near the site drain into the Nemiscau River (2,000 km2 watershed upstream from the site) itself, a tributary to the Rupert River that it joins 70 km downstream from the site. The Nemiscau River is on the western perimeter of the site and the on-site drainage trends west-southwest towards the Nemiscau River. The largest stream on-site is the one that drains Spodumene Lake with a watershed of over 100 km2. The other on-site watersheds are well under 5 km2.

The largest lake in the vicinity of the site is Lac des Montagnes, with a surface area of 1,375 ha. It is located on the western perimeter of the Whabouchi Property and is actually a widening of the Nemiscau River. Spodumene Lake, to the east of the site, is only 61 ha.

Calculations of stream flow rates derived from standard CEHQ models have been made and supplemented with field checks.

A preliminary analysis of the surface water quality data to date indicates that the waters on and near the site are typical for this type of environment. The pH values range from 5.0 to 8.0, depending on water body and depth. Some of the waters are somewhat acidic, common for Shield water bodies. The lowest pH values are for a small lake that is in the process of turning into a bog, again a common value for this situation. Most of the chemical analyses show values under the detection limits, common for poorly mineralized oligotrophic waters. The only exception appears to be the small lake turning into a bog that has normal higher levels of carbon. Alkalinity and calcium levels are low in most stations, indicating some sensitivity to acidification. Most metals are present at levels below the detection limits but some measurable levels of aluminum, iron, chromium, lead, and zinc occur.

A preliminary analysis of sediment quality data taken to date in the lakes and watercourse suggests that they are generally of good quality with some elevated copper values.

Groundwater has been analysed at several well locations and preliminary evaluation suggests that it is generally of good quality. Other analyses are planned.

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20.1.3 Biological Environment a) Aquatic Biology As is typical in northern environments, the aquatic systems are low in nutrients and the growth rate of aquatic organisms is slow because of the low temperature.

Benthic invertebrate studies were undertaken in 2010. Benthic invertebrates are important indicators of the quality of aquatic habitats. Results show a great deal of variability between sites.

Thirteen (13) fish species have been found in the lakes and streams near the site during fieldwork from 2010 to 2011. The species caught most frequently is the lake whitefish. Other common species include white sucker, walleye, and brook trout.

Several of the fish species present are caught and eaten by local residents, notably brook trout, walleye, lake white fish, and pike. In November 2011, over 100 sample fish were taken and preserved for chemical analyses. Preliminary evaluation of the results of chemical analyses of muscle tissue from 40 of these fish suggests that levels of metals are generally low. Mercury levels were above detection limits in all samples. The mercury levels were consistently higher in the lake whitefish samples from Lake Spodumene.

To date, no aquatic species of special status has been found on the site. Lake sturgeon is a species considered as likely to be designated threatened or vulnerable in Quebec (Espèce de la faune susceptible d'être désignée menacée ou vulnerable) and is present in some nearby lakes (i.e., Nemaska Lake) and in the Rupert River system. However, there are no indications of its presence on the Nemiscau River drainage this far up. Local people have stated that they have not found it in this area. b) Terrestrial Biology The site is in the boreal forest vegetation zone that covers much of northern Quebec; more specifically, it is located in the Western spruce-moss bioclimatic zone. Black spruce forests, more or less open, dominate but around 80 percent of the local study area has been burned in the very recent past and is in various stages of regeneration. Forest fires are a key element in the evolution of the vegetation in this area. Wetlands, primarily bogs but also some marshes, cover a significant percentage of the study area. There is a large wetland south of the proposed pit.

The site is well beyond the northern limit of commercial forests according to the MDDEP. However, some plant species of traditional use are likely to be present on the site. No plant species of special status has been found on the site.

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The animal species found in the vicinity are typical of the boreal forest and include moose, black bear, otter, marten, and wolf. There have been field surveys for mammals, birds, reptiles, and amphibians.

One (1) species of particular interest is the woodland caribou, a subspecies of caribou. Two (2) ecotypes of this subspecies may be present at the site. The forest ecotype is present in small numbers throughout the year in this part of Quebec, although the area around the site has not been identified as of particular importance by the Quebec Government. The other ecotype is the tundra ecotype. In the summer, it is present in large herds in the more northern parts of Quebec. In the winter, these herds move down into the more wooded areas in smaller bands. Most of the tundra ecotype caribou remain north of 52 degrees latitude, but radio telemetry data shows that some individuals of the Rivière aux Feuilles herd do come down to the vicinity of the site. Woodland caribou of the forest ecotype are considered vulnerable in Quebec; however, caribou were not sighted during the February 2012 aerial surveys, nor during any of the other field work.

There is a prior record of one (1) sighting of a hoary bat near Spodumene Lake. This species is considered as likely to be designated threatened or vulnerable by the Quebec Government. Analyses of ultra sound recordings taken in June-July 2012 are still ongoing. This is a species that needs trees for roosting and the recent forest fires may have made the local area less attractive for this species.

One (1) bird species, the common nighthawk, listed as “threatened” under the Federal Species at Risk Act, is present at the site and probably breeds on the site. It is a widespread and typically fairly common species but which has seen a marked reduction in population throughout most of its range, hence it being put on the Federal list. Mitigation measures regarding this species will be included in the impact analysis.

The site is not in any protected zone for terrestrial plants or animals, as designated by the Quebec Government.

20.1.4 Water, Rock and Tailings Characterization

The principal lithological units that comprise the deposit and are likely to end up in the waste rock and tailings pile include gabbro, basalt, and pegmatite. The geochemical classification of the lithological units was done using static tests and kinetic tests are actually ongoing. Samples of drill core and mineral processing rejects were tested following the characterization program proposed by the MRNF/MDDEP (Directive 019 sur l’Industrie minière, 2009). The overburden will also be tested soon but will go through a chemical analysis only.

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The geochemical classification program had three (3) objectives: • Classify the waste rock and the tailings according to the MRNF/MDDEP standards; • Identify the surface disposal requirements; • Identify chemicals that could potentially affect future surface water quality. Preliminary results done on 83 samples of rock taken in the drill holes show that some of them contain a small amount of sulfides with no neutralisation potential. The average values of all samples show that the amount of sulfides should not be enough to start any acidic drainage over the time of weathering exposure. However, kinetic tests are ongoing to ensure that the future waste rock will never oxidise in a way that would produce acidic water. Static tests done on the tailings showed that they are non-acid generating. The tailings will be co-disposed with the waste rock by layered co-mingling. Because the tailings will be filtered and then transported by truck and compacted in the waste pile, their hydraulic conductivity will be low. This will have an impact on the infiltration of water within the pile and the movement of oxygen also. This impact is positive and should limit the oxidation of sulfides, if any, and the transport of chemicals in the environment.

Samples of waste rock and tailings have also been tested for their leachability. Three (3) protocols were done (TCLP, SPLP and Shake Flask Extraction). Results showed that most of the samples leached during these static tests. Kinetic tests are ongoing. These are more representative in terms of probably leaching behavior and leachate concentrations generated from waste rock and tailings that will be exposed to weathering conditions.

20.1.5 Run-Off and Mine Water Management

The water management plan is developed to collect, monitor and treat, if required, all contact water from the mine site.

Water will be collected only from the three (3) zones of operation, namely milling and mine infrastructures, waste rock and tailings pile, and open pit. All other run-off water will follow the natural watersheds around these three (3) operation zones and will not be allowed to enter the drainage structures by protecting them with berms.

In general, peripheral berms will prevent run-off water to enter the various drainage structures and mix, dilute, or segregate with mine effluents (compliance with Quebec Directive 019 for the Mining Industry, article 2.1.5).

Surface water from the milling and mine infrastructures area will be collected in ditches into local sedimentation ponds. The treated water will then be recycled for use in processing. Peripheral berms around the sedimentation ponds and ditches will prevent run-off water from entering these structures and, as a result, will increase the recycled

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water volume. Water from this source shall not be released in nature. Only suspended solids are expected in the mine water.

The run-off water from the waste rock and tailings pile and open pit will be collected in ditches and then will flow to sedimentation ponds where water will be treated for suspended solids. Then the overflow will be discharged in nearby streams and/or lakes.

Since preliminary results do not indicate acid generation, construction of impermeable drainage structures (except for the dikes of the sedimentation ponds) is not necessary. The native soil is generally fine silts and sands of relatively low permeability, of thickness ranging from 0.5 to four (4) metres and sitting on rock. Infiltration of water from the ditches and sedimentation ponds towards near-by water bodies will be slow (order of four (4) metres/year) and any fines in the drainage water will be captured by the native soil.

Progressive vegetation of the waste rock and tailings pile will reduce the release of fine materials. Further sedimentation control will be provided along the length of the drainage ditches by introducing various types of small catchment basins (by enlarging and/or deepening the ditches) and silt traps (rip-rap mounts, hay balls, etc.).

The waste rock and tailings pile and open pit sedimentation ponds are designed with low impermeable dikes and will be sufficient to contain the 100-year return of precipitation and snow melt flood. Under “normal” flow conditions, the water will be given sufficient time to sediment in the pond before exiting through a culvert positioned in the dike above a predetermined elevation such as to provide sufficient water residence time required for sedimentation of fines. In the case of an “extreme” precipitation event, the water will surpass the elevation of the culvert to a maximum of one (1) metre below the crest of the dike. This water will subsequently exit the pond through the culvert, which is designed with a gate to control the output flow rate.

The water collected from the waste rock and tailings pile and open pit will eventually reach the Lac des Montagnes in two (2) distinct locations: on the south side of the waste rock and tailings pile (south-west limit of the proposed mine site) through an existing stream and south of the open pit (south-east limit of the proposed mine site), respectively. Channels, armoured with geotextile and rip-rap, are designed to carry the water from the sedimentation ponds to either the creek or the lake. These channels become wider at their exit such that the flow rate and corresponding water velocity are further minimized.

Armouring the “exit” channels will prevent erosion as well as contamination with fines of the previously “polished” water.

In the long term, the permeability of the bottom materials of the ditches and ponds will decrease as fines settle as sediment from the drainage water.

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A minimum freeboard of one (1) metre has been allowed in the design of all drainage structures.

Outflow water from the sedimentation ponds (waste rock and tailings pile and open pit) will be regularly sampled and tested for compliance with Quebec Directive 019 for the Mining Industry (article 2.1.1.2).

“Normal” operation conditions water balance estimates were based on the mean Canadian Normals for the Nemiscau airport. It is assumed that there are no flows in the period of December to March as almost all precipitation is snow and the air temperature is below zero. However, during spring, the snow melt is incorporated in the precipitation run-off as the difference between snow melt quantities between two (2) consecutive months (snow on the ground at month end) until there is no more snow on the ground.

For “extreme” operation conditions, the 100-year return 24-hour rain, plus a 30-day 1:100-year snow melt is employed in the estimation of water run-off volumes.

Table 20.2 – “Normal” Operation Conditions Water Flow Estimates Minimum Water Water Flow Zone Retention Time Remarks (m3 /day) (days) Milling and Mine 44 56 – 189 Water recycled Infrastructures Waste rock and Water released to 21 349 – 1,642 tailings pile nature Water released to Open Pit 23 491 – 954 nature 20.2 Social Studies

Information pertaining to local demographics, economic development, land use, cultural heritage, health and social services, and infrastructure has been collected in order to provide a snapshot of the community’s needs and priorities and to determine how current conditions may be affected by the proposed Whabouchi Project. Results are based largely on data obtained from the local First Nations government (Cree Nation of Nemaska Band Council), local social service providers, educational institutions, law enforcement, and Census Canada, as well as on data obtained from engagement with land users and community stakeholders. Socio-economic data from various Cree agencies (Cree Regional Authority entities) and provincial authorities will continue to be collected and analyzed in order to complete the baseline assessment.

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20.2.1 Socio-Economic Baseline of Nemaska Cree Community

According to the 2006 Canada census data, the Cree community of Nemaska has a population of 642 people. However, estimates from the Nemaska Band Council administration personnel report rapid population growth and the actual population now stands closer to 800 people. The mean population age in 2006 was 25 years old. Elders represent the smallest age group at 4% of the population.

Local economic opportunities in Nemaska have increased over the past several years. Contributing factors include Hydro-Québec contracts being awarded to local contractors for construction projects and environmental monitoring, as well as new community infrastructure projects undertaken by local government and associated economic spin- offs. Taking into consideration the large youth population in Nemaska, the Economic Development branch of the local government is currently rolling out various training and skills development programs (heavy machinery operation, for example) which will play a key role in building local capacity to meet future needs.

A recent development in the community is the emergence of employment at the entrepreneurial level. Individuals are offering such services as equipment rentals (dump trucks, snowmobiles), video rentals, painting, towing and equipment maintenance and repairs. As well, the new Nemaska clinic, band council headquarters and justice facilities are in the process of being erected. However, in spite of efforts to stimulate the local economy, people in the community still report a lack of jobs.

Social service facilities in the community include the wellness centre, Nemaska clinic, social services centre, school, daycare, multi-service centre and youth centre. The wellness centre, social services centre and Nemaska clinic provide those social service programs with the farthest reach within the community. Programming focuses on mental and physical health promotion at large, and on generating awareness around such specific issues as diabetes, substance abuse, bullying, conjugal violence, sexual health, suicide, anxiety, and depression. Resource workers who occupy management positions at these facilities report being without adequate resources to accomplish the work required of them (lack of trained workers, time and funding).

There is very little incidence of violent crime in Nemaska, however assaults and domestic violence do exist. The majority of the crimes dealt with by the police force are alcohol related, with the incidence of crime being higher in the summer. Drunk driving was cited as one (1) of the main criminal activities within the community. Out-of-town employees staying at temporary worker camps are thought to exacerbate alcohol-related incidents.

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20.2.2 Land Use

The Project is located on part of the territory of the Cree Nation of Nemaska, one (1) of the nine (9) James Bay Cree communities, more specifically on the Cree hunting ground known as Trapline R20.

The land use study covers a range of topics that allow a thorough description and understanding of Cree uses of the territory and its resources over a period of approximately two (2) generations. It also includes projected land use by Cree families traditionally associated with the area.

The general description of Cree land use practices is partly based on a literature review and data collection from the community of Nemaska. Land use studies have been conducted in the community of Nemaska in the recent past, namely in the context of Hydro-Québec’s environmental impact assessment for the EM1-A Rupert diversion, which affected the same Cree families. For this reason and to avoid duplication of effort and undue pressure on Cree land users, authorization was obtained from relevant authorities to use results from these past studies.

An important part of the consultations regarding land use consists in conducting in-depth semi-structured interviews with Cree users of hunting grounds located in the vicinity of the Project. A special emphasis is put on consultations with the family of Trapline R20, which will bear the majority of disturbances and infrastructures related to construction and operations of the mine.

Focus groups are also carried out to obtain further community input on the Project, perceptions of impacts and benefits, and to document concerns and expectations.

Efforts are made to create and maintain a collaborative and cordial relation with the community, in particular with families affected by the Project, in order to address issues as they emerge throughout the course of consultations and Project development.

20.2.3 Heritage Resources

An initial archaeological survey has been completed in 2011 and further field work done during the summer of 2012. However, at this time, no significant issues have emerged that would be considered jeopardizing to the Project.

20.2.4 Involvement of Chibougamau Community

Since 2009, the company has engaged with the mayor and executive council of Chibougamau to inform them of the mining project and its potential impacts on the community. Nemaska Lithium and the municipal leadership have since developed an engagement and communication program that is appropriate for anticipated impacts that

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the Project will have on the community. Although the impacts anticipated at this time are mainly positive, ongoing engagement should identify any new potential impacts that have yet to be assessed. Broad community consultation is planned to take place in 2013.

20.3 Permitting

20.3.1 Environmental Assessment and Review Processes

Because of its northern location, the Whabouchi Project falls under the environmental protection regimes of the James Bay and Northern Quebec Agreement (“JBNQA”) and could possibly fall under the Canadian Environmental Assessment Act. The purpose of the Environmental Assessment is to allow the relevant Quebec and federal regulators to properly assess the impact of the Project and to seek input from local stakeholders on the proposed development. In this context, the Whabouchi Project notice was submitted in July 2011 to the James Bay Convention Administrator, Mrs. Diane Jean, sous-ministre from the Ministère du Développement durable, de l’Environnement et des Parcs (“MDDEP”). Following the Directives to be issued by the Evaluation Committee (“COMEV”) via the MDDEP and possibly the Canadian Environmental Assessment Agency (“CEAA”), an ESIA will be completed. Subject to a schedule to be established by the Review Committee of the JBNQA (“COMEX”), and possibly the CEAA, it is currently expected that public hearings will be held, and, upon the satisfaction of all regulatory requirements, the Project will be eligible for the receipt of Quebec Certificates of Authorization. Once the provincial (and federal if necessary) administrators have issued authorizations for Project development, final mine permits will be sought from the MDDEP, the Ministère des Ressources naturelles et de la Faune (“MRNF”), and all relevant federal authorities.

20.3.2 Permitting

Apart from the ESIA, which will lead to the delivery of the global certificate of authorization, many other authorizations are required from municipal, provincial, and possibly federal governments.

At the municipal level, the Project is on the territory of the Municipality of James Bay and a certificate of conformity to their regulations will have to be issued each time a certificate of authorization is requested under sections 22 and 32 of the Environment Quality Act (c. Q-2). Construction permits will also be required before starting any construction work.

At the provincial level, many certificates of authorization will have to be delivered from the MDDEP in accordance with the Environment Quality Act (c. Q-2).

During the construction and the operation, many permits and leases will be required. Nemaska will present the requests when required.

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From the federal point of view, if the Project does not fall under the Canadian Environmental Assessment Act, no specific permit will be required for the construction of the mine.

20.3.3 Status of Permit Applications

The permit application process for the Whabouchi Project was initiated at the end of July 2011 when Nemaska sent the proponent’s preliminary information and Project notice. Nemaska has received the guidelines from the COMEV in February 2012.

20.4 Anticipated Impacts

20.4.1 Environmental

The impact predictions contained in this section are preliminary, subject to modifications in the mine layout and process and the results of ongoing fieldwork and modelling.

