opportunity study for the duncan lake iron ore project

GM 67283OPPORTUNITY STUDY FOR THE DUNCAN LAKE IRON ORE PROJECT

http://gq.mines.gouv.qc.ca/documents/examine/GM67283

http://gq.mines.gouv.qc.ca/documents/sigeom/Licence.pdf

http://gq.mines.gouv.qc.ca/documents/sigeom/license.pdf

Augyva Mining Resources Inc.

OPPORTUNITY STUDY FOR THE

DUNCAN LAKE IRON ORE PROJECT

JAMES BAY AREA QUEBEC — CANADA

FINAL REPORT

December 2010

GM 67283

REÇU AU N0ME

Ressources naturelles et Fe

22 MAI 2013

Dir information géologique

1 0 ANT 2013

DIRECTION DES TITRES MINIERS

Project Number: 28078 .

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Prepared by: RayIu d Jean Senior Geologist 06Q de-5-

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Augyva Mining Resouurces Inc. Opportunity Study for the Duncan Lake Iron Ore Project

Prepared for:

Augyva Mining Resources Inc. 154, des Cascades

La Macaza, Quebec, JOT 1R0

Prepared by:

Met-Chem Canada Inc. 555, boul. René-Lévesque Ouest, 3e étage

Montréal (Québec), H2Z 1B1

Date:

Charles Cauchon, Eng. Senior Process Specialist

Approved by:

Date: 02/7C"-) .)

D agnon, Eng. Manager, Mining Group

Date: ` e1//a/:)D/zs s Cloutier, Eng.

General Manager, Technology & Operations

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i D UR 2012 i .

PIRt`CTION DES TITRES MINERS

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Augyva Mining Resources Inc. Opportunity Study for the Duncan Lake Iron Ore Project

TABLE OF CONTENTS

1.0 EXECUTIVE SUMMARY 1.1

Introduction 1.2 Property Description and Location 1.3 2009 Exploration Work 1.4 Mineral Processing and Metallurgical Testing 1.5 Mineral Resource Estimate 1.6 Other Relevant Data and Information 1.7 Project Schedule 1.8 Capital Cost Estimates 1.9 Operating Costs per Tonne of Concentrate 1.10 Preliminary Economical Analysis 1.11 Work Program 1.12 Recommendations

2.0 INTRODUCTION AND TERMS OF REFERENCE 2.1 Introduction 2.2 Scope of Work 2.3 Sources of Information 2.4 Units and Currency 2.5 Met-Chem's Qualifications 2.6 Disclaimer

3.0 PROJECT INFORMATION 3.1 COREM Test Results 3.2 Mineral Resources Estimate 3.3 Economic Parameters to Simulate Mining Reserves 3.4 Simulation of Mineral Reserves 3.5 Production Rate 3.6 Mining Concept and Infrastructures Location

4.0 CRUSHER 5.0 CONCENTRATOR

5.1 Primary Grinding SAG 5.2 Alternative, Secondary Crushing and High Pressure Grinding Rolls 5.3 Magnetic Separation and Regrinding

6.0 SLURRY PIPELINE 7.0 FILTERING AND PELLETIZING PLANTS

7.1 Filtering 7.2 Mixing 7.3 Balling 7.4 Pellet plant 7.5 Feed System 7.6 Induration and Hearth Layer 7.7 Auxiliary Systems and Infrastructure for the Pellet Plant and Port

8.0 PORT FACILITIES, STOCKPILING, RECLAIMING AND SHIP LOADING 8.1 Concept 8.2 Pellet stockpiling and Reclaiming 8.3 Jetty and Ship Loading Berth

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8.4 Metallurgical Tests 34 9.0 MANPOWER 35 10.0 COMPARABLE IRON ORE PROJECTS IN QUEBEC 36

10.1 LabMag Iron Ore Project 36 10.2 Bloom Lake Project 37 10.3 Baffinland Iron Mines - Mary River Direct Shipping Ore Project 39 10.4 KéMag Iron Ore Project 41

11.0 EXPECTED PROJECT SCHEDULE 44 12.0 CAPITAL COST ESTIMATES 46

12.1 Scope 46 12.2 Cost Estimate 46 12.3 Exclusions 47

13.0 OPERATING COST ESTIMATES 50 13.1 Summary of Estimated Operating Costs 50

14.0 PRELIMINARY FINANCIAL ANALYSIS 51 14.1 Constant Parameters 51 14.2 Variable Parameters 52 14.3 Other Assumptions 52 14.4 Results of the Analysis 53

15.0 WORK PROGRAM 54 15.1 Drilling Proposal of Deposits 3, 4, and 6 54 15.2 Scoping Study 55 15.3 Feasibility Study 58

16.0 CONCLUSION AND RECOMMENDATIONS 61 17.0 REFERENCES 62

LIST OF TABLES

Table 1.1 - Measured and Indicated Mineral Resources by Deposit 3 Table 1.2 - Inferred Mineral Resources by Deposit 3 Table 1.3 - Simulation of Potential Added Tonnage by Deposit 4 Table 1.4 - Capital Cost Estimates in M $ CAN 5 Table 1.5 - Summary of Estimated Operating Costs 5 Table 1.6 - Summary of the Results of the Preliminary Analysis 6 Table 2.1 - List of Abbreviations 10 Table 3.1 - Mean of Chemical Analysis of Product by Deposit 15 Table 3.2 - Measured and Indicated Mineral Resources by Deposit 16 Table 3.3 - Inferred Mineral Resources by Deposit 16 Table 3.4 - Simulation of Potential Added Tonnage by Deposit 17 Table 3.5 - Estimated of Production Cost Figures 17 Table 3.6 - Iron Ore Sale Price 18 Table 3.7 - Pit Wall Angle 18 Table 3.8- Rock Density 18 Table 3.9 - Simulation of Mining Reserves - All Deposits 19 Table 3.10 - Simulation of Mining Reserves Including Potential Tonnage All Deposits 19 Table 3.11 - Grade and Tonnage for low Silica Deposits 20 Table 3.12 - Simulation of Mining Reserves - All Deposits Stripping Ratio Forced 21 Table 10.1 - Howells River Estimated Mineral Resources (2006) 36 Table 10.2 - Howells River Estimated Capital Cost (2006) 37 Table 10.3 - Howells River Estimated Operating Costs (2006) 37 Table 10.4 - Mineral Reserves in Ultimate Pit Design (2006) 38 Table 10.5 - Bloom Lake Estimated Operating Costs (in US$ - 2008) 39

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Table 10.6 — Deposit Nos.1, 2 and 3 Mineral Resources Exclusive of Reserves and Mineral Reserves (Deposit No.1) 39

Table 10.7 — Mary River Project Estimated Capital Cost (2008) 40 Table 10.8 — Mary River Project Estimated Operating Costs (2006) 41 Table 10.9 — Proven and Probable Reserves (Cut-off grade 18% DTWR) 42 Table 10.10 — Summary Estimated Capital Cost of KéMag Project (2009) 42 Table 10.11 — KéMag Project Estimated Operating Costs (2009) 43 Table 12.1 — Capital Cost Estimates in M$ CAN 47 Table 12.2 — Capital Cost Estimates in M$ CAN 48 Table 12.3 — Capital Cost Estimates in M $ CAN 49 Table 13.1 — Summary of Estimated Operating Costs 50 Table 14.1 — Preliminary Data Used in the Analysis 51 Table 14.2 — Other Constant Parameters Used in the Analysis 51 Table 14.3 — Estimated Unit Operating Costs used in the Analysis 52 Table 14.4 — Variable Parameters used in the Analysis 52 Table 14.5 — Summary of the Results of the Preliminary Analysis 53 Table 15.1 — Boreholes Proposal to determine Inferred Resources 54 Table 15.2 — Boreholes Proposal to determine Measured and Indicated Resources 55 Table 15.3 — Work Program Cost Estimate for Scoping Study Level 56 Table 15.4 — Work Program Cost Estimate for Feasibility Study Level 59

LIST OF FIGURES

Figure 1.1 — Project Schedule 4 Figure 2.1 — Duncan Lake Iron Ore Project Location, Québec 11 Figure 2.2 — Regional Location of Duncan Lake Iron Ore Project 12 Figure 3.1 — Economical Pit Envelops of Deposits 1 to 6 20 Figure 3.2 — Bathymetric Map 22 Figure 3.3 — Infrastructures Location Map 23 Figure 3.4 — Port Area Infrastructures 24 Figure 11.1 —Project Schedule 45

LIST OF APPENDICES

Appendix A — Conceptual Flow Sheet and Simplified Block Diagram Appendix B — Details of the Preliminary Financial Analysis Appendix C — Planned Drill Locations

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Augyva Mining Resources Inc. Section 1 Opportunity Study Report for the Duncan Lake Iron Ore Project

Page 1

1.0 EXECUTIVE SUMMARY

1.1 Introduction

Following the presentation of the report entitled "Technical Report 43-101 on Mineral Resources of the Duncan Lake Iron Ore Project", Augyva Mining Resources Inc. (Augyva) requested Met-Chem Canada Inc. (Met-Chem) to prepare an Opportunity Study to get a better appreciation of the entire Project.