The Whabouchi Project has a major positive feature in terms of impact on the natural environment: it is located on an existing road and has nearby support facilities, including an airport. Besides the costs and impacts directly associated with these constructions, there are also long-term impacts from the road, often much longer than the operational life of the mine. The most profound impact of a new road is that it allows ground access to areas which were formerly remote, leading to increased use. The road itself may also lead to habitat fragmentation. This may have long-term effects on the natural environment, primarily on larger mammals. The habitat fragmentation impacts of the Whabouchi Project are lessened because of its location on the Route du Nord.

Many of the impacts associated with the Whabouchi Project are limited to the period of operation of the mine and impacts cease with the mine closure. The development of a good mine closure plan will considerably reduce any residual impacts.

The most evident impact of a mine development on a natural site is that a certain amount of land is removed from natural habitat and is converted into active industrial use for the duration of the mining activities, in this case roughly 25 years. This includes the open pit, the associated facilities, any on-site roads, and the waste rock and tailings pile. For the Whabouchi Project, the estimated area involved amounts to 4 km2 (2 km x 2 km).

The nature of the deposit essentially eliminates the possibility of an underground mine, an option that could reduce the amount of habitat converted. The body of the mineralized material dictates the location and size of the pit itself. Other spatial aspects of the facility have been developed taking environmental concerns into consideration. In particular, nearly all lakes, streams, and wetlands have been avoided. There is only one crossing of a stream by an on-site required road and it will be done following all applicable measures to reduce impacts to aquatic life. The areas to be cleared consist of terrestrial vegetation;

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primarily spruce woods regenerating following recent fires. Some ephemeral watercourses may be eliminated by the waste tailings, something practically impossible to avoid in an area with a drainage network as dense as here. Choosing an optimum time and method will minimize the impact from the clearing and reduce the impact on animals unable to move (i.e., nesting birds). The closure plan will include measures to help bring back the affected areas as useful habitats. The open pit will gradually be filled with water.

The surface water flow near the mine site may be affected in two (2) ways: by the diversion of surface water flow in specific watersheds and by mine dewatering. The surface flow diversions are associated with the mine water management system, which ensures that water coming off of the waste rock and tailings pile, and other facilities are released at two (2) outfalls. This will impact particularly two (2) small watersheds, watersheds so small as to be considered to have only intermittent watercourses associated with them. The ESIA will document the anticipated changes and the potential impacts. Groundwater entering the open pit will have to be pumped to a sedimentation pond. This will tend to depress the water table and could affect water levels in adjacent wetlands, streams and lakes. Further hydrogeological studies are underway to estimate the magnitude of this drawdown effect.

The quality of the surface water in the area of the mining activities could be affected. Studies to quantify and model these potential effects are underway.

Air quality at the mine site could be affected as the result of dust in the air resulting from traffic and mining activities. The air quality impacts will be modelled. The increase of noise levels as a result of the Project will be modelled and the anticipated effects described.

The use of the road to Chibougamau to haul concentrate will result in some increase of noise and dust along the road.

The Project will also be assessed in terms of cumulative impacts. There are few other projects in this area at this time so that the cumulative effects are likely to be limited. However, it will be important to consider these effects carefully because of the remote nature of the area.

20.4.2 Social

The construction, operation, and closure of the mine will alter the physical landscape and will consequently affect land use within the immediate areas. The mining Project will also increase direct, indirect and induced employment and thereby impact the community and regional populations. These changes will affect the communities of Nemaska, Chibougamau, and, perhaps to a much lesser extent Chapais and other communities further afield, both positively and negatively. Public consultations are underway and aim

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to provide a more meaningful assessment of the scope and weight of anticipated impacts. Impacts identified to date or anticipated are summarized below. a) Land Use Impacts The mine site is within the territory of Trapline R20 and the Project will lead to a temporary loss of some hunting and trapping area within R20 because of the mine’s footprint. The mine and its activities will impact resource harvesting temporarily through habitat loss and perhaps disturbance.

There are some campsites, which may include cabins or tent frames, on or near the mine site that will also be impacted.

At the time when this Report was prepared, sites of shared or family heritage (sites imbued with memory and history) have been identified on or near the mine site. Undoubtedly the integrity of these sites and the overall landscape will be impacted. However the degree to which specific sites of importance will be impacted requires further study (new meeting with the users to clarify certain information and determine mitigation measures).

It is perceived by residents that use of the beaches, the bible camp, and possibly the waterways in between, which are nearby and/or within view of the mine, will be negatively affected by the mine’s visual and audible impacts. The extent of these impacts is being further assessed and is the focus of ongoing consultation. The beaches and bible camp are sites valuable to the maintenance of community cohesion. Impacts on these two (2) sites may hinder their use for the potential pursuit of outfitting, tourism and other income generating activities.

The Project’s overburden, waste rock and tailings pile, as well as its noise, will decrease the aesthetic and wilderness value of the landscape to Cree and other land users and create direct visual impacts that may degrade the value and appreciation of certain sites (i.e., the bible camp).

Disturbance from noise will be temporary (i.e., the operating life of the facility), however the visual disturbance from the waste rock and tailings pile will be longer- term. The visual impacts may be mitigated by careful design and re-vegetation. b) Social Impacts The Project will provide training, employment, economic development opportunities, and revenue. These positive impacts will occur at the local level (Nemaska), the regional level (Chibougamau), and the provincial and national levels. These impacts will likely be felt most strongly in Nemaska through the training and employment of local residents and the local procurement of goods and

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services. Chibougamau may also see marked positive impacts as the road haul terminal.

Along with these positive impacts, one must be aware of potentially negative impacts. For example, the study team will be further assessing the potential for these largely positive impacts to increase tension and conflict among Nemaska community members, especially with regards to those who are perceived beneficiaries versus those who are not (i.e., impacting community cohesion).

It is also possible that the increased presence of non-Cree people near Nemaska and the increased frequency of workers leaving and returning to the community may exacerbate existing social problems or create new social issues. For example, the more frequent movement of people in and out of the community may increase the availability of drugs and alcohol, which may thereby exacerbate the social issues associated with their use and abuse.

In terms of community well-being, Nemaska community members have indicated increased social and individual anxiety stemming from fear of Project impacts, an increased sense of exploitation and decline in cultural self-esteem, increased feelings of loss related to the modification of the traditional landscape, and increased feelings of confusion related to the clashing of Cree and non-Cree values.

The population growth resulting from the increased direct, indirect, and induced employment attributed to the mine and its transportation system may put pressure on the Nemaska and Chibougamau community services and infrastructure, depending on whether they are at or near capacity and whether this population growth far exceeds the communities’ planning projections.

20.5 Closure Plan

20.5.1 Introduction

In accordance with the law on mines (L.R.Q., chapter M-13.1, section III, article 232, august 2012) a detailed closure plan must be submitted to the Ministère des Ressources Naturelles et de la Faune (“MRNF”) prior to getting the Global Certificate of Authorization. The Guide and method of preparation for the plan as well as the general requirements for the restoration of Quebec mining sites (Guidelines for Preparing a Mining Site Rehabilitation Plan and General Mining Site Rehabilitation Requirements, 1997, hereafter referred to as Guide), prepared by MRNF in collaboration with the Ministère du Développement durable, de l’Environnement et des Parcs (“MDDEP”), must be followed when preparing the closure plan for a mining site.

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are planned during operation, and if circumstances permit, a description of the restoration and redevelopment planned at the end of the mining operations, an evaluation of the restoration related costs, as well as a description of the financial guarantee provided for restoration of the accumulation areas.

The financial guarantee must cover 70% of the costs related to restoration of the accumulation areas. According to the Guide, accumulation areas include sedimentation ponds, waste rock piles, tailings area, and storage areas of mineralized material concentrate. The goal of restoration is to rehabilitate the site in a satisfactory condition, by: • Eliminating unacceptable health hazards and ensuring the safety of persons; • Limiting the production and the spread of substances that are likely to affect the receiving environment and, in the long term, aiming to eliminate all forms of maintenance and monitoring; • Restoring the site to a visually acceptable state for the community; • Restoring the infrastructures site (excluding accumulation areas) in a state that is compatible for future use. 20.5.2 Revegetation

Several measures will be taken with regards to the closing works for the mine. In this regard, the site of the mine infrastructures, the waste rock and tailings pile, the sedimentation ponds, and any roads which are not useful for post closure monitoring operations, will be restored and covered with vegetation.

A 0.3 metre layer of overburden will be spread on the relatively flat surfaces before revegetating them.

The overburden will be used initially to restore the waste rock and tailings pile and other industrial areas. Revegetation of the waste rock and tailings pile will be done progressively during operation of the mine.

20.5.3 Soil Decontamination

The Regulation on Land Protection and Rehabilitation (L.R.Q., c. Q-2, a. 31, 31.69, 109.1 et 124.1 in effect since March 2003), aims to provide increased protection and land rehabilitation in case of contamination, making several provisions applicable to the new section IV.2.1 of the Act on Environmental Quality (sections 31.42 to 31.69), enacted by Article 2 of Chapter 11 of the statutes of 2002. Moreover, the very definition of land includes the surface water and groundwater. The regulation states the limits for a variety of contaminants and specifies the types of commercial and industrial activities covered including the mining industry.

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The mining company must then proceed with the characterization of the industrial site according to the regulation on Land Protection and Rehabilitation within six (6) months after the end of the operation. In cases where contamination exceeds the criteria established by the regulations, a rehabilitation plan that specifies the protective measures to be put in place, must be submitted for approval to the MDDEP. Note that the waste rock and tailings pile is not subject to this regulation.

The characterization study will identify areas that were likely to have been contaminated by industrial activities and in particular, by the handling of petroleum products.

20.5.4 Dismantling of Buildings and Infrastructures

The walls of the buildings will be dismantled and the foundations will be covered with overburden and will then be revegetated. Wherever possible, the surface infrastructures will be built with the objective of maximizing the reuse of materials during decommissioning.

A program of waste management will be implemented in order to minimize the disposal of residual and demolition materials.

Underground storage tanks and pipes will be removed.

The work areas and parking lots will be covered with overburden and revegetated.

Based on past experience for similar projects, a substantial part of the decommissioning costs of infrastructure will be covered by the sale of equipment and materials themselves (structural steel, copper, aluminum, etc.).

It is estimated that the mine will have an approximate volume of 7,000 m3 of non- recoverable material that will have to be disposed. Because these materials are inert and that they represent a considerable volume, a solids disposal site will be created at the waste rock and tailings pile. The disposed materials will be covered with materials already present on site such as sand and waste rock.

20.5.5 Waste rock and Tailings Pile

The waste rock and tailings pile will be designed to meet the minimum requirements in terms of stability and also, to meet the requirements as defined in the Guide.

The waste rock and tailings pile will be designed so as to enable progressive revegetation during the life of the Project. Revegetation will be completed at the end of operations. The waste rock and tailings pile will be built with flatter slopes than the minimum stable slopes determined from stability analyses performed for the most critical sections. Thus, no stabilization work will be required at closure.

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According to static tests, most of the waste rock and tailings samples that were tested leached during these tests; kinetic tests, which are more representative, are ongoing (see Section 20.1.4 for more details). Nonetheless, as a precaution, perimeter ditches will be constructed to collect the seepage and drainage catchment areas within its site boundaries. The water will then be routed to a sedimentation basin to be polished (removal of fines by sedimentation) and then directed to the Lac des Montagnes.

It is estimated that 36% of the waste rock and tailings area will be revegetated as only the relatively flat surfaces (benches and top surface of the pile) can be revegetated. Given all the activities planned during operations, over 40% of the area occupied by these activities will be restored at the end of the operation period.

20.5.6 Open Pit

Following the closure of the mine, the open pit will be filled gradually by groundwater and precipitation. The access to the pit will be permanently closed by constructing an embankment made of waste rock removed from the waste pile. A ditch around the embankment will intercept surface water from entering the pit.

For the time being, filling the pit is being considered by stopping the pumping system. A hydrogeological study indicates that the water level will stabilise at elevation 288.5 metres and it will overflow to the wetlands on the south-west side of the open pit.

20.5.7 Mining Effluent

At the conclusion of the operation phase, the processing unit and water control structures will remain in operation until the MRNF/MDDEP criteria are met. Then the ditches will be filled with their own excavation material used as protective berms during construction of the mine and revegetated. The dikes of the sedimentation ponds will be breached to allow free circulation of natural surface run-off water. The crests of the dikes and any relatively flatter slopes will be revegetated.

20.5.8 Other Installations

Sanitation, petroleum products, waste, and hazardous waste facilities will be dismantled in accordance with the regulations and the laws currently in place. The same applies for the residues associated with such facilities.

20.5.9 Emergency Plan and Monitoring Program

The closure plan will include the required information regarding the emergency plan and the five-year minimum monitoring program required (integrity of the works, environmental and agronomic monitoring/follow-ups).

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20.5.10 Timing and Economic Considerations

Timing and economic considerations for the proposed mine site restoration will also be covered in detail in the closure plan.

Table 20.3 summarizes the costs estimated for restoration of the Whabouchi mine site including the post closure monitoring. The estimated cost for the restoration of accumulation areas (Table 20.4) totaled $1,441 million. These costs represent only a fraction of the estimated costs that Nemaska Lithium will require to invest for closure of the mine site.

Table 20.3 – Total Restoration Costs Type of Work Cost (CAD$) Stationary Equipment Dismantling or demolition of stationary equipment 700,000 Mine maintenance facilities and plant installations Dismantling or demolition of buildings 700,000 Construction materials debris burying 74,800 Fill drainage ditches and basins (with the protective berm material) 63,000 Overburden transport and installation 135,000 Revegetation 82,500 Open-Pit Restoration Fill drainage ditch (with the protective berm material) 18,000 Protective embankments with ditches 480,000 Closure and securing the open pit 25,000 Waste Rock and Tailings Pile Overburden transport and installation 747,000 Fill drainage ditches (with the protective berm material) 90,000 Revegetation 456,500 Sedimentation Pond 1 (Waste Rock and Tailings Pile) Dam breaching 30,000 Overburden transport and installation (crest) 6,300 Revegetation (crest) 4,000 Sedimentation Pond 2 (Open-Pit) Dam breaching 30,000 Overburden transport and installation (crest) 6,000 Revegetation (crest) 3,500 Respect of environmental regulations (5-year follow-up) 550,000

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Type of Work Cost (CAD$) Characterisation of potentially contaminated sites and decontamination of 600,000 soils Filling and disinfecting septic tanks 25,000 Total 4,826,600

Table 20.4 – Costs for Restoration of the Accumulation Areas Type of Work Cost (CAD$) Waste Rock and Tailings Pile Overburden transport and installation 747,000 Fill drainage ditches 90,000 Revegetation 456,500 Sedimentation Pond 1 (Reject Materials Pile) Dam breaching 30,000 Overburden transport and installation 6,300 Revegetation 4,000 Sedimentation Pond 2 (Open Pit) Dam breaching 30,000 Overburden transport and installation 6,000 Revegetation 3,500 Respect of Environmental Regulations (5-Year Follow-Up) 550,000 Total 1,923,300 According to the Mining Act currently in force, the amount of financial guarantee required presently is equivalent to 70% of the anticipated restoration cost for the accumulation areas, which is $1,346 million. The guarantee is expected to increase to a 100% of the anticipated restoration cost for the accumulation areas, or $1,923 million, after the Mining Act is revised and effective. The annual guarantee estimated payments are presented in Table 20.5 for the anticipated restoration costs for the accumulation areas.

Table 20.5 – Annual Guarantee Payments Year Factor Financial Guarantee 70% 100% 1 0 $0 $961,650 2 0 $0 $480,825 3 0 $0 $480,825

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Year Factor Financial Guarantee 4 0.008 $10,770 $0 5 0.025 $33,658 $0 6 0.041 $55,199 $0 7 0.058 $78,086 $0 8 0.074 $99,627 $0 9 0.091 $122,514 $0 10 0.107 $144,055 $0 11 0.124 $166,942 $0 12 0.141 $189,830 $0 13 0.157 $211,371 $0 14 0.174 $234,258 $0 15 0 $0 $0 Since the restoration costs of the waste rock and tailings pile will be high, Nemaska Lithium will conduct research and development to reduce the costs of restoration of accumulation areas, such as: • Placement of the overburden; • Methods and species for the re-establishment of vegetation. 20.6 Conclusion

Consultations with stakeholders and a preliminary assessment indicate that the impacts identified to date do not present any critical risks to the Project at this time. It is anticipated that mitigation, remediation, compensation, surveillance and monitoring measures and programs, as well as other collaborative initiatives, should address the Project’s impacts in a manner that is commensurate with the standards set by authorities, the industry, and other projects, and which meet stakeholder expectations. Although the assessment of the Project’s social, environmental, and economic impacts is still ongoing, Nemaska Lithium and the Cree Nation of Nemaska are currently negotiating a Resource Development Partnership Agreement (“RDPA”), which sets out an arrangement for the sharing of Project benefits and for addressing Project impacts through a range of measures and programs.

20.7 Hydrometallurgical Plant

For the Valleyfield installation, the discussions with MDDEP indicate that this part of the Project did not require an ESIA.

However, a certificate of conformity to Valleyfield’s City regulations will have to be issued each time a certificate of authorization is requested under sections 22 and 32 of the

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Environment Quality Act (c. Q-2). Construction permits will also be required before starting any construction work.

At the provincial level, certificate of authorization will have to be delivered from the MDDEP in accordance with the Environment Quality Act (c. Q-2).

During the construction and the operation, many permits and leases will be required. Nemaska will present the requests when required.

From the federal point of view, if the Project does not fall under the Canadian Environmental Assessment Act, no specific permit will be required for the construction of the plant.

It is recommended that before the land acquisition a soil characterization and geotechnical survey be completed.

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21.0 CAPITAL AND OPERATING COSTS

The Project scope covered in this Study is based on the construction of a greenfield facility having a nominal daily processing capacity of 3,000 tonnes. The capital and operating cost estimates related to the mine, concentrator, and Whabouchi site infrastructure have been developed by BBA. Capital and operaring cost estimates related to the hydrometallurgical plant and associates infrastructures have been developed by Met-Chem. Engineering quantities related to the water and environmental management were provided by Journeaux Assoc., as well as the site closure plan, including the associated costs. Owner’s Costs were provided by Nemaska. BBA and Met-Chem consolidated cost information from all sources. Table 21.1 presents a summary of the total estimated initial and sustaining capital costs for the Project. The capital and the operating costs are in Canadian Dollars (“CAD”).