In November 2010, Canadian Century Iron Ore Corporation earned a 51% interest in Augyva' s Duncan Lake Property. Augyva holds the remaining 49% of the Property. Canadian Century Iron Ore Corporation was acquired by Red Rock in an all stock deal. To complete the transaction, a newly-incorporated subsidiary of Red Rock will amalgamate with Century Iron Ore Holdings (100% owned by Canadian Century Iron Ore Corporation).

The Study would determine operating production costs as well as a long term sale price to simulate mining reserves, tonnage, grade, and weight recovery factor (WR) using LG module in MineSight software to sustain a long term mining project. Thereafter the Study would evaluate, through comparison with projects disclosed on the SEDAR website and other sources, the expected equipment requirement and cost of a complete mining project including the mine, crusher and concentrator, slurry pipeline, filtering, balling and pelletizing plant, stockpiling and reclaiming, port, ship loading and all support facilities at the mine and port site. The study would include potential financial projections for various production scenarios to assist Augyva in the review of further exploration decisions. A work program and budget estimate would also be included.

1.2 Property Description and Location

The Duncan Lake property is located approximately 570 km north of Matagami, Québec, which is the nearest mining community. Access is by a paved road (Highway 109), connecting Mattagami to Radisson. Road distance from Montreal to Duncan Lake property is about 1,350 km. The property is located 50 km south of Radisson and 10 km south of the LG2 regional airport.

The property is divided into six (6) different blocks (Deposits 1 to 6), located close to Highway 109. Deposits 1, 2, and 5 are located southwest of the highway and Deposits 3, 4 and 6 to the northeast in a lateral extension of the other three deposits.

Radisson (population 500) is the closest municipality. The population may reach 700 with the personnel hired for temporary work by Hydro-Québec all around the sector. Radisson was created by Hydro-Québec in 1974 to provide housing, and to support development of the La Grande hydro-electric complex. Actually, Radisson provides services for the local, permanent population and temporary employees of Hydro-Québec.

The local topography of Duncan Lake is characterized by low plain, rolling terrain, elongated in the northeast direction, with several small hills and low lying swampy areas.

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In general, the terrain is quite swampy and flat, except for a few low rocky hills and unconsolidated glacial eskers and moraines composed of silt, sand, gravel and boulders. The presence of the iron formation band does not influence the topography too much. Deposits 1, 3, 4, and 5 are characterized by low hills and elongated topographic features of less than 10 meters high.

1.3 2009 Exploration Work

Exploration work consisted mainly of ground magnetic surveys on the six deposits and diamond drilling on five of the deposits. From the total of 10,460 m that was drilled, a total of 6,484 m of core have been cut in half and sampled for 1,489 samples collected and analysed for major elements by Chemex Laboratory in Val d'Or. In addition, 218 samples were sent to COREM in Québec City, for metallurgical testing. Analytical tags, subdivided in three parts, were inserted in the plastic bag, inserted and stapled to the bottom of the core interval in the core boxes, and the remaining tag was kept in the assay booklets and entered into the computer by the logging geologist or technician.

1.4 Mineral Processing and Metallurgical Testing

To characterize an iron ore resource for a deposit that needs concentration, not only head assays are required (like iron content, silica and others elements), but also it is important to know the percentage of material which will be concentrated (weight recovery = WR) along with its chemical composition and quality.

Besides mineralogical studies, COREM has performed metallurgical tests such as Davis Tube (DT) and Satmagan analysis of 144 samples from Deposits 1 to 4. All samples were ground to 85% passing -75 microns, prior to testing.

Differences between the deposits have been observed. Deposits 3 and 4 gave better quality concentrate and lower % silica and other chemical elements. The deposits tested 68.93% FeT and 2.58% Si02 and 68.38% FeT and 4.61% Si02 respectively. Deposit 4 compared with Deposit 1 gave a higher WR and a better quality product with identical head grade. Deposits 1 and 2 will probably need a finer grain size to reach better liberation of the silica than the present 65.83% FeT and 7.60% Si02, and 66.11% FeT and 7.11% Si02 respectively. All other chemical elements such as phosphorous, manganese, titanium etc. are reduced to acceptable level in the concentration process. However, the average of the COREM's samples was slightly higher than the average grade of the deposits, based on the mineral resources estimation.

1.5 Mineral Resource Estimate

At the selected cut-off grade of 16% FeT, the Measured and Indicated Mineral Resources are shown, by Deposit number, in Table 1.1. Tonnage of the Measured and Indicated in those categories is low because of the low drilling density and wide drilling spacing. Following NI 43-101 standards the Inferred Mineral Resources, shown in Table 1.2, cannot be added to the previous mineral resources. Inferred resources contain most of the total tonnage in Deposits 1 and 4 (predominantly drilled), and Deposit 3. Deposits 2 and

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5 are very limited for potential tonnage increase; however, Deposit 2 has a better grade than other deposits. Deposit 3 has relatively low drilling density; however, the thickness of the ore lenses appears large, which was also confirmed by the magnetic survey. Deposit 3 has a good potential to increase the tonnage volume because of the anticipated favourable syncline structure.

Table 1.1 — Measured and Indicated Mineral Resources by Deposit

Cut-off: 16% FeTotal

Measured Mineral Resources Tonnes* FeTotal(%) Deposit 1 4,090,000 21.89 Deposit 2 902,000 27.41 Deposit 3 0 Deposit4 708,000 26.19 Deposit 5 0 Deposit 6 0

Total 5,700,000 23.29

Indicated Mineral Resources Tonnes* Ferotai(%)

Deposit 1 13,443,000 21.80

Deposit 2 4,931,000 27.44 Deposit 3 0 Deposit4 7,241,000 25.20 Deposit 5 0 Deposit 6 0

Total 25,615,000 23.84

Total (Measured + Indicated) Tonnes* Ferotai (%)

Deposit 1 17,533,000 21.82

Deposit 2 5,833,000 27.44 Deposit 3 0 Deposit 4 7,949,000 25.29 Deposit 5 0 Deposit 6 0

Total 31,315,000 23.74

*Tonnes are rounded

Table 1.2 — Inferred Mineral Resources by Deposit

Infered Mineral Resources Tonnes* Ferotai(%)

Deposit 1 283,857,000 23.71

Deposit 2 78,795,000 26.90 Deposit 3 154,724,000 24.83 Deposit 4 192,336,000 24.33 Deposit 5 32,658,000 24.48 Deposit 6 78,765,000 25.35

Total 821,135,000 24.56

*Tonnes arc rounded

1.6 Other Relevant Data and Information

Table 1.3 shows the results of a simulation technique to define additional potential tonnage. Deposits 1 and 3 have better chance to substantially increase potential tonnage. Major reasons for this are related to the very long undrilled extension of Deposit 1 and on the very partial and limited drilling information data for Deposit 3, both associated with

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high magnetic susceptibility anomalies. Additional potential tonnage is in the range of 400 to 500 million tonnes. Estimated grade range varies from 23 to 27 % iron, estimated to be the same range grade as the Inferred resources.

Table 1.3 — Simulation of Potential Added Tonnage by Deposit

Simulation by Deposit Range of Potential Tonnage*

(M of Tonnes)

MIN MAX

Deposit 1 100 120

Deposit2 20 30

Deposit 3 160 180

Deposit4 30 50

Deposit5 40 60

Deposit6 50 60

Total 400 500

* Tonnes are rounded by millions

Density = 3.2

1.7 Project Schedule

The EPCM part of the project is likely to require about 36 months; however, nothing can start on the site until the Environmental Assessments Process is completed which will require 24 months or more O.

Figure 1.1— Project Schedule

ID Task Name 2010 2011 2012 2013 2014 2015 2016 1121314 1121314 1121314 1121314 1121314 1121314 1121314

1 Exploration and Testing I 1

2 Pre-Feasibility Study

• 3 Decision

4 Feasbility Study I

5 Environmental Assessment Process

6 Pernik Application For Construction

7 Decision

8 EPCM

9 Power Transmission Lie Canstnic6m

10 Mine Development I I

11 Concentrator Construction I I

12 Concentrator Conrriasioring

13 Pipeline Construction I 1

0 14 Pipeline Commissioning

15 Palet Plant Construction I 7

0 16 Pelet Plant Commissioning

17 Stockyard and Shiploading Construction I 1 _

0 18 Stodryard and Shiploading Commissioning

19 Production Startup I

20 Full Production •

Task I MieatOrre ♦

1.8 Capital Cost Estimates

The capital cost estimates include all costs required for development, start-up and operation of a complete project starting with the mine and finishing with the shipping of

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the iron pellets produced. In order to determine the best potential economic scenario various production rates were analyzed namely 3, 6, 9, 12, 15, 18 and 21 Mtpy operations. The summarized estimated capital costs for the 7 scenarios are shown in Table 1.4.

Table 1.4 — Capital Cost Estimates in M $ CAN

Description 3 Mtpy 6 Mtpy 9 Mtpy 12 Mtpy 15 Mtpy 18 Mtpy 21 Mtpy

Mining Complex 475 $ 685 $ 874 $ 1 020 $ 1 141 $ 1 301 $ 1 427 $

Pipeline 273 $ 291 $ 328 $ 357 $ 364 $ 455 $ 499 $

Pellet Plant 271 $ 400 $ 505 $ 597 $ 680 $ 761 $ 835 $

Port 218 $ 232 $ 242 $ 249 $ 250 $ 318 $ 348 $

Total Direct Estimated Costs

1 237 $ 1 608 $ 1 959 $ 2 223 $ 2 435 $ 2 835 $ 3 110 $

Total Indirect Estimated Costs

298 $ 388 $ 469 $ 536 $ 586 $ 683 $ 749 $

Total Estimated Costs

1 535 $ 1 996 $ 2 418 $ 2 759 $ 3 019 $ 3 518 $ 3 859 $

1.9 Operating Costs per Tonne of Concentrate

The operating cost is estimated at CAD$ 32.60 per tonne of concentrate or CAD$ 40.75 per tonne of pellets. Mining cost appears to be high and this is due to the relatively low grade and high stripping ratio.

Table 1.5 — Summary of Estimated Operating Costs

Description $/tonne $/tonne concentrate

$/tonne pellets

Mining of ore 2.15 7.42 7.42

Mining of waste 1.60 11.54 11.54

Concentrator 10.56 10.56

Pipeline 1.09 1.09

Pellet Plant - 8.97

Stocking/Loading 0.98 0.98

Administration 0.19 0.19

Filtration Plant 0.82 -

Operating Cost Total 32.60 40.75 Note that the Operating Costs are equal except for the Pellet Plant and Filtration. This is because 1 t of concentrate = 1 t of pellets. This will have to be adjusted after Pot Grate test work.

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1.10 Preliminary Economical Analysis

It must be realized that the described preliminary economical analysis is based on data that is not yet compliant with NI 43-101 parameters and that the resources that are used in the analysis are not compliant with NI 43-101 requirements. The results of the analysis are therefore for guidance of the next steps of exploration only.

The parameters used for the preliminary analysis are shown in Section 13.0 and are summarized above. The details of the analysis are summarized in Table 1-6.

Table 1.6 — Summary of the Results of the Preliminary Analysis

Description 12 Mtpy 15 Mtpy 18 Mtpy 21 Mtpy NPV (Rate of 10%) $686.1 M $866.9 M $764.6 M $732.3 M

IRR 13.5% 15.0% 14.3% 14.1%

Payback Year 6 Year 5 Year 5 Year 5

Life of Project* 20.3 years 16.9 years 14.6 years 12.9 years

Ore Volume Required (25 years) 1,039 Mt 1,298 Mt 1,558 Mt 1,818 Mt * Includes 3 year construction period.

1.11 Work Program

A preliminary work program has been prepared describing the exploration phases and the work to be performed on the property in order to bring the project to a Feasibility Study level with the objective to achieve a fast track mine development project. It includes a proposal for the drilling of Deposits 3, 4 and 6 to increase mineral resources of the Inferred category for a S coping Study, and of the Indicated and Measured category for the Final Feasibility Study. It is recommended to start on the eastern part of the property. The work program is divided into two (2) phases, passing directly from the scoping level to the feasibility study level to minimize delays that are generally present between phases for decision making.

The first phase of drilling would consist of some 83 holes for a total of 26,800 m. During this phase, environmental work must be carried out as well as geological mapping. Samples of the drilling must be assayed and tested to obtain additional metallurgical information. At the end of this phase, estimated to take two years, a Scoping study report will be prepared. Total estimated cost for the first phase is approximately $5.1 million.

The second phase, which is recommended to continue at the same time that the Scoping study is being prepared, will consist of additional drilling of some 89 holes for a total of 28,700 m. Additional studies have to be carried out such as mining, metallurgical and pelletizing studies. Bulk sample test work is to be done to obtain data for the concentrator and pellet plant design required for the Feasibility study. At the end of this phase, also estimated to take two years, a Feasibility study must be prepared. Total estimated cost for this second phase is approximately $8.7 million.


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1.12 Recommendations

The project is a challenge, but it is feasible. The Duncan Lake Iron Ore Project (DLIOP)'s deposits were relatively well outlined by drilling and geophysical surveys completed to date. Additional drilling will be required to convert Inferred Resources into Indicated or Measured categories. In addition, and in consideration of the high mining cost estimated in the Opportunity Study, additional tonnage at lower stripping ratio would be a critical key point.

Considering the critical timing of this large project, it is urgent to:

• Initiate a drilling campaign to prove considerably more reserves, preferably in Deposits 3, 4 and 6;

• Initiate the Environmental Assessment studies. This item is critical and ultimately controls when and if work can start at the site;

• Select representative core samples for laboratory test work and a representative, bulk sample for pilot plant test work to confirm the flow sheet from laboratory test work and allow more precise equipment sizing;

• Initiate bathymetry studies of the proposed port area in James Bay; and

• Initiate a Pre-Feasibility followed by a Feasibility Study to further define the main parameters of the project. These studies would also establish a flow sheet, mine study, mechanical/electrical equipment list, preliminary layout for the mine, concentrator and pellet plant and port sites, develop arrangement drawings for each plant and establish all quantities for civil and steel of all types, develop CAPEX and OPEX, and prepare a financial analysis to allow the Owners to make an informed decision on the project.

The Feasibility Study should include budget prices for all the major equipment and also quotations and prices for:

• Slurry pipeline;

• Pellet Plant; and

• Port and ship loading facilities.


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2.0 INTRODUCTION AND TERMS OF REFERENCE

2.1 Introduction

Following the submission in January 2010 of the Technical Report 43-101 on Mineral Resources of the Duncan Lake Iron Ore Project, Augyva Mining Resources Inc. (Augyva) requested from Met-Chem Canada Inc, (Met-Chem) the preparation of an Opportunity Study to have a better appreciation of the entire Project. The Study would contain an estimate of capital and operating production cost as well as a long term price projection to estimate mining resources, tonnage, grade, and weight recovery factor (WR) using LG module in MineSight software to simulate a long term mining project. Thereafter the Study would evaluate, through comparison with projects disclosed on the SEDAR website and other sources, the expected equipment requirement and cost of a complete mining project including the mine, crusher and concentrator, slurry pipeline, filtering, balling and pelletizing plant, stockpiling and reclaiming, port, ship loading and all support facilities at the mine and port site.

The study would also include a projection of potential financial projections for various production scenarios to assist Augyva in the review of further exploration decisions. A work program and budget estimate would also be included.

Augyva is a junior mining company with several exploration properties in Northern Québec. Some of these properties are related to low grade, magnetite iron ore type. The Company does not have any property in production. The Company's head-office is located at La Macaza, Québec.