The capital cost estimate consists of the direct capital costs for the Whabouchi mine and concentrator site and for the Valleyfield plant site and the indirect costs for the two (2) sites. A contingency of 10% is added to Whabouchi site costs. A contingency of 15% is added to the Valleyfield site costs to allow for a PEA accuracy. The estimated Owner’s costs are also included. A Working Capital, equal to two (2) months Operating Costs, has been included as well. Table 21.1 shows the summary of the capital costs.

Table 21.1 – Summary of the Capital Costs Estimate Description Capital Costs CAD $M Whabouchi Site – Mine and Concentrator Total Direct Costs 110.9 Total Indirect Costs (incl. Owner’s Cost) 34.2 Contingencies 14.1 Sub Total Whabouchi Site 159.2 Valleyfield Site – Hydromet Plant Total Direct Costs 203.4 Total Indirect Costs 37.6 Contingencies 36.2 Sub Total Valleyfield Site 277.2 Mine Development Pre-Stripping 2.5 Trust Fund Rehabilitation First Payment 0.9 Working Capital 14.7 Total Capital Cost 454.5

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21.1 Whabouchi Capital Costs

Table 21.2 – Whabouchi Estimated Initial Capital Costs (M$) Total Estimated Project Work Breakdown Structure Initial Capital Costs (M$) 0000 – Off Site 0.9 1000 – Infrastructures 2.3 2000 – Administration and Services 3.8 3000 – Mine 10.8 4000 – Crushing 10.1 6000 – Processing Plant 79.0 7000 – Tailings and Water Management 3.8 8000 – Owner’s Costs Management and Project Development 4.7 Pre-Production Development 1.7 Mining Equipment 1.8 Electrical Power Line – Hydro-Québec 0.9 Accommodations and Travel Inc. Charter 2.5 Flights 9000 – Indirect Costs Engineering and Procurement Services incl. 9.4 Support to Constr. Team Construction Temporary Facilities and 9.1 Operation Pre-Op. – Mechanical Acceptance 4.3 Spares/Freight/Vendors Rep’s Contingency 14.1 Total 159.2 The total initial capital costs, including Indirect Costs and Contingency, have been estimated to be in the order of $159.2 M. Mining costs, including pre-stripping, estimated at $2.48 M. This estimate table does not include the following items, which were accounted for separately: • Leased equipment (mining equipment), estimated value at $7.96 M, which is included in the operating costs (interest of 6% over seven (7) years), with the exception of the initial deposit that was included in the Capital Cost Estimate;

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• The portion of rehabilitation and closure costs required to be disbursed prior to production start-up, estimated by Journeaux Assoc. in the order of $1,009 M; • Sustaining capital (capital expenses incurred in Year 1 of production to the end of mine life) estimated at $25.74M. 21.1.1 Basis of Estimate and Assumptions

The Capital Cost Estimate, pertaining to the processing areas and Whabouchi site infrastructure within the BBA scope, was performed by a professional estimator on BBA’s estimation team. Capital costs for the mine were estimated by BBA’s mining group. Material take-offs, fulfilled by Journeaux Assoc., were rated by BBA.

21.1.2 Dates, Currency, and Exchange Rates

This Capital Cost Estimate is expressed and presented in constant Q1 2012, Canadian Dollars.

Table 21.3 and Table 21.4 show the currency exchange factor used for the study and the splits for the direct costs of the Project.

Table 21.3 – Foreign Exchange Rates Country/Zone Currency Equivalent Europe EUR 1.30 CAD United States USD 1.00 CAD Table 21.4 – Direct Cost Currency Split ($ x 1,000) Country/Zone Currency Equivalent Europe 5,245.9 EUR 6,812.9 CAD United States 5,284.0 USD 5,284.0 CAD 21.1.3 Mining Quantities and Costs

Mining equipment quantities and costs have been developed by BBA’s mining group based on the mine plan developed in this Study. Equipment costs were estimated from BBA’s recently updated database of Vendor pricing. In order to reduce initial mining equipment costs, it is assumed that Nemaska Lithium will lease to own certain equipment that is required for pre-production and for Year 1 of operation.

Pre-stripping costs incurred in the pre-production period have been capitalized. This capital cost estimate is based on pre-stripping tonnage, as defined in the mine plan, and includes costs associated with the mining and hauling of overburden and waste rock.

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21.1.4 Scope and Quantity Development

This capital cost estimate is based on the construction of a greenfield facility having a capacity to process 3,000 tpd mill feed.

The design of the crusher area and the concentrator zone has been developed by BBA and has largely been based on BBA’s experience on other projects. General Arrangement drawings developed in this Study have been used to estimate quantities and generate MTOs for all commodities. Equipment costs have been estimated using Vendor prices based on budget proposals for major process equipment and some mechanical equipment packages. Labour rates have been estimated as described subsequently in this section. Related infrastructure has been estimated by BBA based on the site plan developed. A summary of the methodology associated with each engineering discipline is presented below: • Site Works – Earthwork quantities were measured from drawings and local topographical data; • Concrete – Preliminary design sketches were used to derive the foundations; • Structural Elements – Building sketches were developed to produce take-off; • Architectural – Siding and roofing quantities were taken from General Arrangement drawings; • Mechanical Process Equipment – A detailed mechanical equipment list was developed with capacities and sizing; • Mechanical Bulks – A plate work list was developed with sizing, weights and surface areas, including lining requirements; • Fire Protection and HVAC – MTOs were taken from layout and elevation drawings. An HVAC equipment list was also developed; • Piping – Diameter sizing carried out from preliminary design, while lengths determined from layouts. Lining requirements were also categorized; • Electrical Equipment – An equipment list was developed with capacities and sizing; • Electrical Bulks – MTOs derived from cable schedules and runs, including tray routing layouts; • Automation – Detailed instrumentation list was developed from marked-up PFs; • Telecommunication – Preliminary telecommunication design and equipment list was developed from plant site drawings and local data. Moreover, the mine garage, administration offices, and high performance fabric building (Sprung) were estimated using BBA’s recently updated internal database, as well as budget prices from suppliers.

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21.1.5 Pricing Sources

The pricing in this Estimate was based on a combination of budgetary quotes and/or historical data obtained from similar projects. • Concrete – Unit rates, including formworks and rebars, were built from similar ongoing projects overseen by BBA; • Steelwork – Labour productivity calculated from BBA database and material priced from actual steel market value; • Architectural – Pricing based on recent data from similar projects erected in equivalent region; • Plant Equipment – For all major mechanical equipment packages (23 in total), data sheets were prepared and budget pricing was obtained from Vendors. Whenever recent pricing was available for packages of lesser value, this information was used; • Mechanical Bulks – Pricing was based on recent project experience; • Piping - Material pricing for carbon steel and rubber lined piping was constructed from supplier propositions; • Electrical and Instrumentation Bulks: Pricing bulks are based on current published pricing; • Instrumentation equipment: Pricing was based on recent project experience; • Electrical Equipment: For all major electrical equipment packages, data sheets were prepared and Budget Pricing was obtained from Vendors, or was taken from recent projects. 21.1.6 Labour Rates and Labour Productivity Factors

For the purpose of defining the “Work Week”, all estimated costs for labour are based on ten (10) hours per day, through seven (7) days per week, for a total of 70 hours per week. Rotation is based on a work cycle of three (3) weeks on site and one (1) week off, unpaid. Work is forecasted to be realized on day-shift only.

The present estimate is structured and based on the philosophy that contracts will be awarded to reputable contractors on a lump-sum basis.

The hourly crew rates, used in the present estimate, have been derived from Quebec construction collective agreements dated January 1, 2012. Crew rates include a mix of skilled, semi-skilled and unskilled labour for each trade, as well as the fringe benefits on top of the gross wages. The supervision by the foremen and surveyors is built-in into the direct costs. The indirect costs consist of items such as small tools, consumables, supervision by the general foremen, management team, contractors on site temporary

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construction facilities, mobilization / demobilization, contractor’s overhead and profit. They also include the costs related to the transportation of the employees to and from their residence to construction site, as well as room and board (if required) during traveling. The construction equipment rates are based on the tariff proposed by the Direction générale des acquisitions du Centre de services partagés du Québec, detailed inside edition dated April 1, 2011; and by Ontario Provincial Standard Specifications (“OPSS”) schedule of rental rates for construction equipment. The cost of fuel (diesel), associated to the construction equipment in the present estimate, is estimated at 1.10 $/litre given that we are taking under consideration a potential return on taxes.

Table 21.5 – Labor Rates Used for Capital Cost Estimation Crew Rates Based on 70 hours/work week ($) Labour Rate Construction Typical Crew Total Direct Indirect Equipment Site Works – Civil 70.70 37.50 53.40 161.60 Concrete works 71.90 41.00 11.30 124.20 Metal Works 77.20 44.10 25.40 146.70 Architectural Finishes 72.00 41.10 7.30 120.40 Mechanical – Process 72.30 43.00 27.10 142.40 Mechanical – Building 73.20 43.30 19.90 136.40 Piping 71.30 42.60 19.20 133.10 Insulation 69.60 37.40 7.40 114.40 Electrical 72.60 43.10 4.80 120.50 Automation/Telecom. 72.30 43.00 1.70 117.00 Construction Project Performance is an important concern for Project owners, constructors, and cost management professionals. Project costs and schedule performance depend largely on the quality of Project planning, work area readiness, preparation, and the resulting productivity of the work process made possible in Project execution. Labour productivity is often the greatest risk factor and source of cost and schedule uncertainty to owners and contractors alike.

Many terms are used to describe productivity in the construction industry: performance factor, production rate, unit person-hour rate and others. Traditionally, productivity has been defined as the ratio of input / output, i.e., the ratio of the input of an associated resource (usually expressed in person-hour) to real output (in creating economic value). To restate this definition for use in the construction industry: labour productivity is the physical progress achieved per person-hour, i.e., person-hours per linear metre of conduit laid or person-hours per cubic metre of concrete poured. The two (2) most important measures of labour productivity are:

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• The effectiveness with which labour is used in the construction process; • The relative efficiency of labour, doing what it is required, at a given time and place. A study, carried out by Optima Engineers and Constructors, for the Alberta Economic Development, revealed 208 irregularities affecting the labour productivity on a construction site. BBA summarized the results of this study into the following major factors: • Site location • Weather conditions • Extended overtime • Scattered items of work • Access to work area • Overcrowded / Tight work areas • Height – Scaffolding • Complexity • Availability of skilled workers • Efficiency • Labour turnover • Supervision • Inspection + QA / QC • Revamps / Connections / Tie-ins • Sophisticated specifications • Fast-track requirements • Materials + Equipment + Handling • Safety / Security Table 21.6 presents the labour productivity factors applied in the capital cost estimate for Nemaska Lithium Project.

Table 21.6 – Factors Affecting Labour Productivity Productivity Loss Ratio Activity Factor Site Works – Civil 1.356 Concrete works 1.430 Metal Works 1.543 Architectural Finishes 1.482 Mechanical Works 1.564 Piping/Insulation 1.614 Electrical 1.582 Automation/Telecom. 1.570 21.1.7 Indirect Costs

Indirect Costs for areas under BBA’s responsibility for capital cost estimation were estimated by BBA based on experience on other projects as described below. It is important to note that these factors were estimated based on the assumption that Nemaska

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Lithium will execute the Project using an internal Procurement and Construction Management (“PCM”) team rather than contracting a major PCM firm. • Costs related to the construction of temporary facilities required during the Project construction period comprise of costs incurred for building and maintaining temporary facilities and accesses, which will no longer be required once construction is completed. An itemized list with budget allowance was developed by BBA. • The operations of temporary construction facilities are capitalized. An itemized list with budget allowance was developed by BBA. • Engineering and Procurement. A detailed staffing plan was the basis of the services estimate. • Construction Management (“CM”). The construction management will be performed by Nemaska Lithium and those costs are included in the Owners’ Costs. An allowance for engineering field support services from BBA has been included. • An allowance of 3,000 man-hours for pre-operational acceptance of mechanical equipment was included. • Costs for plant mobile equipment used during construction was based on an itemized list with budget amounts. • Costs for spare parts and initial fill were estimated at 2.5% of equipment value. • The camp facilities, operation & the catering will be provided by CCDC at the location of the Relais Routier Nemiscau Camp on the Route du Nord, which is located about 12 km to the west of the mine site. The requirement for the capacity of the camp has been estimated at 215 based on the Project schedule. Budget pricing was provided by CCDC and has been used in this Study. • Freight: Domestic-sourced materials are FOB job site, and therefore, costs are included with the material pricing. Freight for design/supply packages has been estimated at 6.75% of the equipment costs. • The cost for Vendor representatives for equipment installation and pre-operational support is based primarily on estimated durations and historical daily pricing. 21.1.8 Owner’s Costs

Owner’s costs include items such as owner’s team salaries and expenses, insurance, authorization certificates and permits, geotechnical and surveying costs, laboratory test work, etc. An itemized list with budget allowance was developed by Nemaska Lithium.

An allowance was also included for the power line to be built by Hydro-Québec.

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21.1.9 Contingency

Contingency provides an allowance in the capital cost estimate for undeveloped details within the scope of work covered by the estimate. Contingency is not intended to take into account items such as labour disruptions, weather related impediments, changes in the scope of the Project from what is defined in the Study, nor does contingency take into account price escalation or currency fluctuations. A contingency of 10% of the sum of direct and indirect costs has been attributed to the capital cost estimate developed in this Study for areas estimated by BBA.

21.1.10 Exclusions

The following items are not included in this capital cost estimate: • Inflation and escalation, the estimate is in constant Q1 2012 Canadian Dollars; • Currency fluctuations; • Costs associated with protection against currency fluctuations; • All taxes, duties and levies; • Project financing costs; • Provision for outdoor winter concrete or formwork in winter season; • Disposal of hazardous materials. 21.1.11 Assumptions • Hydro-Québec will provide the permanent power line in time for use as construction power; • It was assumed that all backfill materials will be available from gravel pits, esker or other sources located within a radius of five (5) km; • Mine waste rock is suitable for use in road construction (as established following preliminary assessment); • Mass earthworks and haulage road construction is performed by crews from Nemaska; • Soil conditions will not require special foundation designs such as grouting and piling (as established following preliminary assessment); • All excavated material will be disposed of within the site battery limits; • The external envelope of the mineral processing plant is based on a pre-engineering building design; • Estimate is based on the Project obtaining all relevant permits in a timely manner to meet the Project schedule.

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21.2 Hydrometallurgical Plant Capital Costs

21.2.1 Capital Cost Estimate and Battery Limits

This part of the Project scope is based on the construction of a greenfield facility having a nominal yearly processing rate of 213,558 tonnes to produce lithium hydroxide monohydrate and lithium carbonate. The estimate is based on Met-Chem’s standard methods applicable for a preliminary economic assessment study to achieve the accuracy level of ± 25%.

All duties and taxes are excluded from the capital cost, but are considered in the economic analysis. Escalation and interests incurred during construction are also excluded from the Capital Cost.

The effective date for the cost estimate is Q3 2012. The estimate is expressed in CAD dollars.

The pre-production initial capital cost for the base case scope of work is CAD$ 277.2 M, of which CAD$ 203.4 M is direct cost, CAD$ 37.6 M is indirect cost and CAD$ 36.2 M is contingency.

No provision is required for sustaining capital. Capital costs are summarized in Table 21.7.

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Table 21.7 – Summary of Capital Cost Estimate Total Total Total Initial Sustaining Item Description Capital Capital Capital CAD$ M CAD$ M CAD$ M 0 Total Direct Cost 203.4 0 203.4 20 Hydrometallurgical Process 177.8 177.8 30 Tailings and Water Management Facilities 1.8 1.8 40 General Services 2.1 2.1 50 Infrastructures 3.7 3.7 60 Power and Communication 9.3 9.3 70 Service Vehicle and Equipment 0.8 0.8 (Allowance) 80 Contractor’s Costs 7.9 7.9

I Indirect Cost (including financial costs) 37.6 0 37.6 C Contingency 36.2 0 36.2 T Total Capital Cost 277.2 0 277.2 21.2.2 Basis of Estimate for Direct Capital Cost a) Currencies Updated indices were used for quotation received before Q3 of 2012. No provisions for escalation or currency fluctuation are included. The exchange rates used when quotations were received in foreign currencies are 1.00 CAD / 1.00 USD and 1.24 CAD / 1.00 EUR. b) Construction Labour

The labour rate was developed for a typical crew from a detailed table of current rates developed by the Corporation des Entrepreneurs Généraux du Québec. The all-inclusive hourly rate is CAD$ 100, and includes the basic hourly rates for the tradesman, social benefits and employer’s burden, industrial site premium, direct supervision, small tools and consumables, and contractor’s overhead and profit.

Indirect supervision is excluded from the hourly rate but is included in the construction contractor’s site management provision.

In addition to the labour cost, a construction allowance based on delivered equipment cost was established from similar projects to cover for construction material, sub-contract and mobile cranes to be paid by the owner. Few heavy or

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complex crane lifts as well as limited specialized worker trades are expected. Therefore, a factor of 3.5% is applied.

The working calendar was defined as one (1) shift per day, eight (8) hours per day and five (5) days per week for a total of 40 hours per week. Realistic construction schedule, good site conditions as well as limited number of contractors on site, limited work outside in winter and also limited work in overtime is expected. Therefore, a factor of 1.07 is applied for productivity loss.

The estimate is based on construction contracts attributed on the base of competitive bidding process amongst qualified contractors. Availability of local qualified contractors and skilled workers is expected. It is also expected that an average level of site management, contract administration, quality control and adequate safety requirements will be required from the contractors by the construction management. c) Freight, Duties and Taxes Based on recent surveys and studies and when not included in the cost, the freight was accounted for by adding a factor of 10% to the value of the goods. All other duties and taxes are excluded from the capital cost, but are considered for the economic analysis. d) Contractor’s Costs Provisions have been included to cover for contractor’s major equipment and supplies, including owned and rented construction equipment, vehicles and other facilities such as trailers, tool cribs, power panels, containers, maintenance of area, janitorial and clean-up, and also the mobilisation and demobilisation. Special installation tools, cranes, scaffolding, cribbing and dunnage are also included as well as work place weather protection. Workers transportation within the construction site is also included.

Provision have also been made for construction contractor’s site management including supervision and support staff such as administration and procurement, coordination and scheduling, and quality and safety. e) Process The process facilities at the Hydrometallurgical Plant include the main process building consisting of the process area, the mechanical shop, the office, change room and lunch room and the laboratory areas as well as railcar unloading facility.

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i) Main Building

The cost for the building is estimated based on the type and dimensions of the building and unit cost from recent similar projects, adjusted for the location of the Project.

ii) Process Equipment

The process equipment list was derived from the flow sheets. For major equipment, budget prices were obtained from qualified suppliers based on data tables or preliminary technical description. The remaining equipment was estimated from databases from recent similar projects or in house estimation.

Installation was estimated for each area by factorisation on delivered process equipment.

iii) Piping and Pipelines

Process piping cost was established for each area by factorisation on delivered process equipment. Small bore pipelines and pipe racks for tailings and gypsum by-product was estimated based on quantities from layout and unit rates from databases.

iv) Electrical

Electrical equipment list and quantities were derived from the single line diagram. Budget prices were established based on databases from recent projects. No provision was made for redundancy on the power supply installations. Quantities and costs for material as well as man-hours were also established based on recent similar projects. Installation was estimated using hourly rate as described above.

v) Instrumentation

Instrumentation, automation and communication material and equipment quantities were derived from the flow sheets. Budget prices were established based on databases from recent projects. Installation was estimated using hourly rate as described above.

vi) Buildings Services and Supplies

For each process area, buildings services and supplies were estimated by applying a factor to delivered equipment cost, based on recent similar

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projects. Consideration was given that some services will be provided by the City of Valleyfield. f) Tailings and By-products Management Facilities The management and storage facilities were designed for one (1) year only of storage in the gypsum by-product pond, one (1) year only of storage in the aluminium silicate by-product stockpile and 20 years of storage in the SIR tailings pond. It is expected that markets will be found within one (1) year to then dispose of gypsum and aluminium silicate by-products. Costs were based on material take- offs from the layouts and unit cost from databases. g) General Services, Infrastructure and Ancillary Buildings General services such as natural gas and general fire protection were estimated based on recent similar projects. Services such as sanitary waste disposal and potable water were not included as they will be supplied by the City of Valleyfield. However, a provision is made for interconnection to public networks based on a budget price provided by the City of Valleyfield.

Site preparation, peripheral fencing and site roads cost estimate were established based on allowance, quantities derived from general layouts and budget unit price based on recent similar projects. h) Ancillary Buildings and Facilities Preliminary requirements were established from the layouts and also process data for the storage warehouse, the shipping warehouse, the railcar unloading system cover structure and the guard house.

Unit rates and allowances based on recent similar projects were applied to estimate the costs of the storage warehouse and the guardhouse, while budget prices were obtained from a qualified supplier for prefabricated structures for the shipping warehouse and the railcar unloading system cover structure.

Facilities such as offices, change rooms, laboratory, as well as mechanical and electrical rooms, are located within the main process building and included in its cost.

Finally, because the Project is centrally located, there are no provisions for permanent or construction camps.

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i) Site Access Quantities were calculated from the layout for the construction of 850 m railway connection to the existing CN line and some turnouts. Unit rates obtained from a qualified contractor were applied to estimate the cost. j) Power Supply, Main Substation and Communication Preliminary requirements were established for a main power line. Estimation of the cost was based on quantities and prices for material as well as installation man- hours, based on recent similar projects.

Preliminary requirements were established for the main substation based on the power demand. Equipment budget prices and costs for material and installation were established based on databases from recent similar projects.

Preliminary requirements were established for the site distribution power lines. Estimation of the cost was based on quantities and prices for material as well as installation man-hours, based on recent similar projects.

Preliminary requirements were established for emergency power supply. A budget price was estimated based on recent similar projects. k) Service Vehicles and Equipment An allowance was made for the service vehicles based on preliminary requirements and recent similar projects.

21.2.3 Basis of estimate for Indirect Costs

Preliminary estimation for indirect costs is included in Table 21.7. This provision was established by factorisation on the total direct cost and typically covers for the following major items detailed here under: Project Development, Project Implementation, Financial Costs and Closure Costs. a) Project Development Project Development mainly includes: Permitting Process, Land Acquisition, Administration, NSR buyout, Engineering Studies as Feasibility studies and Independent Review, Metallurgical Testing, Geotechnical and Occupational Hazard studies, Social Impact Studies and Community Relations, Preproduction Operation Group and Legal Fees. No provision is made for development, implementation and operation of a demonstration plant.

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b) Project Implementation Project Implementation mainly includes: EPCM, Spares, First Fills, Commissioning and Owner’s Costs.

“EPCM” includes Detailed Engineering, Procurement, Construction Management and Commissioning Assistance and Site Assistance.

Spares and First Fills are excluded from the initial capital cost but are considered as working capital. Commissioning include commissioning spare parts, dry and wet commissioning as well as vendors representative on site.

Construction owner’s costs include mainly construction indirect costs and Owner’s Project team. c) Financial Costs Financial Costs includes mainly insurances, escalation and interests incurred during construction. Escalation and interests are excluded from the Capital Cost. Working capital, taxes and duties are also excluded from the Capital Cost estimate but are considered in the financial analysis. d) Closure and Rehabilitation Costs Preliminary indication is that no provision is required for closure and rehabilitation of the site.

21.2.4 Contingency

Based on the level of development stage of the Project as well as assessment of major risks, a factor was applied to estimate the provision for contingency. No contingency provision was made to cover technological risks associated with the process. It is expected that sufficiently developed engineering, adequate project management and tight cost control will be implemented in order to meet the budget for the Project.

21.2.5 Sustaining Capital Expenditures

There are no sustaining capital expenditures for the Hydrometallurgical Plant.

21.3 Operating Costs

The operating costs have been estimated for the Whabouchi operations that include the mine and concentrator operations and the tailings and waste disposal. They are estimated based on the average over the life of the Project. The mine operating costs include the costs for leasing the major equipment. The truck transport of the concentrate from the site to Chibougamau will be handled by a contractor and the transport to Valleyfield will be

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by train. The operating costs at the Hydrometallurgical Plant have also been estimated. A summary of the costs is shown in Table 21.8.

Table 21.8 – Summary of the Operating Costs Estimate Costs $/t $/t $/t LiOH- Description Mineralized $/t Li CO Concentrate H O 2 3 Material 2 Mining 15.35 78.74 Processing (Concentrator) 15.49 79.44 Transport of Tailings 0.91 4.68 G&A Whabouchi 6.98 35.81 Concentrate Transport 11.70 60.00 Hydro-Metallurgical Process 44.22 229.18 Total Cost 94.65 485.54 3,400 3,495 Royalties 0.32 1.67 21.3.1 Mining Operating Costs

Table 21.9 presents a summary of total estimated average operating costs for producing concentrate in Nemaska. Those costs are presented in Canadian Dollars (CAD$) per tonne of concentrate produced. Operating costs were estimated based on the average over the life of the mine. Operating costs include the estimated cost of leased equipment over the life of the lease.

Table 21.9 – Total Estimated Average Operating Costs for Production of Concentrate Project Operating Costs $/t Milled $/t Concentrate Mining Operating Costs 15.87 81.37 Haulage of Tailings and Middlings 0.91 4.68 Process Plant Operating Costs 15.49 79.44 G&A Operating Costs 6.91 35.45 Total F.O.B. Mine 39.19 200.94 The total estimated operating costs are in the order of $200.94/t of concentrate produced. These costs include the cost of leasing mining equipment for the life of the lease. The operating costs developed were based on power costs that are in line with local rates and fuel costs that were obtained from a major fuel supplier in the region.

Royalties are not included in the Operating Cost Estimate presented but are treated separately in the economic analysis.

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a) Mine Equipment Maintenance Costs The equipment operating costs are derived from supplier quotations (2011-2012), historical numbers and values from BBA’s internal database. The equipment operating costs are split between fuel costs and unit operating costs, which usually include maintenance, parts, repairs, etc.

The maximum $/hr unit operating costs for the major equipment (excluding fuel and labour) are as follows: • $52.88/hr for the CAT390 hydraulic shovel; • $40.13/hr for the CAT772 haul truck; • $49.83/hr for the CAT988H loader; • $105.87/hr for the DTH drill. The yearly mine maintenance operating costs are summarized for all primary, secondary and auxiliary equipment in Table 21.10. The equipment operating costs (excluding fuel) total $55.85 M over the LOM. b) Mine Fuel Consumption Costs Mine fuel consumption was based on supplier quotations from 2011, as well as historical trends from an internal database. The fuel consumption costs are calculated using a local fuel cost on site of $1.1046/L.

The yearly fuel operating consumption is summarized for all primary, secondary and auxiliary equipment in Table 21.11. As well, the total operating fuel costs for each year are summarized in one line at the bottom of the table. The unit fuel consumptions are provided in the first column of the table in litres per hour. The total consumption calculations are dependent on the Gross Operating Hours (GOH) for the specific type of equipment, on the total shifts required in all materials, and on the utility factor.

The total equipment fuel costs over the LOM total $67.62 M.

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Table 21.10 – Annual Mine Equipment Operating Costs (Excluding Fuel Costs) (M$)

PP Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19 Total

Primary Equipment Hydraulic Excavator (CAT390D- 6m³) 0.07 0.23 0.31 0.33 0.35 0.38 0.41 0.48 0.55 0.55 0.64 0.55 0.52 0.50 0.45 0.34 0.26 0.18 0.16 0.01 7.28 Wheel Loader (CAT988-H) 0.05 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.03 3.34 Haul Truck (CAT772-51 ton-46 tonnes) 0.12 0.44 0.74 0.89 0.91 1.10 1.12 1.34 1.65 1.64 1.66 1.64 1.64 1.64 1.59 1.30 1.01 0.73 0.64 0.04 21.79 Drill DTH (Cubex QXR 920 - 6") 0.01 0.07 0.14 0.19 0.32 0.35 0.28 0.34 0.41 0.42 0.50 0.43 0.41 0.39 0.35 0.27 0.21 0.15 0.13 0.01 5.34 Total Primary Equipment 0.24 0.93 1.37 1.59 1.76 2.02 1.99 2.34 2.80 2.79 2.98 2.80 2.75 2.71 2.57 2.10 1.67 1.25 1.11 0.09 37.76 Secondary Equipment Wheel Dozer 0.02 0.07 0.07 0.07 0.07 0.07 0.07 0.07 0.07 0.07 0.07 0.07 0.07 0.07 0.07 0.00 0.00 0.00 0.00 0.00 0.95 Track Dozer (D7) 0.05 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.00 0.00 0.00 2.98 Motor Grader (Caterpillar 14M) 0.05 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.19 0.03 3.50 Water Truck 0.02 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.01 1.61 Total Secondary Equipment 0.13 0.53 0.53 0.53 0.53 0.53 0.53 0.53 0.53 0.53 0.53 0.53 0.53 0.53 0.53 0.46 0.46 0.28 0.28 0.05 9.04 Auxiliary Equipment Fuel / Lube Truck 0.03 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.13 0.02 2.35 Service Truck ( 250 HP 22,000 GVW) 0.01 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.01 0.89 Tire Changer (attachment for 988-H) 0.01 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.00 0.00 0.00 0.82 Pick Up Truck (4x4 crew cab 0.02 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.07 0.07 0.04 0.04 0.01 2.31 Chevrolet 2500) Pick Up Truck (4x4 single cab 0.02 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.07 0.07 0.04 0.04 0.01 2.31 Chevrolet 2500) Light Plant (1000 W diesel generator) 0.01 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.02 0.00 0.00 0.37 Total Auxiliary Equipment 0.10 0.55 0.55 0.55 0.55 0.55 0.55 0.55 0.55 0.55 0.55 0.55 0.55 0.55 0.55 0.40 0.40 0.27 0.25 0.04 9.05 Total OPEX - Maintenance 0.47 2.00 2.45 2.66 2.83 3.09 3.06 3.42 3.87 3.87 4.05 3.87 3.83 3.79 3.65 2.96 2.53 1.80 1.64 0.18 55.85

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Table 21.11 – Annual Mine Equipment Operating Costs (Fuel Costs) (M$)

l/hr PP Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19 Totals

Haul Truck (CAT772-51 ton-46 tonnes) 57 0.16 0.62 1.05 1.26 1.29 1.57 1.59 1.90 2.35 2.33 2.35 2.32 2.33 2.34 2.25 1.85 1.43 1.04 0.90 0.06 30.95 Hydraulic Excavator (CAT390D- 6 m³) 64 0.09 0.28 0.38 0.40 0.43 0.47 0.49 0.58 0.67 0.66 0.78 0.67 0.63 0.60 0.54 0.41 0.32 0.22 0.20 0.01 8.82 Drill DTH (Cubex QXR 920 - 6") 65 0.01 0.11 0.16 0.17 0.19 0.21 0.20 0.25 0.30 0.30 0.36 0.31 0.30 0.28 0.26 0.20 0.15 0.11 0.10 0.01 3.96 Wheel Loader (CAT988-H) 50 0.05 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.03 3.35 Wheel Dozer 25 0.03 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.00 0.00 0.00 0.00 0.00 1.57 Track Dozer (D7) 40 0.04 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.00 0.00 0.00 2.86 Motor Grader (Caterpillar 14M) 33 0.04 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.02 2.65 Water Truck 20 0.02 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.01 1.61 Fuel / Lube Truck 25 0.02 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.02 1.68 Service Truck ( 250 HP 22,000 GVW) 10 0.005 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.003 0.335 Tire Changer (attachment for 988-H) 10 0.005 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.018 0.000 0.000 0.000 0.298 Pick Up Truck (4x4 crew cab Chevrolet 6 0.01 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.04 0.04 0.02 0.02 0.00 1.38 2500) Pick Up Truck (4x4 single cab Chevrolet 6 0.01 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.04 0.04 0.02 0.02 0.00 1.38 2500) Light Plant (1,000 W diesel generator) 4 0.002 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.007 0.000 0.000 0.127 Mobile Pump (125 HP diesel) 4 0.002 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.015 0.007 0.007 0.001 0.252 Total Fuel Consumption 0.50 2.05 2.62 2.86 2.94 3.27 3.31 3.76 4.34 4.33 4.52 4.33 4.29 4.25 4.08 3.29 2.74 1.96 1.77 0.18 61.22 Total Fuel Cost 0.55 2.26 2.89 3.16 3.24 3.62 3.66 4.15 4.80 4.79 4.99 4.79 4.74 4.69 4.51 3.64 3.02 2.16 1.96 0.19 67.62

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c) Blasting Costs Blasting and loading services will be carried out under a blasting contract. Costs under this proposed contract have been validated with local suppliers. The emulsion and accessories costs are outlined below Table 21.12.

Table 21.12 – Blasting Costs by Category Category Blasting Costs Fuel Price ($/L) 1.1046 Emulsion ($/100 kg) 103 Unitronic Double Prime ($) 25.25 Pentex D16 Booster ($) 8.9 Surface Connector ($) 0.908 The powder factors for mineralized material and for waste are 0.30 and 0.26 kg/t, respectively. The total unit cost for blasting mineralized material and waste are 0.34 and 0.32 $/tonne (of blasted material), respectively. The powder factor is calculated by taking the total bulk emulsion mass per hole (i.e., kg/hole) and dividing by the expected rock mass per hole. The total unit cost is calculated by taking the sum of the total required emulsion cost, accessories cost and fuel cost over the life of the mine, and then dividing by the total mine plan tonnage.

Other mine operating costs related to blasting are limited to the costs associated for the blasting contractors. The following salaries are allocated to the blasting contractors on a monthly basis, over the life of the mine: • Site supervisor: $14,943/month; • MMU Operator: $10,775/month; • Mechanic (blast): $12,425/month. d) Mine Personnel The mine personnel salaries were obtained from the 2011 Infomine Canadian Mine Salaries, Wages and Benefits Survey. For both hourly and salaried personnel, middle ranges were chosen from a summary of surface mined in Eastern Canada.

The total operating costs allocated to hourly personnel are approximately $93.53 M, whereas the total operating costs for salaried personnel reach a total of $44.49 M. This accounts for a total of $141.02 M in personnel operating costs over the life of the mine. A complete list of personnel salaries can be seen in the Table 21.13.