In November 2010, Canadian Century Iron Ore Corporation earned a 51% interest in Augyva's Duncan Lake Property, in accordance with an option and joint venture agreement entered into between the parties in May, 2008. Augyva holds the remaining 49% of the Property. Century Iron Ore Corporation may increase its interest in the Duncan Lake property to 65%.

Canadian Century Iron Ore Corporation was acquired by Red Rock in an all stock deal. To complete the transaction, a newly-incorporated subsidiary of Red Rock will amalgamate with Century Iron Ore Holdings (100% owned by Canadian Century Iron Ore Corporation).

The Duncan Lake property is located approximately 570 km north of Matagami, Québec, which is the nearest mining community. Access is by a paved road (Highway 109), connecting Mattagami to Radisson. Road distance from Montreal to Duncan Lake property is about 1,350 km. The property is located 50 km south of Radisson and 10 km south of the LG2 regional airport (see and ).

The Duncan Lake Project is composed of six (6) different blocks or properties named Deposits 1 to 6. Drilling program 2008-2009 was realized on five (5) deposits, mainly on Deposit 1, with few testing boreholes on Deposits 2, 3, and 5, and tentative confirmation

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boreholes (twin holes) on Deposits 4 which was previously drilled in 1973. Deposit 6 previously drilled in 1973 was not drilled in 2009. The 2008-2009 drilling program consisting of fifty-two (52) diamond drill holes (10,460 meters) was completed in April 2009. Its objectives were to validate Deposit 4, 1973 historical boreholes data, convert the historical mineral resources into 43-101 compliant resources and to define additional tonnage with definition resource on Deposit 1. The 2008-2009 drilling program was under the supervision of independent geological consultants assisted by Augyva's employees.

In September 2009, Met-Chem was mandated to estimate the resources of the Duncan Lake Project based on the information of the latest drilling campaign as well as on a new field magnetometer survey. Mineral resource estimate was performed in accordance with "National Instrument 43-101 Standards of Disclosure for Mineral Projects" and the "CIM Definitions Standards on Mineral Resources and Mineral Reserves" adopted by CIM Council (2004).

2.2 Scope of Work

The scope of services provided by Met-Chem consisted of:

• Estimation of mineral resources to support a long term mining project;

• Elaboration of a typical flow sheet for a mine and a magnetite concentration plant complete with slurry pipeline, pelletizing and port/shipping facilities; and

• Development of the potential OPEX and CAPEX of the project based on published and in house information, for various production scenarios.

2.3 Sources of Information

The documentation used in the preparation of this Opportunity Study was supplied by Augyva, retrieved from publicly available reports on the SEDAR website on similar projects, or in-house information.

Metallurgical information and tests results are from tests at the COREM laboratory, Québec and are based on core samples selected during the 2008-2009 drilling campaign.

2.4 Units and Currency

Metric units are used throughout the text, figures and drawings of this report. Table 2.1 shows the abbreviations list used in this Report. All currency amounts are in Canadian Dollars ($CAD) unless otherwise stated.


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Table 2.1— List of Abbreviations

Abbreviation Description DDT Davis Tube Test g gram g/t gram/tonne m metre kg kilogram km Kilometre km2 square kilometre Mt million metric tonnes ppm, ppb part per million, part per billion tonnes or t metric tonne %WR % weight recovery dmtu % Fe units per dry metric tonne


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Figure 2.1 — Duncan Lake Iron Ore Project Location, Québec


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a mET cHEm P:\28078\Texte\Rapports\Opportunity Study Report\Final Report\28078 Opportunity Study Final Report-Rev.4.doc



Page 13

2.5 Met-Chem's Qualifications

Met-Chem Canada Inc. is based in Montreal and was incorporated in October 1969 as a Canadian subsidiary of UEC Technologies LLC a subsidiary of US Steel.

Since then, Met-Chem's mandate has been to offer a complete range of consulting and engineering services to the mining industry, including exploration, geology, resources evaluation, mine design, process design, feasibilities studies, detailed engineering, etc. Met-Chem has a very long history on the study of iron ore, related to low grade magnetite and coarse hematite ore type in Canada, USA, India and other countries.

2.6 Disclaimer

The present report was originally written for Augyva by Met-Chem, and is based on information available to Met-Chem at the time of preparation of the report in the months prior to July 2010.

While the services performed by Met-Chem to create this report were in accordance with good engineering principles as practiced by competent engineering organizations, the analysis, information, opinions and recommendations provided by Met-Chem in this report are advisory only, and the owners of the Duncan Lake Project shall retain control over its operations. The present report is directed solely to the development and presentation of data and evaluations to permit the owners of the Duncan Lake Project to reach informed decisions on the project.

It should further be understood that the Opportunity Study presented in this report is based on a comparison with other large iron projects, and not on actual test work on ore from this project or any third party consultation for specific areas of the project, such as slurry pipeline, pelletizing or port facilities. The estimates of the size and grade of the deposits are based on limited information from drilling performed on a wide spacing and recent field magnetic survey calibrated on the magnetic susceptibility of the core analyzed and sampling and the assumptions and the parameters currently available.

The level of confidence in the Opportunities Study depends upon a number of uncertainties. These uncertainties include, but are not limited to, ore resources tonnages and qualities, beneficiation characteristics including liberation size, concentrate grade, impurities content, and recovery rates from those projected, as well as changes in project parameters and environmental considerations. These may have a serious impact on the reserves and on the size and life of the project and ultimately on its economic viability

There was no environmental review carried out under Met-Chem, and the limited relevant information already available is described in the report.

Metallurgical concentration tests consist of Davis Tube only (DT) performed by COREM laboratory in Quebec City. The metallurgical tests and the process control were not verified by Met-Chem; however, results obtained are in accordance with geology and


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mineralogy observed at Duncan Lake, and is comparable with other similar iron ore deposit.

Met-Chem's estimates are based on methodologies and procedures consistent with accuracy levels as stated in the Report. This report, or portions of it, is not to be reproduced or used other than for internal information by the owners of the Duncan Lake Project; any other use of this report by any third party is at that party's sole risk.

This Opportunity Study is not intended to be used by Augyva, Century Iron Holdings Inc, or Red Rock Capital Corp as a Technical Report with Canadian Securities regulatory Authorities pursuant to provincial securities legislation. Except for the purposes contemplated under provincial securities laws, any other use of this Technical Audit Report by any third party is at that party's sole risk.


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3.0 PROJECT INFORMATION

3.1 COREM Test Results

Based on Davis Tube tests done by COREM in Quebec City on 144 samples, (see Table 3.1 below), Deposits 3 and 4 produced, at 75 microns, the best concentrate grade; more than 68% Fe with silica at less than 5%. Deposits 1 and 2 produced about 66% Fe with silica of more than 7.0%; however, the average grade of the selected samples for testing is higher than the average iron grade of the mineral resources estimate for each deposit. The WR factor is closely associated to iron content.

The concentrate obtained from Deposits 3 and 4 is quite good for Blast Furnace (BF) pellet feed and could be made acceptable for Direct Reduction (DR) pellets by lowering the silica. Deposits 1 and 2 will need a finer grind to liberate the iron unit and possibly flotation to dispose of the silica. It is to be noted that the COREM results are somewhat higher than those received from a different laboratory. The reasons for the discrepancy have not been investigated at this point. The higher concentrate grade and lower silica from Deposits 3 and 4 are due in part to the slightly coarser grain size and lower silica content in the ore of these deposits. Deposit 6 is expected to be similar to Deposit 3 and Deposit 4.

Table 3.1- Mean of Chemical Analysis of Product by Deposit

Location Head Assay Concentrate Assays WR

Deposit Samples FeT FeT 5102 A1203 Fe203 MgO CaO Na20 K20 TiO2 MnO P205 Cr203 LOI

# Qty % % % % % % % % % % % % % % %

1 54 27.18 65.83 7.60 0.29 94.18 0.14 0.24 0.048 0.111 0.025 0.026 0.021 0.012 -2.87 34.01

2 57 31.01 66.11 7.11 0.27 94.58 0.24 0.11 0.038 0.122 0.035 0.015 0.024 0.014 -2.80 38.77

3 23 29.45 68.93 2.58 0.08 98.62 0.08 0.07 0.022 0.028 0.024 0.013 0.014 0.015 -2.46 31.08

4 10 27.11 68.38 4.61 0.12 97.83 0.07 0.08 0.027 0.053 0.020 0.017 0.019 0.014 -3.11 34.13

Total 144

In addition 42 tests on option 1

Based on the quality of the metallurgical test results obtained from Deposits 4 and 3, it was proposed to initiate the mining study of the deposits located on the eastern side of the property associated to coarser grain size.

3.2 Mineral Resources Estimate

At the selected cut-off grade of 16% FeT, the Measured and Indicated Mineral Resources are shown in Table 3.2, all by Deposit number. The Measured and Indicated tonnage is low due to a low drilling density and wide drill hole spacing.


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Table 3.2 — Measured and Indicated Mineral Resources by Deposit

Cut-off: 16% FeTotal

Measured Mineral Resources Tonnes* Ferotai(%) Deposit 1 4,090,000 21.89 Deposit 2 902,000 27.41 Deposit 3 0 Deposit 4 708,000 26.19 Deposit 5 0 Deposit 6 0

Total 5,700,000 23.29

Indicated Mineral Resources Tonnes* FeTotal (%)

Deposit 1 13,443,000 21.80


Total 25,615,000 23.84

Total (Measured + Indicated) Tonnes* Ferotai (%)

Deposit 1 17,533,000 21.82


Total 31,315,000 23.74

*Tonnes are rounded

Following the NI 43-101 standards, the Inferred Mineral Resources cannot be added to the Measured and Indicated Mineral Resources. The Inferred Resources contain all tonnage, mainly associated to Deposits 1 and 4 (mostly drilled), and Deposit 3. Deposits 2 and 5 are very limited for tonnage extension; however, Deposit 2 has a higher grade than other Deposits. Deposit 3 has a relatively low drilling density; however, the thickness of the ore lenses appears large that was also confirmed by the magnetic survey. Deposit 3 has good potential to increase the tonnage because of the anticipated favourable syncline structure.

Table 3.3 — Inferred Mineral Resources by Deposit

Infered Mineral Resources Tonnes* Fe rota' (%)

Deposit 1 283,857,000 23.71

Deposit 2 78,795,000 26.90 Deposit 3 154,724,000 24.83 Deposit 4 192,336,000 24.33 Deposit 5 32,658,000 24.48 Deposit 6 78,765,000 25.35

Total 821,135,000 24.56 *Tonnes are rounded

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The potential for additional tonnages has been identified and estimated in the Table 3.4; however, is not considered as mineral resources as per NI 43-101 standards.

Table 3.4 — Simulation of Potential Added Tonnage by Deposit

Simulation by Deposit Range of Potential Tonnage*

(M of Tonnes)

MIN MAX

Deposit 1 100 120

Deposit 2 20 30

Deposit 3 160 180

Deposit 4 30 50

Deposit 5 40 60

Deposit 6 50 60

Total 400 500

* Tonnes are rounded by millions

Density = 3.2

3.3 Economic Parameters to Simulate Mining Reserves

The determination of economic pit limit and reserves for each deposit were made using the EPIT module of MineSight mining software. The EPIT module uses the Lerch Grossman (LG) pit optimization method to obtain the optimum net value pit. A number of parameters are input into the software to calculate the results, including mining, processing and administration costs, product sales price, and maximum pit slope angles. The software then applies these parameters to the ore and waste blocks of the 3D block models to calculate the optimal net revenue pit.

Production costs are used by the optimizer to evaluate optimal pit configuration and are usually expressed as a cost per ton of ore (or waste) basis. Production figures were estimated by Met-Chem based on data analysis of similar project in public domain.

Table 3.5 — Estimated of Production Cost Figures

Mining &

Processing

Mining Cost Ore ($/t) 2.15

Mining Cost Waste ($/t) 1.60

Processing Cost ($/t conc.) 11.00

Fixed Administration Cost ($/t conc.) 3.10

For LG simulations and optimal pit limits, sales price per type of material produced and sold, is required. The selection of an accurate sales price which is not too optimistic or too pessimistic is very important. Iron ore prices vary depending on the quality of the final product and distances from the source to the user. In the fourth quarter of 2008, iron ore prices have fallen dramatically from their peaks based on the world economic downturn. The price for iron ore, however, has already increased in 2009, and is expected to be very high in 2010 and 2011, increasing by almost 50%. The selected price takes into account the metal unit factor estimated at 65% FeT for the concentrate. The


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sale price is estimated at CAN $57.90 FOB located closer to the port facility. This value is equivalent to a long term price of US 0.85 cents/dmtu for a concentrate fixed at 65% FeT and 5% of moisture.

In 2009, the average price was in order of US$53/t or US 0.85 cents/dmtu; however, iron ore pricing is expected to increase by 50% in 2010 and 2011, close to US$80/t. For conservative approach, iron ore price has been fixed at CAN 57.90$/t of concentrate in regard to long term.

Table 3.6 — Iron Ore Sale Price

Price at FOB I Sales Price ($/t conc.) 57.9 I

The economic pit limits were estimated using overall pit slopes of 55 degrees for each deposit. No geotechnical data and study is available to define precisely the pit slopes of the simulated pit shell. Rock types observed at Duncan Lake Iron Ore Project are strong, hard with joints sets fabric. The selected pit wall angle of 55 degrees is achievable, comparing with other mining operators located in northern Quebec. To reach this objective, DLIOP needs to obtain geotechnical data, study and monitor the wall excavations, as well as the drilling and blasting process to avoid damaging the final pit wall.

Table 3.7 — Pit Wall Angle

Pit Wall Angle 'Degrees, all direction 55 I

There are no density tests made on the ore and waste rock. Density for ore and waste is estimated from similar projects taking the geology description into consideration. In fact, the density factor will change with the %FeT. The densities used for this study are shown in Table 3.8.

Table 3.8— Rock Density

Density Density - Ore (t/m3) 3.2

Density - Waste (t/m3) 2.8

3.4 Simulation of Mineral Reserves

Mineral reserves need to be defined to support a mining project. To initiate the mineral reserve simulation, it is necessary to define operating cost of mining and milling, fixed administrative costs and an FOB iron ore price as the minimum parameters for LG algorithm using MineSight mining software to define mineral reserves, tonnage and grade based on WR defined by metallurgical tests.


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The determination of economic pit limits and reserves for each deposit were made using the EPIT module of MineSight mining software. The EPIT module uses the Lerch Grossman (LG) pit optimization method to obtain the optimum net revenue value pit. A number of parameters are input into the software to calculate the results, including mining, processing and administration costs, product sales price, and maximum pit slope angles. The software then applies these parameters to the ore and waste blocks of the 3D block models to calculate the optimal net revenue pit. Parameters were discussed in the previous section.

A simulation of the mining reserves has been done with the measured, indicated and inferred mineral resources. Total tonnage is estimated at 720 Mt at 24.65% FeT, weight recovery of 28.9% and stripping ratio of 2.26, details are shown in Table 3.9.

Table 3.9 - Simulation of Mining Reserves - All Deposits

DUNCAN LAKE IRON ORE DEPOSITS - Mineable Resource - Measured, Indicated & Inferred

Deposit Ore FeTotal WR Si02 Mn0 P205 S Waste Strip Ratio

(#) (Mt) (%) (%) (%) (%) (%) (%) (Mt) (o/w)

Deposit 1 261.1 23.87 27.72 51.04 0.054 0.143 0.518 584.1 2.24

Deposit 2 68.0 27.02 32.37 49.22 0.046 0.135 0.517 210.5 3.09

Deposit 3 132.6 24.81 29.11 34.32 0.038 0.096 0.463 296.6 2.24

Deposit 4 183.4 24.56 28.74 27.32 0.023 0.070 0.180 327.9 1.79

Deposit 5 10.4 24.56 28.73 49.63 0.071 0.172 0.613 33.5 3.21

Deposit 6 64.8 25.26 29.77 n/a n/a n/a n/a 175.0 2.70

TOTAL 720.4 24.65 28.88 n/a n/a n/a n/a 1627.6 2.26

A simulation of the mining reserve has also been done with all mineral resource categories including the potential tonnage. Although the potential tonnage is not considered part of the by the NI 43-101 standards acceptable mineral resource categories, for purposes of this simulation, the potential tonnage was taken into consideration. Total tonnage is estimated to 1,074 Mt at 24.59% FeT, weight recovery of 28.8% and stripping ratio of 2.47. Details are shown in Table 3.10.