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Table 21.13 – Complete Personnel Yearly Salaries (M$) Position PP Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19 Totals Shovel / Loader Operator 0.10 0.38 0.38 0.38 0.38 0.38 0.38 0.48 0.58 0.58 0.67 0.58 0.58 0.58 0.38 0.38 0.38 0.19 0.19 0.02 7.96 Haul Truck Operator 0.16 0.66 0.82 0.99 0.99 1.32 1.32 1.56 1.89 1.89 1.89 1.89 1.89 1.89 1.81 1.48 1.15 0.99 0.82 0.08 25.43 Drill Operator 0.02 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.09 0.02 3.21 Wheel Dozer Operator 0.08 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.00 0.00 0.00 0.00 0.00 4.82 Track Dozer Operator 0.08 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.00 0.00 0.00 5.49 Grader Operator 0.08 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.34 0.03 6.17 Water Truck Operator/ Snow Plow/ Sanding 0.04 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.01 2.85 Other Auxilliary Equipment 0.04 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.01 3.08 General Labour 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.00 0.00 0.00 1.25 Janitor 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.00 0.00 0.00 0.00 0.00 1.09 Hourly Open Pit Operations Total 0.62 2.72 2.88 3.05 3.05 3.38 3.38 3.72 4.15 4.15 4.24 4.15 4.15 4.15 3.87 3.13 2.80 2.02 1.77 0.17 61.35 Field Gen Mechanics 0.02 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.09 0.03 3.21 Field Welder 0.03 0.11 0.11 0.11 0.11 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.11 0.11 0.02 3.30 Field Electrician 0.03 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.11 0.02 3.84 Shovel Mechanics 0.03 0.22 0.22 0.22 0.22 0.44 0.44 0.44 0.44 0.44 0.44 0.44 0.44 0.44 0.44 0.22 0.22 0.11 0.11 0.02 5.91 Shop Electrician /Millright 0.05 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.11 0.02 3.87 Shop Mechanic 0.05 0.22 0.22 0.22 0.22 0.22 0.22 0.33 0.44 0.44 0.44 0.44 0.44 0.44 0.44 0.22 0.22 0.11 0.11 0.02 5.39 Mechanic Helper 0.02 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.00 0.00 1.46 Welder-machinist 0.03 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.11 0.11 0.00 0.00 0.00 2.53 Lube/Service Truck 0.02 0.08 0.08 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.17 0.08 0.08 0.08 0.08 0.08 0.01 2.47 Electronics Technician 0.03 0.11 0.11 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.11 0.11 0.11 0.00 0.00 3.19 Hourly Mine Maintenance Total 0.31 1.55 1.55 1.74 1.74 2.07 2.07 2.18 2.40 2.40 2.40 2.40 2.40 2.40 2.31 1.66 1.66 1.22 0.72 0.14 35.17 Mine Superintendant 0.04 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.03 2.75 Mine Shift Foreman 0.03 0.56 0.56 0.56 0.56 0.56 0.56 0.56 0.56 0.56 0.56 0.56 0.56 0.56 0.56 0.23 0.23 0.11 0.00 0.00 8.50 Drill & Blast Foreman 0.00 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.00 0.00 0.00 1.81 Blaster 0.02 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.00 0.00 0.00 Dispatcher 0.00 0.09 0.09 0.09 0.09 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.09 0.09 0.00 0.00 0.00 2.29 Salaried Open Pit Operations Total 0.09 1.01 1.01 1.01 1.01 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 0.67 0.67 0.36 0.15 0.03 16.99 Maintenance Superintendant 0.04 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.14 0.00 0.00 0.00 2.35 Maintenance Foreman 0.03 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.22 0.11 0.11 0.11 0.11 0.02 3.47 Maintenance Clerk/ Admin Assistant 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.06 0.06 0.06 0.00 0.00 1.82 Salaried Mine Maintenance Total 0.06 0.48 0.48 0.48 0.48 0.48 0.48 0.48 0.48 0.48 0.48 0.48 0.48 0.48 0.48 0.31 0.31 0.17 0.11 0.02 7.65 Chief Engineer 0.04 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.00 0.00 2.53 Senior Mine Planning Engineer 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.12 0.00 0.00 2.12

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Position PP Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19 Totals Open Pit Engineer 0.03 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.10 0.02 1.90 Mining Engineering technician 0.02 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.09 0.09 0.09 0.09 0.01 2.85 Mine Surveyor 0.02 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.18 0.09 0.09 0.09 0.00 0.00 2.76 Salaried Mine Engineering Total 0.11 0.73 0.73 0.73 0.73 0.73 0.73 0.73 0.73 0.73 0.73 0.73 0.73 0.73 0.73 0.55 0.55 0.55 0.19 0.03 12.16 Chief Geologist 0.04 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.15 0.00 0.00 2.58 Grade Control Geologist 0.03 0.21 0.21 0.21 0.21 0.21 0.21 0.21 0.21 0.21 0.21 0.21 0.21 0.21 0.21 0.10 0.10 0.10 0.10 0.02 3.33 Sampler 0.01 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.11 0.06 0.06 0.06 0.00 0.00 1.79 Salaried Geology Total 0.08 0.47 0.47 0.47 0.47 0.47 0.47 0.47 0.47 0.47 0.47 0.47 0.47 0.47 0.47 0.31 0.31 0.31 0.10 0.02 7.70 Total Personnel 1.27 6.96 7.12 7.48 7.48 8.22 8.22 8.68 9.32 9.32 9.42 9.32 9.32 9.32 8.96 6.63 6.30 4.63 3.04 0.40 141.02

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e) Mine Operating Cost Summary The operating costs are summarized by category on a unit operating cost basis, and on a yearly total basis alongside the mine schedule. Table 21.14 demonstrates the division of mine unit operating costs; Table 21.15 demonstrates the complete operating cost summary over the life time.

Table 21.14 – Summary of Mining Unit Operating Costs by Category $/t mined $/t milled Personnel 1.81 7.19 Salaried 0.57 2.27 Hourly 1.24 4.92 Equipment 1.59 6.30 Maintenance+Parts 0.72 2.85 Fuel 0.87 3.45 Blasting 0.37 1.45 Others (Blasting Contractors) 0.11 0.43 Total (Cash Cost Unit OPEX) 3.88 15.37 Leasing 0.13 0.50 Total (Cash Cost +Leasing) Unit OPEX 4.00 15.87

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Table 21.15 – Summary of Mine Operating Costs (M$)

Open Pit Production Units PP Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19 Total

Mill Feed Tonnes Mt 0.05 0.84 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 0.13 19.64

Li2O % (Grade) % 1.460 1.528 1.514 1.542 1.454 1.426 1.483 1.542 1.576 1.533 1.437 1.452 1.454 1.468 1.444 1.483 1.495 1.530 1.529 1.528 1.494 Waste rock Mt 0.11 1.25 1.90 2.09 2.58 2.95 2.70 3.63 4.62 4.74 5.88 4.89 4.58 4.30 3.79 2.62 1.77 0.88 0.67 0.04 55.95 OB Mt 0.43 0.26 0.21 0.23 0.02 0.04 0.45 0.39 0.23 0.11 2.36 Waste + OB Tonnes Mined Mt 0.54 1.50 2.12 2.32 2.60 2.98 3.15 4.02 4.85 4.85 5.88 4.89 4.58 4.30 3.79 2.62 1.77 0.88 0.67 0.04 58.31 Stripping Ratio 10.88 1.79 1.93 2.12 2.38 2.72 2.88 3.67 4.43 4.43 5.37 4.46 4.18 3.93 3.46 2.40 1.61 0.80 0.61 0.48 2.97 Total Tonnes Mined Mt 0.64 2.34 3.21 3.41 3.70 4.08 4.25 5.11 5.95 5.94 6.97 5.98 5.67 5.39 4.89 3.72 2.86 1.98 1.76 0.12 77.87 Equipment Maintenance + ALL M$ 0.47 2.00 2.45 2.66 2.83 3.09 3.06 3.42 3.87 3.87 4.05 3.87 3.83 3.79 3.65 2.96 2.53 1.80 1.64 0.11 55.96 Equipment Fuel M$ 0.55 2.26 2.89 3.16 3.24 3.62 3.66 4.15 4.80 4.79 4.99 4.79 4.74 4.69 4.51 3.64 3.02 2.16 1.96 0.13 67.75 Blasting M$ 0.07 0.89 1.20 1.26 1.42 1.53 1.45 1.75 2.07 2.11 2.47 2.15 2.05 1.97 1.80 1.43 1.16 0.87 0.81 0.05 28.52 Personnel M$ 1.27 6.96 7.12 7.48 7.48 8.22 8.22 8.68 9.32 9.32 9.42 9.32 9.32 9.32 8.96 6.63 6.30 4.63 3.04 0.20 141.22 Other M$ 0.11 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.46 0.00 8.36 Re-handling 0.50 0.03 0.03 Total OPEX (Cast Cost) M$ 2.48 12.57 14.12 15.02 15.43 16.92 16.86 18.46 20.51 20.54 21.39 20.59 20.40 20.22 19.38 15.11 13.47 9.93 7.91 0.53 301.83 CASH COST UNIT OPEX Cost Per Tonne Mined $/t 3.86 5.36 4.39 4.40 4.17 4.15 3.97 3.61 3.45 3.46 3.07 3.44 3.60 3.75 3.96 4.06 4.70 5.02 4.49 4.49 3.88 Cost Per Tonne Milled $/t 14.97 12.90 13.72 14.09 15.46 15.40 16.85 18.73 18.76 19.53 18.80 18.63 18.47 17.69 13.80 12.30 9.07 7.22 4.04 15.37 Leasing Cost $ 0.91 1.43 1.43 1.43 1.43 1.43 1.27 0.51 9.83 CASH COST + LEASING UNIT OPEX Cost Per Tonne Mined $/t 3.86 5.75 4.84 4.82 4.56 4.50 4.30 3.86 3.54 3.46 3.07 3.44 3.60 3.75 3.96 4.06 4.70 5.02 4.49 9.64 4.00 Cost per Tonned Milled $/t - 16.06 14.20 15.02 15.39 16.76 16.70 18.02 19.20 18.76 19.53 18.80 18.63 18.47 17.69 13.80 12.30 9.07 7.22 8.68 15.87

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21.3.2 Haulage of Coarse Tailings

The estimated cost for the haulage of the coarse tailings is summarized in Table 21.16.

Table 21.16 - Estimated Operating Costs of Coarse Tailings Material Operating Costs $/t milled Handling of Coarse Tailings 0.91 The coarse tailings operating costs are separated from the mining costs, since they are not directly related to the LOM plan, and do not require the same haulage schedule. The tailings operating costs are calculated based on a few assumptions: • Under one (1) Mt of coarse tailings material is re-handled each year from the conveyor; • The coarse tailings follows the same haul route as the waste rock material; • The time for loading coarse tailings material from conveyor is ten (10) minutes; • One (1) truck is required to re-handle all of the coarse tailings material from the conveyor. The operating costs for the one (1) truck required for the handling of coarse tailings material incorporate the total fuel cost, the equipment operating cost, and the cost associated to the operators who are assigned to that truck.

21.3.3 Processing Operating Costs

The estimated annual operating costs for the process plant are summarized in Table 21.17 for unit costs. The breakdown of the major components such as manpower, electrical loading, consumables, reagents, and maintenance supplies are shown by sector. The corresponding unit cost per tonne of mineralized material milled is also provided.

The plant operating costs were derived from Project specific process engineering, from supplier information, BBA’s internal database, and benchmarked from similar operations.

Table 21.17 – Process Plant Operating Cost by Sector $/tonne $/tonne of Sector milled concentrate 4100 Primary Crushing 0.20 1.03 Primary Crusher Liners 0.09 0.44 Power 0.11 0.59 5100 Secondary & Tertiary Crushing and Fine Ore Storage 0.47 2.39 Secondary and tertiary crusher liners 0.18 0.92

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$/tonne $/tonne of Sector milled concentrate Screen Decks 0.03 0.13 Power 0.26 1.34 6100 DMS Circuit 0.78 3.99 Ferrosilicon 0.44 2.28 DMS Cyclones 0.05 0.24 Screen Decks 0.07 0.35 Power 0.22 1.12 6210 Grinding 2.04 10.48 Mill Liners 0.17 0.85 Grinding Media 1.54 7.92 Power 0.33 1.70 6220/6230 DeSliming and Flotation Circuit 3.89 19.97 Dispersant and Flotation Reagents 3.56 18.26 Steam Power 0.08 0.40 Power 0.25 1.31 6240/6500 Filtering and Final Product 0.05 0.25 Filtration media 0.01 0.07 Power 0.04 0.18 6610/6630 Reagent Preparation 0.04 0.22 Power 0.04 0.22 6800 Tailings Disposal 0.31 1.57 Flocculant 0.20 1.05 Power 0.10 0.52 6970/6980/6990 Water and Compressed Air 0.37 1.88 Power 0.37 1.88 Process Plant General Operating Costs 7.35 37.67 Manpower (hourly) 4.42 22.64 Manpower (staff) 1.28 6.56 Maintenance Costs (3% of Mechanical Equipment 0.86 4.43 Cost) Laboratory Supplies 0.07 0.37 Building Heating, Electricity (HVAC) 0.40 2.03 Effluent Water Treatment, Reagent 0.32 1.63 Process Plant Operating Cost 15.49 79.44

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a) Plant Manpower In the process plant, it is estimated that 78 employees will be required for operation. This includes the supervisory staff for the process plant, crushers and assay laboratory, as well as hourly staff for the operation, laboratory, mechanical and electrical repairmen.

Selected manpower costs presented in Table 21.18 demonstrate the salaries and manpower loading rates across various salary grades estimated for this Report.

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Table 21.18 – Process Plant Manpower Basic Yearly Total Yearly Total Annual Benefits Unit Cost Process Plant Manpower No. Salary Salary Cost $/y $/t milled $/y $/y $/y Staff Mill Superintendent 1 122,000 51,000 173,000 173,000 0.16 Mechanical/Electrical Superintendent 2 106,000 45,000 151,000 302,000 0.28 Metallurgist 1 81,000 34,000 115,000 115,000 0.11 Shift Foreman 5 70,000 29,000 99,000 495,000 0.45 Chief Chemist 1 84,000 35,000 119,000 119,000 0.11 Work Planner 2 70,000 29,000 99,000 198,000 0.18 Subtotal 12 1,402,000 1.28 Hourly Secretary/Clerk 2 49,000 18,000 67,000 134,000 0.12 Crushing and Loader Operator 10 53,000 20,000 73,000 730,000 0.67 DMS Operator 5 53,000 20,000 73,000 365,000 0.33 Laborer/Helper/Loader Operator 10 44,000 16,000 60,000 600,000 0.55 Grinding Operator 5 53,000 20,000 73,000 365,000 0.33 Flotation Operator 5 53,000 20,000 73,000 365,000 0.33 Tailings Operator 5 53,000 20,000 73,000 365,000 0.33 Reagent Operator/Concentrate Filtration Operator 5 53,000 20,000 73,000 365,000 0.33 Mechanical Maintenance 8 62,000 23,000 85,000 680,000 0.62 Electrician / Instrumentation 6 62,000 23,000 85,000 510,000 0.47 Laboratory Technician 2 58,000 21,000 79,000 158,000 0.14 Lab Labour-Sample Prep Etc 2 44,000 16,000 60,000 120,000 0.11 Metallurgical Technician 1 58,000 21,000 79,000 79,000 0.07 Subtotal 66 4,836,000 4.42 Total Process Plant 78 6,238,000 5.70

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b) Plant Electrical Electrical power is required for assorted equipment in the process plant, including: crushers, grinding mills, conveyors, pumps, agitators, services (compressed air and water), lighting, heating, etc.

The unit cost of electrical power was established at 0.048$/KWh. Details of the annual electrical costs for the process plant are given in Table 21.19. The annual estimated cost is $2.62 million. The average unit cost is $2.25/t milled.

Table 21.19 – Process Plant Electrical Power Costs

Sector Sector Sector 5000 6000 HVAC 4000 Secondary Unit Mineral Process Total Primary and Processing Plant Crushing Tertiary Plant Crushing Load kW 295.8 672.7 3,493.7 1,364.2 5,485.3 kWh Consumption per Year 2.6 5.89 30.6 11.95 51.04 (x106) Consumption kWh/t 2.4 5.4 28.0 10.91 46.7 Annual Cost $/y 125,359 285,104 1,480,613 578,153 2,622,360 Unit Cost $/t milled 0.11 0.26 1.35 0.53 2.25 c) Crushing Consumables Crushing consumables include the liners for the jaw crusher, secondary and tertiary cone crusher, and the screen decks for cone crusher scalping screen. The wear life and cost for these components were obtained from manufacturer life estimation and quote respectively. The unit cost for the crushing consumables is estimated at $0.30/t. Freight and applicable surcharges to the mine site were estimated at 3% of the consumable costs. Table 21.20 shows the cost for the crushing consumables.

Table 21.20 – Consumables Costs for Crushing Consumption Consumables Costs Unit Cost Sets/y $/set $/y $/t milled Jaw Crusher Liners 6 15,750 94,500 0.09 Secondary Cone Crusher Liners 12 9,563 114,753 0.10 Tertiary Cone Crusher Liners 8 10,291 82,325 0.08 Cone Crusher Scalping Screen 3 9,402 28,205 0.03 Decks Total 542,668 0.30

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21.3.4 DMS Consumables

DMS consumables include the ferrosilicon consumption, spare cyclones and screen decks for the product separation screens. Ferrosilicon consumption is based on laboratory test work. An additional DMS cyclone is required for every operating cyclone. There are three (3) cyclones in the first stage of DMS and two (2) in the second stage. A total of five (5) spare cyclones are required. For each of the two (2) separation screens, a spare set of screen decks is required. The cost for these consumables and transport to mine site was obtained from the manufacturer’s quotation. The unit cost for the DMS consumables is $0.56/t. Table 21.21 shows the costs for the DMS consumables.

Table 21.21 – DMS Consumables Costs Consumption Consumables Costs Unit Cost kg/hr $/t $/y $/t milled Ferrosilicon 6 1,543 486,676 0.44 sets/y $/set $/y $/t DMS Cyclones 5 10,300 51,500 0.05 Screen Decks 2 37,400 74,800 0.07 Total 612,976 0.56 a) Grinding Consumables Grinding consumables include rod mill liners, ball mill plate liners and lifter liners, and grinding media for the rod and ball mill. Wear life for the liners was obtained from the manufacturer’s life estimation. The grinding media consumption has been estimated using the Bond Abrasion Index equation and the power input into the material. Media will be added as required in each mill. The cost for the grinding consumables was obtained from the supplier’s quotation. The price includes delivery to the mine site and applicable surcharges. The unit cost for the grinding consumables is estimated at $1.71/t. Grinding consumables with their respective cost estimate is detailed in Table 21.22.

Table 21.22 – Grinding Consumables Costs Consumption Consumables Costs Unit Cost kg/t $/t $/y $/t milled Grinding rods 0.48 1,175 617,465 0.56 Grinding balls 0.69 1,430 1,073,210 0.98 sets/y $/set $/y $/t Rod mill liners 0.5 157,500 78,750 0.07 Ball mill plate liners 1.0 28,900 28,900 0.03 Ball mill lifter liners 2.0 37,400 74,800 0.07 Total 1,871,626 1.71

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b) Reagents The reagents are primarily used in the desliming, flotation and thickening process. The reagents include: Dispersant D618, Soda Ash (Na2CO3), Caustic (NaOH), Flocculant Magnafloc 10, Collector LR19, Collector Armac C and Fuel oil.

The Collector LR19 accounts for the highest reagent cost. LR19 is used during flotation and is added to the conditioner tank and flotation cleaner cells for lithium recovery.

The consumption rate for these reagents was obtained by laboratory test work. The cost and transport to the mine site were obtained from the supplier’s quotation. The unit cost for the reagents is $3.77/t. Table 21.23 shows the cost for the reagents.