Table 3.10 - Simulation of Mining Reserves Including Potential Tonnage All Deposits

DUNCAN LAKE IRON ORE DEPOSITS - Mineable Resource - Measured, Indicated, Inferred & Potential Tonnage

Deposit Ore FeTotal WR Si02 MnO P205 S Waste Strip Ratio

(#) (Mt) (%) (%) (%) (%) (%) (%) (Mt) (o/w)

Deposit 1 360.3 23.72 27.49 51.15 0.054 0.142 0.524 926.6 2.57

Deposit 2 80.4 27.01 32.35 49.23 0.047 0.135 0.519 255.7 3.18

Deposit 3 298.2 24.81 29.11 34.59 0.039 0.097 0.459 728.8 2.44

Deposit 4 211.6 24.50 28.65 26.95 0.022 0.070 0.177 392.4 1.85

Deposit 5 12.5 24.60 28.80 49.58 0.071 0.171 0.607 41.3 3.30


TOTAL 1,074.4 24.59 28.79 n/a n/a n/a n/a 2652.2 2.47


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V

Robert-Bourassa Resevoir

r 0 1 2 3 4

Kilometers


Page 20

indicates the surface area of the pit envelopes of the mineral resources simulation of the measured, indicated and inferred mineral resources only. Areas of the potential tonnage are not included in this figure.

Figure 3.1— Economical Pit Envelops of Deposits 1 to 6

Taking into account only the eastern deposits (Deposits 3, 4 and 6) with coarser grain and associated better liberation, the Pit Economics Study gave the following results (Table 3.11).

Table 3.11— Grade and Tonnage for low Silica Deposits

Total Reserves (Measured + Indicated + Inferred)* + (Potential)**

Block # Max. Tonnage** Tonnage* FeT WtR Stripping

Block 3 98.2 132.6 24.8 29.1 2.2 Block 4 211.6 183.4 24.6 28.7 1.8 Block 6 111.4 64.8 25.3 29.8 2.7 Total in Mt 621.2 380.8 Weighted Ave. 24.79 29.03 2.09 At 12 Mtpy conc. ROM @ 29% WR 41,341,575 41,341,575 Mine Life, years 15.03 9.21 Waste @ 2.09 90,640,064.32 86,504,639

Total Moved/year 131,981,639 127,846,214


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Alternatively, it is possible to force the software to calculate various pit limits based, not necessarily on the optimum, but on mine life, maximum stripping ratios, etc. This was performed in order to evaluate a number of economic pit shells by varying the stripping ratio in order of 1 to 1; thereby, affecting the ore and waste tonnage for each pit, the net margin and mine life. As shown in Table 3.12, a lower stripping ratio mining development causes a major decrease in the mining reserve. By decreasing the stripping ration from 2.26 to 1.09, the tonnage decreases from 720 mt to 366 mt.

Table 3.12 - Simulation of Mining Reserves - All Deposits Stripping Ratio Forced

DUNCAN LAKE IRON ORE DEPOSITS - Mineable Resource - Measured, Indicated & Inferred

Deposit Ore FeTotal WR Si02 MnO P205 S Waste Strip Ratio

(#) (Mt) (%) (%) (%) (%) (%) (%) (Mt) (o/w)

Deposit 1 121.5 25.02 29.42 50.47 0.051 0.145 0.499 138.2 1.14

Deposit 2 28.6 27.12 32.51 49.18 0.045 0.140 0.499 29.7 1.04

Deposit 3 71.8 24.90 29.24 33.82 0.038 0.095 0.465 75.8 1.06

Deposit 4 115.5 24.84 29.15 25.20 0.020 0.062 0.163 125.3 1.08

Deposit 5 5.3 24.59 28.79 49.27 0.072 0.161 0.591 5.1 0.96


TOTAL 366.2 25.17 29.64 n/a n/a n/a n/a 399.9 1.09

No detailed pit design and mine plans have been made on the simulated pit shell envelop.

3.5 Production Rate

Because of the high investment required to bring a project like this to reality, the project would start with the three low silica Deposits 3, 4, and 6 as there is less grinding and no flotation required which would reduce the initial costs. A production rate of 12.0 Mtpy of pellets has been assumed, without correction for the weight gain due to additives and to magnetite conversion or the loss due to the calcinations of carbonates in the pelletizing process.

Three other pellet production scenarios were used for the analysis of the potential of the project, namely 15 Mtpy, 18 Mtpy and 21 Mtpy.

3.6 Mining Concept and Infrastructures Location

Deposits 1 to 6 extend over more than 20 km South-West to North-East. For the opportunity study, the mine development is to start on the eastern part in order to optimize the infrastructure location and because of expected better iron ore liberation. Consequently Deposits 3, 4 and 6 become the areas of the initial project mine development.

Mining and milling facilities are assumed to be located between Deposits 3 and 4, closer to Deposit 3 because this deposit has a better chance of tonnage increase, primarily based on geophysical magnetic anomalies. Consequently the iron ore coming from Deposits 4 and 6 will have to be transported over a greater distance. Also this area is located near major hydro electricity power lines and very close to the main paved road.

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Some other considerations lead to the location of the other infrastructures. The concentrated iron ore needs to be transported closer to a navigable point. The closest and possible port location was selected north of Chisasibi Village, with a natural penetrating dyke reaching a water depth that is needed for ocean going boat transportation and named Wastikun I. A bathymetric map shows that this natural dyke gives access to a water depth of 18-20 meters (see).

Figure 3.2 — Bathymetric Map


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obert-Q Bourassa

Res,ervoir

Kilometers


Page 23

The iron ore concentrate has to be transported to the port facility area. Based on a short analysis, the pipeline method was selected as the best transportation method when compared with train and truck transport over an estimated distance of some 140 km. The pipeline will be located along the actual road up to the LG1 power station, crossing the La Grande River on the bridge again following a gravel road in a north-west direction up to the Wastikun I port area, shows the location of the major infrastructures of the mine and milling complex, closest cities, major road and pipeline as well as the planned port location.

Figure 3.3 — Infrastructures Location Map

Figure 3.4 shows the conceptual layout of the major infrastructures closer to the port facility in the area of Wastikun Island.

3-mEr•cHEm December 2010

OPF-009-12/B


James Bay

Pellet Plant

d

Port ,a.

Storage Access

Pipeli e Rltra Plant

,;Ç)— .. C2 °

2 3 4 5

Kilometers


Page 24

Figure 3.4 — Port Area Infrastructures

kET•CHEnI December 2010

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4.0 CRUSHER

It is considered at this point to use two identical gyratory cone crushers. The crushers are housed in an enclosed building with two (2) ROM ore pockets of sufficient capacity to receive the ore from the mine trucks, an overhead crane, a maintenance area, a control room and electrical, sanitary and other services, a dust collection system and hydraulic rock breakers. The crushers will have a close setting of about 200 mm. The crushed ore will fall into a pocket from which it will be extracted by 2 apron feeders onto two common belt conveyors to be carried out to a conical surge pile.

The ore will be extracted from the surge pile by apron feeders, located in a concrete tunnel below the surge pile, and carried by a conveyor system to a covered rectangular ore storage building with a capacity of about 1 day. The storage building will have access for mobile equipment. A tripper conveyor will distribute the ore across the storage building.


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5.0 CONCENTRATOR

5.1 Primary Grinding SAG

It is considered in this study that 2 (two) SAG mills are used. Each will be fed by a conveyor by 3 variable speed apron feeders located in a concrete tunnel, directly below the ore storage building. The -200 mm crushed ore will be fed directly into the SAG mill along with pulping water to maintain the solids at about 70%. Each SAG mill will be equipped with a 6 mm trommel screen with pocket lifter to return the oversize directly inside the mill by means of a sluice through the discharge trunnion of the SAG mill. The trommel undersize is pumped to an agitated slurry mixing tank located ahead of the concentration process. The mill will be operated from the central plant control room. Test work will be required to define the energy requirement and size the SAG mills.

5.2 Alternative, Secondary Crushing and High Pressure Grinding Rolls

An alternative to SAG mills would be to have a second stage of crushing with a number of cone crushers set at about 50 mm. These crushers would be located in a separate building erected between the surge pile and the ore storage building. The -50 mm crushed ore will be conveyed to the tripper conveyor and distributed across the ore storage building.