Table 21.23 – Reagent Consumables Costs Consumption Reagent Costs Unit Cost g/t t/y *$/t **$/y $/t milled Dispersant D618 249.6 156.7 3,020.0 489,071 0.45 Soda Ash 162.4 101.9 375.0 45,136 0.04 Caustic 278.4 174.7 410.0 83,511 0.08 Flocculant Magnafloc 10 276.0 50.8 4,300.0 223,497 0.20 Collector LR19 1,132.9 711.1 4,310.8 3,065,262 2.80 Collector Armac C 63.3 39.7 4,980.0 197,853 0.18 Fuel Oil 21.6 13.6 1,415.4 19,189 0.02 Total 4,123,520 3.77 * Transport costs not included ** Transport costs included

c) Other Consumables Other consumables include activated carbon for water treatment; filter cloths for the tailing filter press, and laboratory supplies. Activated carbon consumption and laboratory supplies are estimated based on recent similar projects. Tailings filter cloth consumption is based on the manufacturer’s estimation. Consumable costs and transport are based on the supplier’s quotation and BBA’s internal data. Table 21.24 shows the cost for these consumables.

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Table 21.24 – Other Consumables Costs Consumption Consumables Costs Unit Cost kg/t $/t $/y $/t milled Activated Carbon 0.504 630 347,684 0.32 sets/y $/set $/y $/t Filter Cloth 2 7,784.7 15,569 0.01 Laboratory Supplies 80,000 0.07 Total 0.40 d) Maintenance Supplies The annual costs for plant maintenance supplies are estimated by applying a ratio of 3% to the mechanical cost of the process plant equipment, excluding installation. The anticipated annual costs for the maintenance supplies are estimated to be $0.86/t.

21.3.5 General and Administration

The General and Administration (“G&A”) costs include all materials and personnel costs associated with the site administration and camp support.

The G&A costs for the Project, including camp, are estimated at 7.58 M$/y of operation and include costs for manpower, as well as costs for material and services related to administration, as summarized in Table 21.25.

Table 21.25 – G&A Operating Costs by Sector Description Employees Costs Unit Cost $/y $/t milled Administration – Manpower 14 1,297,000 1.18 Administration – Material & Services 1,306,000 1.19 Camp Cost 3,676,653 3.35 Employee Transport 1,300,000 1.19 Total 7,579,653 6.91 a) General and Administration – Manpower It is estimated that 14 employees will be required for the Administration, Safety and Security manpower. The unit cost for G&A – Manpower is 1.18$/t. Table 21.26 shows the summary of G&A manpower costs.

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Table 21.26 – G&A Manpower Basic Total Total Unit Administration, Safety & Yearly Benefits Yearly Annual No. Cost Security Salary $/y Salary Cost $/t milled $/y $/y $/y General Manager 1 182,000 76,000 258,000 258,000 0.24 HR and H&S Superintendent 1 84,000 35,000 119,000 119,000 0.11 First Aid/H&S Technician 2 64,000 27,000 91,000 182,000 0.17 Purchasing Agent 1 77,000 32,000 109,000 109,000 0.10 Secretary/Clerk 2 49,000 18 000 67,000 134,000 0.12 Security 5 49,000 21,000 70,000 350,000 0.32 Trainer 1 49,000 21,000 70,000 70,000 0.06 IT technician 1 53,000 22,000 75,000 75,000 0.07 Total 14 1,297,000 1.18 21.3.6 General and Administration – Material and Services

The unit cost for G&A – Material and Services is estimated at $1.19/t. The G&A estimated annual costs for materials and services are provided in Table 21.27.

Table 21.27 – Process Plant Operating Costs by Sector Average Unit Cost Item Annual Cost $/t milled $/yr Mining Leases 60,000 0.05 Site Insurance 400,000 0.37 Consulting Fees 80,000 0.07 Telecommunication Fees 45,000 0.04 Office and IT Supplies 50,000 0.05 IT Maintenance 20,000 0.02 Warehouse Miscellaneous Supplies 10,000 0.01 Recruiting Fees 50,000 0.05 Safety Equipment 10,000 0.01 Medical Supplies 10,000 0.01 Employee Relations 40,000 0.04 Community Relations 50,000 0.05 Environmental Sampling and Supplies 125,000 0.11 Internships 42,000 0.04 Mining Association Membership 24,000 0.02

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Average Unit Cost Item Annual Cost $/t milled $/yr Infrastructure Maintenance 200,000 0.18 Light Vehicle Maintenance and Fuel 80,000 0.07 Small tools 10,000 0.01 Total 1,306,000 1.19 a) Camp Costs Camp costs were obtained from supplier quotation. The camp is sized for a total of 125 rooms. The unit cost of the camp is estimated at $3.35/t. The camp breakdown presented in Table 21.28.

Table 21.28 – Camp Operating Costs by Sector Cost Unit Cost

$/y $/t milled Permanent Camp Rental 1,536,000 1.40 Camp Maintenance 144,000 0.13 Catering 1,825,000 1.67 Camp Security 108,000 0.10 Camp Electricity 54,750 0.05 Total 3,667,750 3.35 b) Employee Transport The cost for the transportation of employees is for chartering a 50-passenger airplane between Montreal or Quebec and the Nemiscau airport, with stopovers in Chibougamau and Saguenay. A budget price was obtained at $25,000/week. It amounts to a total of 1.3 M$/y.

21.3.7 Concentrate Transport

The cost for concentrate transport includes truck transport from the mine site to the Chibougamau rail track, loading of concentrate into railcars, rail transport to the Hydrometallurgical plant in Valleyfield. The cost for storage at Chibougamau is not included in the operating costs since Nemaska plans to build a pad for storage and this is therefore included in the capital cost estimate. The cost for the concentrate transport was obtained from different transport and handling service companies.

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21.3.8 Hydrometallurgical Plant Operating Costs

Operating costs were estimated for the Hydrometallurgical Plant and cover the costs related to the transformation of spodumene concentrate into lithium hydroxide and lithium carbonate.

The operating costs are based on a concentrate annual feed rate of 213,558 tonnes, a lithium hydroxide monohydrate production of 20,734 t/y and lithium carbonate production of 10,000 t/y.

The sources of information used to develop the operating costs include in-house databases and outside sources particularly for reagents and consumables. a) Operating Costs Summary The plant life average operating cost estimate is summarized in Table 21.29.

Table 21.29 – Average Operating Cost Estimate ($/year) Average Operating Cost Area ($/year) Manpower 6,517,339 Electrical Power 10,561,465 Consumables and Wear Parts 10,275,374 Reagents and Chemicals 14,939,150 Natural Gas 5,222,295 Site Material Handling 1,423,241 Total Operating Costs 48,938,864 b) Manpower Requirement Table 21.30 presents the estimated personnel requirements for the hydrometallurgical plant operation by area. Administration employees responsible for both the concentrator and the hydrometallurgical operations are based in Valleyfield and are accounted for in the hydrometallurgical plant operating costs.

Table 21.30 – Total Personnel Requirement Area Number Administration 14 Operations 43 Maintenance 14 Metallurgy 10 Total Manpower 81

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Manpower consists of 31 staff employees and 50 hourly employees. Total annual costs for the above manpower including base salary, expected overtime and fringe benefits have been estimated at $ 6.5 M. c) Electrical Power The hydrometallurgical plant total operating electrical power is 27 MW resulting in annual costs of $ 10.6 M. Approximately 80% of the demand is from the electrodialysis process. Electricity cost ($/kWh) was provided by Hydro-Quebec. d) Consumables and Wear Parts Consumables include belt filter cloths, kiln liner replacement, ion exchange resin replacement, final product bags, electrodialysis membranes, etc. Also included in this cost category are potable water make-up costs (purchased from the City of Valleyfield) and liquid tailings disposal costs. Potable water ($/m3) and tailings disposal ($/kg of impurity) costs were provided by the City of Valleyfield.

Annual consumables and wear parts costs total $ 10.3 M. Similarly to the electrical costs, about 80% of the consumables and wear parts costs are related to the electrodialysis process (membrane replacement). e) Reagents and Chemicals Lime, soda ash, sulfuric acid, hydrochloric acid and carbon dioxide are required by the hydrometallurgical process. Reagents and chemicals costs total $ 14.9 M annually. Lime and sulfuric acid costs represent almost 80% of the total. f) Natural Gas Natural gas is used to fire the kiln, both product dryers and the boiler. Annual costs reach $ 5.2 M. Natural gas cost ($/m3) was provided by Gaz Metro.

21.3.9 Site Material Handling

Mobile equipment is used for continuous plant work, concentrate unloading and reclaim and final product handling. Personnel pick-up trucks costs are also included in this cost category. Annually, an estimated $ 1.4 M is spent for site material handling and personnel transportation on site.

21.3.10 Operating Costs per Product

In order to determine operating costs for each of the hydrometallurgical plant final products, lithium hydroxide solution production costs were first established. Once this base value was known, additional processing costs for converting this lithium hydroxide

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solution into solid lithium hydroxide monohydrate and lithium carbonate were calculated. The operating cost summary per product is presented in Table 21.31.

Table 21.31 – Average Operating Cost Estimate (CAD$/year) Cost Cost Adder Operating Cost % of Cost Cost Adder Cost ($/y) ($/tonne ($/t LiOH- Area Total Costs ($/t LiOH) ($/t Li2CO3) milled) H2O) Manpower 6,517,339 30.52 13.3% 291.88 14.35 14.04 Electrical Power 10,561,465 49.45 21.6% 472.23 18.27 34.72 Consumables & 10,275,374 48.12 21.0% 455.88 26.99 22.06 Wear Parts Reagents & 14,939,150 69.95 30.5% 659.65 0.00 120.00 Chemicals Natural Gas 5,222,295 24.45 10.7% 163.99 68.19 39.29 Site Material 1,423,241 6.66 2.9% 34.17 17.16 35.58 Handling Total 48,938,864 229.16 100.0% 2,077.78 144.97 265.69 Based on the above values, yearly production costs were calculated. The LiOH production cost (based on 20,828 t/y) is 43,276,098 $/y and the cost adders for LiOH- H2O (20,734 t/y) and Li2CO3 (10,000 t/y) are 3,005,861 $/y and 2,656,905 $/y respectively.

To determine each final product’s total production cost, the LiOH cost was distributed according to the solution split between the hydroxide monohydrate and the carbonate circuit (73.6% to LiOH-H2O and 26.4% to Li2CO3).

The LiOH distributed cost and the adders are combined and the operating cost is obtained by dividing this combined cost by the annual production tonnage. Results are shown in Table 21.32.

Table 21.32 – Hydrometallurgical Plant Operating Cost Estimate (CAD$/tonne product) Operating Cost Operating Cost ($/t LiOH-H2O) ($/t Li2CO3) 1,681 1,408

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22.0 ECONOMIC ANALYSIS

22.1 General

A preliminary economic analysis of the Project has been carried out using a cash flow model. The model is constructed using annual cash flows in constant money terms (third quarter 2012). No provision is made for the effects of inflation. As required in the financial assessment of investment projects, the evaluation is carried out on a so-called “100% equity” basis, i.e. the debt and equity sources of capital funds are ignored. Results are presented before and after taxation.

The model reflects the base case macro-economic and technical assumptions given in this Report.

22.2 Assumptions

22.2.1 Macro-Economic Assumptions

The main macro-economic assumptions used in the base case are given in Table 22.1.

The price forecasts for Lithium Hydroxide Monohydrate (LiOH-H2O) and Lithium Carbonate (Li2CO3) are based on projections from Roskill and SignumBOX studies.

The sensitivity analysis examines a range of prices 30% above and below these base case prices.

Table 22.1 – Macro-Economic Assumptions

Item Unit Base Case Value Lithium hydroxide monohydrate (LiOH-H O) 2 USD/tonne 8,000 Price Forecast (F.O.B. hydro-met plant) Lithium carbonate (Li CO ) 2 3 USD/tonne 6,500 Price Forecast (F.O.B. hydro-met plant) Exchange Rate CAD/USD 1.00 Discount Rate % per year 8 Discount Rate Variants % per year 6 and 10 The current Canadian tax system applicable to Mining Resources Income is used to assess the Project’s annual tax liabilities. This consists of federal and provincial corporate taxes as well as provincial mining taxes (mining taxes in Québec were revised in the 2010 budget). The federal and provincial corporate tax rates currently applicable over the Project’s operating life are 15.0 % and 11.9 % of taxable income, respectively. The rate applicable for the purpose of assessing mining taxes is 16 % of taxable income. It is uncertain at this time whether the 13 % processing allowance rate (in the Quebec Mining

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Tax legislation) associated with processing to a more advanced stage within the province would be applicable in this instance. For this reason, the more conservative processing allowance rate of 7 % is used until confirmation can be obtained from the “Ministère des Ressources naturelles et de la Faune”.

Apart from the base case discount rate of 8 %, two (2) variants of 6 and 10 % are used to determine the net present value of the Project. These discount rates represent possible weighted-average costs of capital to the investor.

22.2.2 Mineral Royalties

The present financial analysis incorporates a net profit royalty agreement. The royalty payment is based on 3 % of the “unit value” of the concentrate produced at the mill. The unit value of the concentrate is calculated from the pre-established sales price of $300 and operating cost of $200 per tonne, representing unit revenue and operating expenses, respectively, less all capital expenses incurred over the life of the Project, expressed per tonne of estimated concentrate production. Based on the capital expense estimates associated with this Project, the unit value of the concentrate amounts to $55.51 per tonne, leading to a royalty payment of $1.665 per tonne of concentrate produced.

22.2.3 Technical Assumptions

The main technical assumptions used in the base case are given in Table 22.2.

Table 22.2 – Technical Assumptions Total Resource Mined M tonnes 19.6

Average Mill Head Grade % LiO2 1.49 Design Mining Rate M tonnes/year 1.095 Mine Life years 19.1 Process Recovery % 78.3

Concentrate Grade % LiO2 6.0 Total Concentrate Production M tonnes 3.828 Loss of Concentrate during Transport % 1.0 Total Concentrate Processed at Hydro-met Plant M tonnes 3.790 Average Mining Costs ($/tonne mined) 15.86 Mill Processing Costs ($/tonne concentrate) 79.44 Tailings & Middlings Haulage Costs ($/tonne concentrate) 4.68 General & Administration Mine Site Costs ($/tonne concentrate) 35.43 Concentrate Transport Costs ($/tonne concentrate) 60.00 Hydro-metallurgical Processing Costs ($/tonne concentrate) 229.16

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A reduced production of 894 Kt milled in the first production year provides for a ramp-up to the full capacity of 1 095 Kt milled per year. The amount of concentrate produced is a function of the mill head grade and milling process recovery and thus, varies from 178 Kt to 225 Kt per year.

22.3 Financial Model and Results

A summary of the base case results is given in Table 22.3. The cash flow statement for the base case is given in Table 22.4.

The summary and cash flow statement indicate that the total pre-production capital costs are evaluated at $438.8 M and the sustaining capital requirement is evaluated at $8.8 M, for a total of $447.7 M over the life of the Project. The mine rehabilitation trust fund payments are estimated at $1.9 M.

For taxation purposes, the indirect costs and contingency components of the pre- production capital cost estimate were redistributed by area as shown in the cash flow statement of Table 22.4. The cash flow statement shows a capital cost breakdown by area and provides a preliminary capital spending schedule over the 2-year pre-production period of the Project. Working capital requirements are estimated at 2 months of total annual operating costs. As operating costs vary annually over the mine life, additional amounts of working capital are injected or withdrawn as required.

The total revenue is estimated at $4,083.1 M or $207.91/tonne milled. The total operating costs are estimated at $1,875.2 M or $95.48/tonne milled.

The financial results indicate a before-tax Net Present Value (“NPV”) of $567.2 M at a discount rate of 8%. The before-tax Internal Rate of Return (“IRR”) is 23.3% and the payback period is 3.9 years.

The after-tax Net Present Value is $330.5 M at a discount rate of 8%. The after-tax Internal Rate of Return is 18.9% and the payback period is 4.0 years.