This alternative uses a number of High Pressure Grinding Rolls (HPGR) to reduce the ore size from — 50 mm down to about 6 mm. Each HPGR will be fed, trough a surge bin, by a conveyor with 3 variable speed apron feeders located in a concrete tunnel directly below the ore storage building. Each of the HPGR will be in close circuit with a wet screen which will return the oversize through a conveyor system to each HPGR surge bin. The wet screens undersize (-6 mm) will be pumped to agitated slurry mixing tank located ahead of the concentration process.

The main advantage of an HPGR system is a much lower power requirement per tonne of ore ground, and lower metal wear resulting in lower operating cost. From an investment point of view, because of the need to screen and return the oversize to the HPGR surge bin, the HPGR cost will probably exceed the cost of SAG mills. This alternative was not evaluated in this Opportunity Study, but it should be considered at the Pre-Feasibility of Feasibility Study stage. Test work will be required to determine the energy requirement and allow sizing of the equipment.

5.3 Magnetic Separation and Regrinding

A substantial amount of test and pilot plant work will be necessary to define the Duncan Lake project flow sheet. For this appraisal of costs and for the process description, we have considered a typical taconite flow sheet in Appendix A (drawing A1-28078-0201-F). A simplified block diagram representing the process flow is also included in Appendix A.


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The slurry from the mixing tank will be pumped to 2 or more magnetic separation lines. The first step of magnetic separation ("cobbing") will reject a large portion of the liberated non-magnetic gangue. This gangue (final tails), will flow to a hydroseparator for desliming. The hydroseparator overflow will be directed to a tailings thickener. The underflow will be pumped to the tailings pump box where it will, along with the tailings thickener underflow, be pumped to the tailings pond.

The tailings thickeners receive the hydroclassifier overflow, as well as the rougher and the finisher tails. Underflow from the tailings thickener is pumped to the tailings pump box, and to the pond, while the thickener overflow will be directed to the process water tank and back to the plant.

Water will be returned from the tailings pond to the process water tank to be used again in the plant.

The cobber concentrate is received in an agitated slurry tank and pump to banks of cyclones, the finer cyclone overflows report along with the primary screens oversize to a second hydroclassifier where the fine silica overflows to the tailings thickener. The coarser cyclone underflow flows to ball mills.

The ball mills will regrind the cyclone underflow, the rougher concentrate and the secondary stacksizer screens oversize to further liberate the magnetite. The liberation size is not known precisely at this time, but there are indications from the COREM limited laboratory Davis Tube test work that it could around 80% passing 75 microns or less for Deposits 3, 4 and 6. The ball mill discharges are pumped to a group of primary stacksizer screens.

The primary stacksizer screens oversize is directed, along with the cyclone overflow, to the second hydroclassifier. The undersize from the screens is fed to rougher magnetic separators.

The rougher tails are final tails and go directly to the tailings thickeners while the concentrate, which may not be completely liberated, is returned to the ball mills to be reground.

The finisher magnetic separators receive the second hydroclassifier underflow and reject a final tails which is directed to the tailings thickener while the finisher concentrate, which may still contain unliberated particles of magnetite with gangue material, is sent to the secondary stacksizer screens.

The secondary stacksizer screens reject the coarser concentrate to the ball mills for regrinding while the undersize is final concentrate. It is pumped to the concentrate thickener.

The concentrate thickener underflow will be pumped at 70% solids to 2 or more agitated slurry storage tanks ahead of the slurry pipeline. The thickener overflow will be returned to the process water tank. This complete the magnetic concentration process to recover


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12 million tonnes per year of magnetite (Fe304) in a final concentrate containing + 65.0% Fe and less than 5.0% silica. Non-magnetic, hematite, iron and a minor amount of magnetite will be lost to tails along with the gangue minerals. Emergency Diesel generators are included for life line equipment.


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6.0 SLURRY PIPELINE

The total length of the pipeline from the plant at Radisson to the slurry reception point at James Bay is estimated at 150 km with a negative difference of elevation of about 400 meters to the sea shore. It is assumed that the line from the plant will intercept the highway going to Chisasibi and follow it to La Grande where it will cross over the dam to the north side of the river and follow the road to Bay Wastikun. The annual production to be transported through the pipeline will be 12 Mtpy at about 70% solids or about 2,000 m3/h assuming an effective pumping time of 16 hours per day. See routing sketch as, and Error! Reference source not found. included.

The pipeline feed is the concentrate thickener underflow at about 70% solids. It will be pumped to agitated slurry storage tanks with enough combined volume to store a minimum of 8 hours of plant operation to cover minor production upset or pumping delay. Each tank will be connected to a manifold on the intake side of large, variable speed, high pressure diaphragm pumps in series located in a building adjacent to the concentrator and equipped with adequate lifting and maintenance equipment.

Provision will be made to automatically add dilution water, as required, to maintain a specific density in the line. Provision will also be made to go to full water flow to push the slurry or flush the line and to add additive to prevent line corrosion. For maintenance purposes, by-passes will be provided to allow isolation of each pump in the series (one at a time) without affecting the pumping capacity or the pressure requirement. Communication, instrumentation, control and protection systems, including slurry temperature measurement, customary on installations of this nature are included, and will be located in the concentrator control room. In case of power failure the pumps will also be connected to the concentrator emergency Diesel generator plant.

It is likely that the line will be buried for the full length except for river and above ground crossings where the pipe will be insulated, and possibly heat traced, to reduce the risk of freezing in case of upset operating conditions. At the Pre-Feasibility or Feasibility Stage, freeze protection of the pipeline will be considered. Options may be slurry heating or in case of stopping, flushing with glycol and water.

The slurry will be received at the pellet plant via a distribution tower, and will be directed to agitated slurry tanks or to a concentrate thickener (beginning and end of batches are lower percent of solids). The slurry from the tank will be pumped to the filter plant located adjacent to the pellet plant from where it will be pumped and de-watered ahead of the pellet plant.

Communication, instrumentation, control and protection systems customary on installation of this nature are included.

Test work with concentrate will be necessary to define the pumping characteristics of the slurry and allow sizing of pumps and pipeline. This should also include visit of the proposed pipeline line corridor.

r December 2010



Augyva Mining Resources Inc

Section 7 Opportunity Study Report for the Duncan Lake Iron Ore Project

Page 30

7.0 FILTERING AND PELLETIZING PLANTS

7.1 Filtering

The filtering plant is part of the pellet plant complex. The concentrate received from the slurry pipeline will have been ground at 80% - 75 microns at the concentrator. It will be pumped from the slurry tanks to the filtering plant, adjacent to the pellet plant, at about 70% solids where it will be received in a mixing tank were it will be blended to the client specifications with dolomite and/or limestone slurry. These two flux materials can be ground at the pellet plant site in a small ball mill and stored in slurry tanks to be metered out, as required, to the mixing tank.

The filtering plant cost is included with the, two lines of 6 Mtpy each, pellet plant. For this reason two separate filtering lines are considered. Each line will be composed of several filters (number, to be defined by filtration test work), external tube vacuum disc filter, 2.74 m, 12 discs, (Scanmec or equivalent). The vacuum for each filter will be supplied from a manifold connected to water ring vacuum pumps. The filtrate from the filter will be pumped to the concentrate thickener. The cake, discharged at about 9% moisture, will be conveyed to a filter cake bin with an 8 hours storage capacity in an arrangement similar to that found at the ArcelorMittal QCM plant.

Cake from the filtering lines can be by-passed to a cake storage building to have the option to sell concentrate or in case of shut down of one or both pelletizing lines.

It is possible that the cake moisture content is too high. In this case, sparging of steam (as done at Rio Tinto's IOCC plant) in the mixing tank to raise the temperature of the pulp could help to reduce moisture as may filtration additives. If these fail, a pressure filter will be required.

Communication, instrumentation, control and protection systems, customary in installations of this nature, are included and will be located in the pellet plant control room. In case of power lifeline equipment the pumps will be also connected to the pellet plant emergency Diesel generator plant.

Filtration test done at Pre Feasibility or Feasibility Study stage will allow decisions on the most suitable filtration system.

7.2 Mixing

To be able to make balls with the fine iron concentrate prior to induration, a binder is required. Bentonite is generally used because of price, but organic binding agents can be used, if silica level in the pellets is a problem. It will be necessary to import bentonite to the pellet plant site. It can be purchased as either crude bentonite at -35 mm or ground bentonite. A bentonite storage building will be required in all cases, but, if crude bentonite is used, a dry roller mill system with associated dust collection, storage bins, air conveying system etc. will be required.