Table 22.3 – Project Evaluation Summary – Base Case Value Item (CAD$ M) Total Revenue 4,083.1 Total Operating Costs 1,875.2 Pre-production Capital Costs 438.8 Sustaining Capital Costs 8.8 Mine Rehabilitation Trust Fund Payments 1.9 Total Before-tax Cash Flow 1,758.4

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Value Item (CAD$ M) Before-tax NPV @ 6% 752.9 Before-tax NPV @ 8% 567.2 Before-tax NPV @ 10% 424.1 Before-tax IRR (%) 23.3 Before-tax Payback Period (years) 3.9 Total After-tax Cash Flow 1,123.8 After-tax NPV @ 6% 455.1 After-tax NPV @ 8% 330.5 After-tax NPV @ 10% 234.1 After-tax IRR (%) 18.9 After-tax Payback Period (years) 4.0

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Table 22.4 – Cash Flow Statement

Years -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total All Monetary Values in ‘000 CAD Resources Mined ('000 t) 894 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 1,095 130 19,639

Grade (% LiO2) 1.528 1.514 1.542 1.454 1.426 1.483 1.542 1.576 1.533 1.437 1.452 1.454 1.468 1.444 1.483 1.495 1.530 1.529 1.528 1.494 Waste ('000 t) 642 1,601 2,119 2,320 2,605 2,983 3,154 4,016 4,851 4,848 5,878 4,888 4,580 4,299 3,793 2,624 1,767 881 665 62 58,575

Stripping ratio (w : o) 1.792 1.935 2.118 2.379 2.724 2.880 3.668 4.430 4.427 5.368 4.464 4.182 3.926 3.464 2.397 1.614 0.805 0.607 0.479 2.983

Concentrate Production ('000 t) 178 216 220 208 204 212 220 225 219 205 207 208 210 206 212 214 219 218 26 3,828 Less Handling Losses ('000 t) 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 0 38 Concentrate Processed ('000 t) 176 214 218 206 202 210 218 223 217 203 205 206 208 204 210 211 216 216 26 3,790 LiOH Production ('000 t) 17 21 21 20 20 20 21 22 21 20 20 20 20 20 20 21 21 21 3 370

LiOH-H2O product ('000 t) 17 21 21 20 19 20 21 22 21 20 20 20 20 20 20 20 21 21 2 366

Li2CO3 product ('000 t) 8 10 10 10 9 10 10 10 10 10 10 10 10 10 10 10 10 10 1 177

LiOH-H2O Sales 136,346 165,556 168,618 158,995 155,933 162,166 168,618 172,336 167,634 157,136 158,776 158,995 160,526 157,902 162,166 163,478 167,306 167,196 19,888 2,929,571

Li2CO3 Sales 53,687 65,189 66,394 62,605 61,400 63,854 66,394 67,858 66,007 61,873 62,519 62,605 63,208 62,175 63,854 64,371 65,878 65,835 7,831 1,153,539 Total Revenue 190,033 230,745 235,012 221,601 217,333 226,020 235,012 240,194 233,641 219,010 221,296 221,601 223,734 220,076 226,020 227,849 233,183 233,031 27,719 4,083,110 Total Revenue 190,033 230,745 235,012 221,601 217,333 226,020 235,012 240,194 233,641 219,010 221,296 221,601 223,734 220,076 226,020 227,849 233,183 233,031 27,719 4,083,110

Mining Costs 14,345 15,555 16,459 16,871 18,349 18,271 19,729 21,047 20,563 21,407 20,582 20,428 20,227 19,356 15,100 13,452 9,921 7,903 1,859 311,425 Processing Costs 14,154 17,187 17,504 16,505 16,188 16,835 17,504 17,890 17,402 16,313 16,483 16,505 16,664 16,392 16,835 16,971 17,368 17,357 2,065 304,123 Tailings & Middlings Haulage 834 1,013 1,031 972 954 992 1,031 1,054 1,025 961 971 972 982 966 992 1,000 1,023 1,023 122 17,917 Costs Concentrate Transport Costs 10,690 12,981 13,221 12,466 12,226 12,715 13,221 13,512 13,144 12,321 12,449 12,466 12,586 12,381 12,715 12,818 13,118 13,109 1,559 229,700 G&A Mine Site Costs 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 7,566 901 137,095 Hydro-metallurgical Processing 40,422 49,082 49,990 47,137 46,229 48,077 49,990 51,092 49,698 46,586 47,072 47,137 47,591 46,813 48,077 48,466 49,601 49,569 5,896 868,527 Costs Royalty Payments 297 360 367 346 339 353 367 375 365 342 345 346 349 344 353 356 364 364 43 6,374 Total Operating Costs 88,309 103,744 106,139 101,865 101,852 104,809 109,409 112,538 109,763 105,496 105,469 105,422 105,966 103,817 101,638 100,629 98,962 96,890 12,444 1,875,160 Total Operating Costs 88,309 103,744 106,139 101,865 101,852 104,809 109,409 112,538 109,763 105,496 105,469 105,422 105,966 103,817 101,638 100,629 98,962 96,890 12,444 1,875,160

Operating Profit 101,724 127,001 128,874 119,736 115,481 121,211 125,604 127,657 123,878 113,514 115,827 116,178 117,768 116,259 124,382 127,220 134,222 136,141 15,274 2,207,950

Mine Site Pre-production

Capital Expenditure MINE DEVELOPMENT : Pre- 0 2,476 2,476 Stripping Infrastructure 1,321 1,981 3,302 Administration and Services 2,182 3,273 5,455 Mine Equipment and 6,948 10,422 17,370 Infrastructure Crushing 5,800 8,699 14,499 Processing Plant 46,281 69,422 115,704 Tailings and Waste 1,148 1,723 2,871 Management Hydro-Metallurgical Plant

Capital Expenditure Process 100,831 151,246 252,077 Tailings and Water 982 1,473 2,456

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Years -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total Management Facilities General Services 1,207 1,810 3,017 Infrastructure 2,113 3,169 5,282 Power and Communications 5,282 7,923 13,205 Service Vehicles and 454 681 1,134 Equipment Total 174,548 264,299 438,847

Working Capital 0 14,718 2,572 399 -712 -2 493 767 521 -462 -711 -5 -8 91 -358 -363 -168 -278 -345 -14,074 -2,074 0

Sustaining Capital Expenditure General – Mine Site - - 0 965 738 250 738 738 738 738 250 250 2,160 250 250 250 250 250 0 0 0 8,815 Total - - 0 965 738 250 738 738 738 738 250 250 2,160 250 250 250 250 250 0 0 0 8,815 Total Capital Expenditure 174,548 279,017 2,572 1,364 26 248 1,231 1,505 1,259 276 -461 245 2,152 341 -108 -113 82 -28 -345 -14,074 -2,074 447,662

Trust Fund Rehabilitation 0 962 481 481 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 1,923 Payments – Mine Rehabilitation Costs – Hydro- 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 Metallurgical Plant

Federal Corporate Income Tax 0 0 0 0 0 3,226 15,850 16,315 16,688 16,820 16,270 14,914 15,112 15,146 15,340 15,146 16,168 16,524 17,414 17,662 2,103 230,697 Provincial Corporate Income 0 0 0 0 0 2,559 12,575 12,944 13,239 13,344 12,908 11,832 11,989 12,015 12,169 12,016 12,826 13,109 13,815 14,012 1,669 183,020 Tax Quebec Mining Tax 0 0 0 733 5,739 7,644 9,376 11,930 13,780 14,911 14,889 13,636 14,202 14,486 14,900 14,770 16,149 16,658 17,830 18,172 1,083 220,889 Total Corporate Income and 0 0 0 733 5,739 13,429 37,801 41,189 43,708 45,075 44,067 40,381 41,302 41,647 42,409 41,932 45,143 46,292 49,059 49,845 4,856 634,606 Mining Taxes

BEFORE-TAX CASH FLOW -174,548 -279,979 98,670 125,156 128,848 119,488 114,251 119,707 124,344 127,381 124,339 113,268 113,675 115,838 117,876 116,372 124,300 127,248 134,567 150,215 17,348 1,758,364 Cumulative B-T CF -174,548 -454,527 -355,857 -230,700 -101,852 17,635 131,886 251,593 375,937 503,318 627,657 740,925 854,600 970,437 1,088,314 1,204,686 1,328,986 1,456,234 1,590,801 1,741,016 1,758,364 0

Payback period work area 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0

AFTER-TAX CASH FLOW -174,548 -279,979 98,670 124,424 123,109 106,059 76,450 78,518 80,637 82,306 80,272 72,887 72,373 74,191 75,467 74,441 79,157 80,956 85,508 100,370 12,493 1,123,759 Cumulative A-T CF -174,548 -454,527 -355,857 -231,433 -108,324 -2,265 74,185 152,702 233,339 315,645 395,917 468,804 541,176 615,367 690,834 765,275 844,432 925,388 1,010,896 1,111,266 1,123,759 0

Payback period work area 0 1.00 1.00 1.00 1.00 1.00 0.03 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0

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22.4 Sensitivity Analysis

A sensitivity analysis has been carried out, with the base case described above as a starting point, to assess the impact of changes in total pre-production capital expenditure (“CAPEX”), operating costs (“OPEX”) and product prices on the Project’s NPV @ 8 % and IRR. Each variable is examined one-at-a-time. An interval of ±30% with increments of 10% was used for all three variables.

The before-tax results of the sensitivity analysis, as shown in Figure 22.1 and Figure 22.2, indicate that, within the limits of accuracy of the cost estimates in this study, the Project’s before-tax viability does not seem significantly vulnerable to the under- estimation of capital and operating costs, taken one at-a-time. As seen in Figure 22.1, the net present value is more sensitive to variations in operating expenses than pre- production CAPEX, as shown by the steeper slope of the OPEX curve. As expected, the net present value is most sensitive to variations in price (both product prices are varied together). The net present value becomes marginal at the lowest price variation in the interval, i.e., at -30 % (this corresponds to a price combination of $5,600 and $4,550 per tonne for lithium hydroxide monohydrate and lithium carbonate, respectively). A reduction of about 31% in both prices is found to result in a break-even net present value @ 8%. Figure 22.1, showing variations in internal rate of return, provides the same conclusions. As seen, the IRR becomes marginal (just above the 8% IRR break-even dashed line) at the lowest price variation of the interval. Compared to Figure 22.1, which shows linear variations in net present value for the three (3) variables studied, variations associated with internal rate of return are not linear. Because of the different timing associated with pre-production CAPEX versus OPEX, the IRR is more sensitive to negative variations in pre-production CAPEX than OPEX, but remains less sensitive for positive variations, as evidenced by the shape of the CAPEX curve.

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Figure 22.1 – Before-Tax NPV8%: Sensitivity to Capital Expenditure, Operating Cost and Price

1200

1000 mil.)

800 ($

600 @

NPV 400

T ‐ B

200

0 ‐30 ‐20 ‐100 102030 RELATIVE VARIATION (%)

CAPEX OPEX PRICE

Figure 22.2 – Before-Tax IRR: Sensitivity to Capital Expenditure, Operating Cost and Price

40.0

35.0

30.0

25.0 (%)

20.0 IRR

T ‐

B 15.0

10.0

5.0

0.0 ‐30 ‐20 ‐100 102030 RELATIVE VARIATION (%)

CAPEX OPEX PRICE

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The after-tax results of the sensitivity analysis are shown in Figure 22.3 and Figure 22.4. Figure 22.3 indicates that the Project’s after-tax viability is mostly vulnerable to a price forecast reduction while being less affected by under-estimation of capital and operating costs. A reduction of about 28% in both price forecasts (down to $5,760 and $4,680 per tonne for lithium hydroxide monohydrate and lithium carbonate, respectively) results in a break-even net present value @ 8%.

Figure 22.3 – After-Tax NPV8%: Sensitivity to Capital Expenditure, Operating Cost and Price

800

600 mil.)

($ 400

200 NPV

T ‐ A 0

‐200 ‐30 ‐20 ‐100 102030 RELATIVE VARIATION (%)

CAPEX OPEX PRICE

Figure 22.4, showing variations in internal rate of return, provides the same conclusions. Here, the price curve intersects the 8% IRR break-even dashed line at the same relative variation of about -28%.

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Figure 22.4 – After-Tax IRR: Sensitivity to Capital Expenditure, Operating Cost and Price

30.0

25.0

20.0 (%)

15.0 IRR

T ‐ A 10.0

5.0

0.0 ‐30 ‐20 ‐10 0 10 20 30 RELATIVE VARIATION (%)

CAPEX OPEX PRICE

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23.0 ADJACENT PROPERTIES

Significant adjacent properties in the immediate surroundings of the Whabouchi property include ten (10) properties owned by Monarques Resources Inc. (“Monarques”), the advanced stage Nisk property which hosts the Nisk-1 Ni-Cu-Co-PGE deposit, and nine (9) early stage properties to the east and to the west of the Property, which show potential for hosting magmatic and volcanogenic sulfides mineralization as well as spodumene- bearing pegmatites. Tucana Lithium owns an exploration property located in between the Whabouchi Property and the Nisk property. Figure 23.1 shows the location of adjacent properties surrounding the Whabouchi property.

Figure 23.1 – Location Map Showing Adjacent Mineral Properties

Monarques’ Lac Levac property hosts the Nisk-1 deposit, an elongated body of serpentinized ultramafic rocks that intrude paragneiss and amphibolite sequences. The ultramafic rock intrusion is interpreted as a sill composed of at least two (2) distinct ultramafic lithological units: a grey serpentinized peridotite with magnetite veinlets, and a black serpentinized peridotite with chrysotile veinlets hosting the Ni-Cu-Co-PGE sulphide mineralization. The Nisk-1 deposit hosts NI 43-101 compliant mineral resources totalling 1.25 Mt grading 1.1% Ni, 0.6% Cu and 1.3 g/t Pt+Pd in the measured category,

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0.78 Mt grading 1.0% Ni, 0.5% Cu and 1.2 g/t Pt+Pd with an additional 1.0 Mt grading 0.8% Ni, 0.3% Cu and 1.5 g/t Pt+Pd (Théberge – June 2011).

Other exploration properties also developing spodumene-bearing pegmatites are located in the region surrounding the Whabouchi Property. The Rose-Pivert property owned by Critical Element Corp. and located 47 km northwest of Whabouchi hosts an NI 43-101 compliant mineral resources of 11.4 Mt grading 1.34% Li2O in the Indicated category with an additional 2.2 Mt grading 1.27% Li2O in the Inferred category at 0.75% Li2O cut- off grade (InnovExplo – Dec 2010). Galaxy Resources Inc. lithium property located 105 km northwest of Whabouchi hosts an NI 43-101 compliant with in-pit mineral resources of 11.7 Mt grading 1.30% Li2O in the Indicated category with an additional 10.5 Mt grading 1.20% Li2O in the Inferred category at 0.75% Li2O cut-off grade (SRK – November 2010). At 125 km southeast of Whabouchi, the Moblan property jointly owned by Perilya Ltd and Soquem hosts an NI 43-101 compliant with inpit mineral resources totalling 4.5 Mt grading 1.66% Li2O in the Measured category and 4.9 Mt grading 1.43% Li2O in the Indicated category, with an additional 1.1 Mt grading 1.42% Li2O in the Inferred category at 0.6% Li2O cut-off grade (Perilya NR – May 2011).

The information is not necessarily indicative of the mineralization on the Property that is the subject of the technical report. The Qualified Person has been unable to verify the information related to the NI 43-101 mineral resources reported in this section.

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24.0 OTHER RELEVANT DATA AND INFORMATION

24.1 Project Schedule

The Project implementation schedule covers all the areas of the Project and includes the engineering, procurement, construction and commissioning of the facilities, including the main substation, the processing installations and the site infrastructures. Pre-production mining, initial water management, and related activities are also part of the Project implementation schedule.

The processing facility is the most important element in terms of scope and magnitude. As a result, the overall schedule is governed by the engineering, equipment delivery, construction, commissioning, and other activities related to the start-up of the processing facility.

The schedule assumes that the Project execution team will start in the second quarter of 2013. The schedule also assumes that the environmental permits, required to start construction work at the site, will be received by April 1, 2014. The planned production start-up is for the fourth quarter of 2015. Figure 24.1 presents a summary of the Project Schedule.

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Figure 24.1 – Project Schedule

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24.2 Whabouchi Construction

24.2.1 Construction Infrastructure

A construction camp and a permanent camp are required for the Project. These two (2) camps will be built at the location of the Compagnie de Construction et de Développement Crie Ltée (“CCDC”) site. The CCDC will be mandated for the two (2) camps and all required services.

Initially, the permanent camp will be used as a construction camp. A supplemental construction camp (to accommodate the expected peak workforce) will be built as an extension to the permanent camp. In total, 125 rooms in the permanent camp and 90 rooms in the supplemental construction camp will be built for a total of 215 rooms.

The permanent camp will be in place in April 2014 for the start of construction. The supplemental construction camp will be built for September 2014 and will be used until the end of the construction and commissioning period.

Permanent power will be delivered by Hydro-Québec in time for the start of construction. Until the time when the main substation will be commissioned, construction power will be supplied through a temporary installation.

24.2.2 Mining

In 2013, initial orders for mining equipment are expected to be placed. Deliveries will start in the third quarter of 2014 and assembly will have taken place in time for pre- production to start in the first quarter of 2015. Logistics for on-site assembly plans will be made to ensure that all equipment will be operational in time to perform the mine production tonnages corresponding to the mine planning requirements.

24.2.3 Site Infrastructure

Detailed engineering related to site infrastructures is scheduled to start with the award of the EPCM services. Related procurement activities are to follow three (3) months later. Construction of site facilities and infrastructure, such as change house, offices, maintenance facilities etc., is expected to be completed in the third quarter of 2014.

24.2.4 Concentrator Processing Plant a) Processing Plant Implementation Plan The processing plant implementation plan takes into account that detailed engineering is required to start within 12 months prior to the construction start-up. The implementation plan covers the period from the rod mill and ball mill specification preparation up to the end of the processing plant commissioning. The

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schedule is derived from major equipment delivery information obtained from suppliers during this Study. b) Engineering In order to achieve a proposed plant commissioning and start-up in the fourth quarter of 2015, proper sequencing of detailed engineering is fundamental. Enclosing the building before winter conditions prevent efficient mechanical and electrical installation is key to the success of this plan. To achieve this, a pre- engineered building shell-type construction has been selected. Key to this type of installation is freezing the equipment layout and building sizing early. In order to achieve this, all major mechanical equipment with potential impact on these elements will be designed and procured at the onset of the Project.

The detailed engineering for the processing plant is estimated at 18 months and will include the mobile crushing equipment and the process plant. c) Long Lead Items As part of this Study, budgetary prices and delivery periods have been requested from suppliers for major equipment. The equipment fabrication times for the pressure filters, ball and rod mills have a delivery period from 12 to 14 months, followed by the leach tank agitators with a fabrication time of up to 12 months. The implementation schedule allows for 14 months for the crusher and grinding mills to be delivered on site. d) Procurement Plan In order to complete engineering, certified drawings are required. For this reason, efforts will be made to purchase all other equipment early. This has a secondary advantage of securing early fabrication slots, hence reducing the impact of potential fabrication delays. The procurement of equipment will include the preparation of comprehensive engineering and tender documents. e) Construction The first priority is site preparation and temporary installations such as construction management camp and temporary construction power distribution. Early during the site preparation, the concrete contractor will install a batch plant to be ready for the first concrete pour. Construction activities will start with concrete foundations for the grinding mills and concentrator building. The building shell, complete with siding panels, will be erected. Other equipment foundations, elevated concrete slabs and finally the slab on grade in the main process building will precede the erection of the balance of the steel. The structural steel work of the main process building will follow, starting from the grinding area moving northeast along the plant site.

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The building roofing and siding will follow to complete building closure that will allow workers to work in an enclosed area. The internal equipment steel supports and platforms will then precede the grinding mills installation and mechanical, piping and electrical work will continue to completion in June 2015. The commissioning will follow. All other work areas not on the critical path will be optimized, based on resources leveling and availability. f) Commissioning and Start-Up The commissioning activities will be performed by a team comprised of construction managers, contractors, engineers, and operation personnel in order to maximize efficiency. A six (6) month commissioning period has been scheduled. The dry commissioning will start while the construction is still on-going. A detailed commissioning schedule will identify process systems that can be commissioned prior to construction completion and turned over to Operations to reduce the last minute turnover procedures. Commissioning will start with individual equipment evaluation followed by a dry run for a short period. Systems will then be flushed and run with water, where possible, to check functionality of controls and equipment. Any discovered deficiencies or leaks will then be repaired. Once all process deficiencies are corrected, systems will be turned over to the production personnel to start normal operation.

24.2.5 Water Management

Detailed engineering for water management will be initiated in the fall of 2013. Actual earthworks, dykes and trenches, required for the start-up of operations, will be carried out between both the summer and fall of 2014 and 2015.

24.3 Valleyfield Construction

This part of the Project will follow the same schedule basis. Delivery periods have been requested from suppliers for the major equipment. The long lead equipment area as follows: • Electrodialysis cells delivery 26 to 65 weeks After Reception of Order (“ARO”); • Kiln and cooler delivery 52 weeks ARO; • Crystallizer delivery 48 weeks ARO. The construction and commissioning are aligned with the completion of the Wabouchi Plant since the concentrate from this Plant is required for start-up.

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25.0 INTERPRETATION AND CONCLUSIONS

25.1 Conclusion

The Whabouchi Lithium Deposit and Hydromet Plant Project consists in the development of an open mine 300 Km North of Chibougamau and a lithium compounds production complex to be built in Valleyfield.

A computed cash flow analysis was developed by Met-Chem from the technical aspects and on metal prices projections made for lithium hydroxide and carbonate from two (2) market studies.

As it stands, the Whabouchi deposit contains an economic Mineral Resource.

Consequently, Met-Chem and BBA conclude that Whabouchi Lithium Deposit and Hydromet Plant Project is technically feasible as well as economically viable. The authors of this Technical Report consider the Whabouchi Project to be sufficiently robust to warrant moving it to the feasibility.

25.2 Risks Evaluation

Several aspects of the Whabouchi Lithium Deposit and Hydrometallurgical Plant Project are fairly well defined, however some aspects and assumptions need to be better defined and confirmed during the feasibility stage. Met-Chem and BBA made several recommendations to be followed up to for the next phase in Section 26.

The most significant risks in the Project identified are in technology, markets and environmental as explained in the following paragraphs.

The size and grade of the Whabouchi deposit are well defined. The process to produce spodumene concentrate is also well known and should not be problematic. However the processes of making lithium hydroxide monohydrate from spodumene concentrate by

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electro dialyse and lithium carbonate from lithium hydroxide in solution, have been developed by Nemaska with the assistance of SGS Minerals. These processes, for which Nemaska has filed 2 patent applications, have not been used on a commercial basis and therefore there are risks that the production and capital costs related to this portion of the processes differ from the ones used in the economic analysis which could have a direct impact on the economics of the Project.

Lithium is considered as an industrial mineral and the sales prices for the different lithium compounds are not public. Sales agreements are negotiated on an individual and private basis with each different end-user. Therefore it is possible that the sales prices used in the financial analysis be different than the actual market when Nemaska is in fact in a position to sell lithium compounds. In addition, there are a limited number of producers of lithium compounds and it is possible that these existing producers try to prevent new comers in the chain of supply by increasing their production capacity and lowering sales prices. In such case the economics of the Project could be affected.

One important uncertainty of the Project is the delay to obtain the required authorization (Certificat d’autorisation) to start construction of the mine and concentrator at Whabouchi, as well as the underlying conditions to the Authorization certificate from MDDEP. Consequently it is possible that the capital and operating costs forecasted in the financial model be different and increased in such a way as to affect the economic of the Project.

The delays to obtain construction permits, therfeore beginning of production of concentrate at Whabouchi, may allow other project to begin production before Nemaska and therefore reduce the portion of the market that Nemaska is targeting, which would impact sales level and economic of the Project.

Nemaska intends to produce mainly lithium hydroxide monohydrate to address the increasing demand for that compound due to use in the making of cathodes for rechargeable batteries. If cathode manufacturers use less hydroxide than expected or if the demand for rechargeable batteries, mainly in the electric and hybrid vehicles, is less than forecasted, it could have an effect on the sales price of that compound and the need for new production.

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26.0 RECOMMENDATIONS

This Preliminary Economic Assessement showed that Whabouchi Lithium Deposit and Hydromet Plant Project is technically feasible as well as economically viable.

The following recommendations should be considered to support the feasibility:

26.1 For Whabouchi Site • Perform a detailed investigation of the discharge water quality for both discharge points (reject pile and pit dewatering/run-off); • Continue kinetics testwork on waste rock and tailings samples; • Continue environmental investigation to comply with regulations; • Continue local communities discussions to better present the Project; • Continue discussions with Hydro-Quebec to firm-up power supply and costs to site; • Prepare to award long lead items contracts at the beginning of detailed engineering; • Continue discussions with CN Rail to firm-up a potential agreement for concentrate transportation; • Start recruitment of Owner’s Construction Management personnel early. Include sufficient float in schedule to minimize delays risks. 26.2 For Hydrometallurgical Plant • Prior to land acquisition, soil characterization survey should be completed; • Prior to land acquisition, geotechnical survey should be completed; • Complete metallurgical test works to better define size of process equipment; • Further define aluminium silicates and gypsum by-product characteristics in order make these products attractive to potential clients; • Continue environmental investigation to comply with regulations; • Continue local communities discussions to better present the Project; • Continue discussions with Hydro-Quebec to firm-up power supply and costs to site; • Prepare to award long lead items contracts at the beginning of detailed engineering; • Complete a trade of study of electrodialysis versus electrosynthesis.

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27.0 REFERENCES

Section 4 – Property Description and Location

Nemaska Exploration Inc., 2009: Initial Public Offering Prospectus document dated December 18, 2009.

Section 6 – History

Babineau, J. 2002: Spodumene Lake Project, Quebec, June 12-15, 2001, Rock Sampling and Assaying. Assessment Report NTS 32O/12. Inco Ltd., GM59815.

Beaupré, M.A. 2008: Examen de la propriété et échantillonnage, visite de terrain, propriété du lac Levac situé sur le territoire de la Baie James. Golden Goose Resources, GM 63939.

Bertrand, C., 1978: Rapport sur une pegmatite à spodumène, lac des Montagnes. Projet 402-1378-31. SDBJ, GM38134.

Brunelle, S. 1987: Report on Geophysical Surveys, Lac des Montagnes Property, Quebec. Muscocho Explorations Ltd., GM44641.

Burns, J.G. 1973: Summary Report, Geological Reconnaissance, July-August 1973. James Bay Nickel Ventures. Canex Placer ltd. GM 34021.

Charbonneau, R. 1982: Relevés géophysiques, électromagnétiques et magnétiques au sol, secteur de la bande sédimentaire de Nemiscau, comté d’Ungava, province de Québec. S.D.B.J. Programme Lac des Montagnes. GM 39991.

Elgring, F.H., 1962-63: Diamond Drilling, Lithium Occurrence. Township 1917, Quebec. Canico GM 57880.

Fortin, R. 1981: Rapport final, levé géophysique aéroporté, régions de Elmer Eastmain, Lac des Montagnes, Lac du Glas. Projet S80-5117 par Questor Surveys Ltd et Les Relevés Géophysiques inc., S.D.B.J. GM 38445.

Gilliatt, S. 1987: Report on VLF-EM Survey, Over the Lac des Montagnes Claim Group. Muscocho Explorations Ltd., GM 46065.

Gleeson, C.F. 1976: 126 plans d’un levé géochimique (sediments de lac), région du lac Bereziuk, rivière Eastmain et rivière Rupert. SDBJ. GM 34047.

Goyer, M., Picard, M., Lavoie, L., Larose, P.Y, 1978: Projet vérification d’anomalies géochimiques. Permis SDBJ 3. SDBJ. GM 34175.

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McConnell, T.J. 1987: Dighem III Survey, for Westmin Resources Ltd., Nemiscau Project Quebec. Dighem Surveys and Processing inc., GM 45242.

MRNF 1998: Résultats d’analyse de sédiments de fond de lacs, grand nord du Québec. MRNF, DP 98-01.

Otis, M., 1980: Projet Lien (402-1379-31). S.D.B.J. GM 37998.

Pride, C. 1974: Lake Sediment Geochemistry. SDBJ. GM 34044. Gleeson, C.F. 1975: Geochemical Report on a Lake Sediment Survey, Bereziuk Lake, Eastmain River and Rupert River Areas. GM 34046.

Théberge, D. 2009: NI 43-101 Qualifying Report, Whabouchi Property, James Bay area, NTS sheet 32O/12. Prepared by Solumines for Nemaska Exploration Inc.

Valiquette, G. 1964: Preliminary Report, Geology of Lemare Lake Area, Mistassini Lake Territory. Department of Natural Resources, Quebec, RP 518.

Valiquette, G. 1965: Preliminary Report, Geology of Cramoisy Lake Area. Mistassini Territory. Department of Natural Resources, Quebec, RP 534.

Valiquette, G. 1975: Région de la rivière Nemiscau. Ministère des Richesses Naturelles du Québec RG158

Zuiderveen, J. 1988: Diamond Drill Record, Lac des Montagnes Property. Muscocho Explorations Ltd. GM 47429.

Section 7 – Geological Setting and Mineralization

Card, K.D., and Ciesielski, A. 1986: DNAG #1. Subdivisions of the Superior Province of the Canadian Shield. Geoscience Canada.

Hocq, M. 1994: La Province du Supérieur; in Geologie du Québec. (ed.) M. Hocq, P. Verpaelst, T. Clark, D. Lamothe, D. Brisebois, J. Brun, G. Martineau. Les publications du Québec, p. 7-20. MM94-01.

Valiquette. L. A., Legault, M., Boily, M., Doyon, J., Sawyer, E., Davis, D.W. 2002: Synthèse géologique et métallogénique de la ceinture de roches vertes de la moyenne et de la basse Eastmain (Baie James). Ministère des Ressources Naturelles du Québec., ET 2002-06, ET 2007-01.

Valiquette, G. 1975: Région de la rivière Nemiscau. Ministère des Richesses Naturelles du Québec RG158

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Section 13 – Mineral Processing and Metallurgical Testing

Ameridia, June 8, 2012. Production of Lithium Hydroxide from Lithium Sulfate with Three-Compartment Bipolar Membrane Electrodialysis. 9 pages.

Feeco International, December 2, 2011. Spodumene Roasting and Acid Roasting Kiln Pilot Plant Test Report. 22 pages.

SGS Lakefield Research Limited, November 14, 2011: An Investigation into DMS Plant Testing on Material from the Whabouchi Lithium Deposit, Project 12486-004 Final Report for Nemaska Exploration Inc. 84 pages.

SGS Lakefield Research Limited, September 16, 2011, Project Update Report #2 – 12486-004 – Nemaska DMS for Nemaska Exploration Inc.

SGS Lakefield Research Limited, September 30, 2011, Project Update Report #3 – 12486-004 – Nemaska DMS for Nemaska Exploration Inc.

SGS Lakefield Research Limited, September 9, 2011: Project Update Report #1 – 12486- 004 – Nemaska DMS for Nemaska Exploration Inc.

SGS Lakefield Research Limited, July 14, 2010: An Investigation by High Definition Mineralogy into the Mineralogical Characteristics of Six Composite Samples from the Whabouchi Pegmatite Deposit. Project 12440-001, MI5011-MAY Final Report for Nemaska Exploration Inc. 78 pages.

SGS Lakefield Research Limited, October 20, 2011: An Investigation into the Recovery of Spodumene from the Whabouchi Property. Project 12486-001 – Final Report for Nemaska Exploration Inc. 171 pages.

SGS Lakefield Research Limited, November 14, 2011: An Investigation into the Grindability Characteristics of Samples from the Whabouchi Lithium Deposit. Project 12486-003 – Grindability Report for Nemaska Exploration Inc. 171 pages. 124 pages.

SGS Lakefield Research Limited, January 9, 2012: An Investigation into the Mineralogical Characterization of Two Composite Samples From the Shimanovskogo Project. Project 13219-001 – Mineralogy Progress Report 1 for BBA. 87 pages.

SGS Lakefield Research Limited, April 2, 2012: A Pilot Plant Investigation into the Recovery of Spodumene from the Whabouchi Property. Project 12486-003 – Final Report for Nemaska Lithium Inc. 799 pages.

SGS Lakefield Reseach Limited, May 8, 2012. The Piloting of the Whabouchi Property Hydrometallurgical Flowsheet. Project 13339-001 – Phase 1 Final Report. 51 pages.

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SGS Lakefield Research Limited, May 17, 2012. The Piloting of the Whabouchi Property Hydrometallurgical Flowsheet: Phase 2: An investigation into the Membrane Electrolysis of Lithium Sulphate to Produce Lithium Hydroxide. Project 13339-001 – Phase 2 Final Report. 101 pages.

SGS Lakefield Research Limited, May 18, 2012. An Investigation into the Synthesis of Lithium Carbonate from Lithium Hydroxide. Project 13339-001 – Phase 3 Report. 27 pages.

SGS Lakefield Research Limited, May 1, 2012. An Investigation into the Liquid Solid Separation and Rheological Response of a PIR Discharge Sample from the Whabouchi Property Hydrometallurgical Piloting. Project 13339-001 – Final Report. 132 pages.

SGS Lakefield Research Limited, May 10, 2012. Preliminary Results of Lithium Hydroxide Crystallization from LiOH Solution. 2 pages.

Section 14 – Mineral Resource Estimate

Talison Lithium Ltd, May 11, 2011. Management’s Discussion & Analysis of the Financial Condition and Results of Operations of Talison Lithium Limited as at and for the Interim Period Ended March 31, 2011. 23 pages.

Section 19 – Market Study and Contracts

Roskmil Consulting Group, Nemaska Lithium Inc., Battery grade lithium hydroxide market study, 21st September 2012, page 280.

SignumBOX Inteligencia de Mercados, March 2012. Lithium Minerals Market, Final Report. 73 pages.

Section 20 – Environmental Studies, Permitting, and Social or Community Impact

Genivar, 2010, Propriété Whabouchi, Étude préliminaire de caractérisation environnementale de base, Qualité de l’eau et des sédiments, inventaire des poissons et des invertébrés benthiques, 222 pages.

MDDEP 2005, révisé 2012. Directive 019 sur l’industrie minière. 105 p. http://www.mddep.gouv.qc.ca/milieu_ind/directive019/directive019.pdf.

Section 23 – Adjacent Properties

Perilya, 2011. News Release dated May 31, 2011 Reporting RPA JORC Mineral Resources. InnvExplo, 2011. Technical Report on the Pivert-Rose Property, January 24, 2011

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SRK, 2010. Mineral Resource Evaluation James Bay Lithium Project, James Bay, Quebec, Canada. December 30, 2010.

Théberge, 2011. NI 43-101 Technical Report Pertaining to the Lac Levac Property James Bay Area, Québec, June 1, 2011.

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Appendix A – Drawings

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F E A B D C H G AE REV. cm 10 A1 9 SIZE: 1 8 m 100 SHEET: 7 6 WAREHOUSE MINE OFFICE FUEL STATION MINE GARAGE LAKE DRAFTED BY: APPROVED BY: DATE: 2011-08-15 CRUSHING PLANT MAIN GATEHOUSE 5 DESCRIPTION SPODUMENE ADMINISTRATION OFFICE ADMINISTRATION 50 CONCENTRATE STORAGE ELECTRICAL SUB-STATION MINERAL PROCESSING PLANT POTABLE WATER PUMPHOUSE POTABLE 4 WHABOUCHI

286 25 3 INFRASTRUCTURES 0 2

GENERAL PLANT SITE PLAN

1:2500 288

2 288 292 1

2111 3121 2120 2910 3410 3430 3450 4100 5200 6000 1615 294

290

SUB-AREA

286

286 290 DESIGNED BY: VERIFIED BY: SCALE: 1:1 250 DRAWING No.: 3073002-001000-41-D20-0002 PROJECT: TITLE: 2000 2000 2000 3000 3000 3000 3000 4000 5000 6000 1000 AREA *???(1*?&,9,/?'5$:?Be0,66,21?)25,1)250$7,21?':*?'5$(':*

288 3 3

294 288

292 292

286

294 288 4 4

290 CLIENT:

286

290 288 294 ELEVATION 297.0 (NOM.) CONCENTRATOR PAD

288 FIRE PROTECTION PUMPING STATION (95.0 l/s CAPACITY)

SERVICE WATER TANK 292 5 5

294

STATION TO THICKENER CLEAR WATER PUMPING

288

296

WELL PUMPING STATION 294

5(48,5('&$3$&,7< PóK 5(48,5('&$3$&,7< 292

286

298 288 290 SEAL:

296

290 300

298 302

6 6

292

288 304

¡ ¡ 294

296 286

290

306

306 298

300 308

REVISIONS

302

306 304 7 7 2910 ¡

308 308

302

304 292

8 8 306

302 300 306 DESCRIPTION ¡ 9

9 308

304 306

306

REFERENCE DRAWINGS 302

304

306

292 300 ELEVATION 306.5 (NOM.) OFFICE PAD

¡ 298 10 10 294

¡ 302 DRAWING No.

296

300

/ / / / / /

/ / / 292 / / / / /

/ / / / / /

3121 304 CLEAR WATER LIFT STATION

296 306 11 11 298 AVAILABLE 85 PARKING SPACES 3073002-001000-41-D20-0002 TREATMENT BASIN TRIBUTARY AREA LIMITS

292

LEGEND 306 PILE/PIT LIMIT E: 442 000 NEW ROADS

294

35x35x2m

SEWAGE SEEPAGE FIELD 304 12

292 : 302 TOPOGRAPHIC DATA ARE FROM LASER SURVEY (LIDAR) PROVIDED THE CLAIM LIMITS REFERS TO GENIVAR REPORT, DECEMBER 2010 COORDINATE SYSTEM IS UTM NAD83, ZONE 18. SMALLER LAKE 1,2,3 AND STREAM A,B,C,D IDENTIFICATIONS WITHIN BY GROUP PHB AUGUST 10, 2011.

TREATMENT BASIN No1 300 2. 3. NOTES 1.

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N: 5 727 000 727 5 N: 298

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