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For each balling line, the filter cake will be extracted from the cake bin by a belt feeder, weighed on the conveyor. Bentonite will be added in the right proportion, and the combined material will be discharged into a high intensity plough or bowl type mixer where the bentonite and the fine concentrate will be intimately mixed. The mixed material will be conveyed by a conveyor system to the top of the silo ahead of each balling unit. V ploughs will discharge, in sequence, the mixed balling feed to each storage silo.

7.3 Balling

For each balling module, the mixed balling feed is extracted from the silo by a belt feeder or table feeder, weighed and conveyed to a balling module. Balling can be done with balling drums or balling discs. Balling drums, which require a larger building and more conveying, produce a tighter sized product because the roller screen is in a closed circuit. Balling discs, which require less space and less material handling gives a large size distribution, unless equipped with a roller screen. Both systems are used in the industry and in the end it becomes a matter of cost and client preference. The green balls produced by each balling unit are collected on a conveyor and directed to the induration equipment.

7.4 Pellet plant

The pellet plant, to be located near the James Bay at Wastikun Bay, Quebec is composed of two identical filtering, balling and induration lines each able to produce 6 Mtpy of blast furnace pellet of various "basicity" to meet international client demand. Utilization is expected to be over 90% of total time. The best operated plants achieve better than 94% utilization. The cost appraisal is based on two straight grate units but grate-kiln units could also be used. Both produce quality pellets. The straight grates are used in about 65% of the market while the grate-kilns are used in about 35% and shaft furnaces have about 5% (mostly in China).

7.5 Feed System

For each induration unit, the green balls from the balling section are received on a reciprocating conveyor. This conveyor deposits them in consecutive rows on a 4 m wide belt, feeding a roller screen which removes the — 8 mm fines generated during the transport left behind, if there was a previous screening operation. If open circuit balling disc are used, a double deck roller screen is installed to remove the + 18 mm oversized green balls and return them, after disintegration, along with the fines to the cake bins ahead of the balling. The green balls, sized at +8 to -18 mm, discharge from the roller screen to the hearth layer covering the pallet car grate bars. The height of the green ball is measured by ultrasonic probes at the bridge of the machine. These adjust the machine speed to maintain a constant bed height.


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7.6 Induration and Hearth Layer

Each 4 meter wide straight-grate will have the capacity to produce 6 Mtpy of fluxed blast furnace pellets, and is expected to have an effective grate area of about 600 m2 based on the expected utilization of 90% mentioned at item 7.4 above. The straight grate induration section of the machine has generally five zones, up-draft and downdraft where the green balls are dried slowly, pre-heat, firing and after firing, where the temperature is brought progressively to about 1250° C resulting in bonding of the grains of minerals, calcining of part of the fluxes and oxidizing of the magnetite to hematite in an exothermic reaction which contributes significantly to the induration process fuel requirement. The fired pellets then move to the cooling and second cooling where the pellets exit the hood at about 100° C. The thermal energy to indurate the pellets comes in part from Bunker C heavy oil that is burned in several burners on the down corners on each side of the induration section and from the oxidation of the magnetite. Refractory of various type and thickness are provided all along the induration machine.

The hot gases from the cooling zone are efficiently recirculated though a recuperation header located above the furnace hood while the cooler second cooling gases are pushed back to the front end of the machine to the up-draft zone where lower temperature are required to dry the green balls. Thermally the straight grate is quite efficient with 5 fans with dampers supplying, recirculating or exhausting the air/gases in the process.

Depending on their source, the rejected gases will be cleaned first through a multi-tube dust collector to remove the coarser dust and finally in one or more electrostatic precipitator ensuring a clean exhaust stack. Three wet scrubber will collect the dust from the hearth layer system, the machine discharge and the segregation bin areas and return it as slurry to the concentrate thickener.

To reduce the temperature at the grate and improve gas distribution, the grate bars of the pallet cars are covered with about 100 mm of hearth layer (recirculated pellets) as a bed for the green balls. In a similar manner, side layer protect the sidewall of the pallet cars. The hearth layer pellets come from a segregation bin located at the discharge of the product conveyor that removes fired pellets from the machine discharge. In the segregation bin the larger pellet roll to the side of the bin and are diverted to the hearth layer conveyor and returned to a bin at the feed end of the straight grate. The larger diameter of the pellets reduces plugging of the grate. The pellet product is removed from the bottom of the segregation bin, by a variable speed conveyor and proceeds to the stock pile conveyor system. The speed of the conveyor removing the product from the bottom of the segregation bin is controlled by the level of pellet in the segregation bin. If justified some pellet screens could be added to screen the pellets discharged from the bottom of the segregation bin. In this case a screw classifier and a pellet chips regrind mill will need to be added. The screw classifier fines and the reground chips will be pumped to the concentrate thickener. An emergency overflow at the segregation bin will permit operation in case of failure of the material handling system.

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Communication, instrumentation, control and protection systems customary on installation of this nature are included and will be located in the pellet plant control room. In case of power lifeline equipment the pumps will be also connected to the pellet plant emergency Diesel generator plant.

Balling test work will be required to define the balling characteristics of the concentrate and pot grate test on the produced green balls will be needed to establish the physical and metallurgical quality of produced pellets. Tests for the induration cycle are required (induration temperature per zone, gas flow, residence time etc.) to define the grate factor. These tests should be done at the Feasibility Study stage of the Project. They will determine, in addition to the balling characteristics, whether or not further grinding is required beyond that necessary for the concentration process in order to produce quality pellets.

7.7 Auxiliary Systems and Infrastructure for the Pellet Plant and Port

A number of support facilities will be required for the pellet plant and port. Amongst them:

• Port and pellet plant offices;

• Shop and warehouse;

• Steam plant;

• Emergency generator station;

• Bunker C storage and pumping;

• Propane tanks for pellet plant start up;

• Raw bentonite storage, grinding and distribution system;

• Limestone stockpiling, reclaim and grinding and ground product storage tank;

• Dolomite stockpiling, reclaim and grinding and ground product storage tank;

• Fire fighting/protection and emergency;

• Vacuum and process air plant;

• Environmental protection;

• Process and potable water system;

• Excess water treatment plant;

• Roads and accesses;

• Communication;

• Waste water and sewage disposal; and

• Mobile equipment.


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8.0 PORT FACILITIES, STOCKPILING, RECLAIMING AND SHIP LOADING

8.1 Concept

The port facilities include a stockyard of about 3 million tonnes capacity with stacker and reclaimer, a conveying system with a capacity compatible with ship demurrage cost, a ship loading berth complete with a ship loader, and a year-round port able to accommodate ocean going ships. A second berth with truck access may have to be provided to unload smaller ships which will bring in the Bunker C heavy oil for the pellet plant, the bentonite, the dolomite/limestone, if no local source can be found, the plant operation supplies such as grinding balls, mill liners, spare parts and general supply. It is likely that some dredging will be required for the port and the turning basin. Considerable amount of bathymetry will have to be done to define the water depth in the port and turning basin area. The capacity of the stockyard will be better defined once the length of the shipping season is known while considering the used of reinforced ships assisted by an ice breaker.

8.2 Pellet stockpiling and Reclaiming

The conveyor from the pellet plant will feed at a rate of over 1500 tph a lofting and slewing stacker running on rails on top of a compacted berm. Each type of pellet will be discharged at a specific location on both sides of the berm. The pellets will be reclaimed by a bucket wheel reclaimer to the ship loader conveyor system. Sampling of pellets will be provided.

8.3 Jetty and Ship Loading Berth

The Opportunity Study considers that the jetty will be made of a reinforced concrete deck supported by large diameter concrete filled steel piles. The jetty will support a radial ship loader. The sea going ships will move along a line of "dolphins" to permit complete loading. The main jetty will be wide enough to accommodate the conveyor and supply lines and man access. The offloading jetty will be in shallower water and be equipped with unloading equipment for bulk material (bentonite) and lifting equipment for supplies. It will be serviced by trucks and an intermediate covered storage will be provided, on shore, for bentonite and supplies. A pipeline will be provided to offload Bunker C and Fuel oil to a tank farm near the pellet plant. The light fuel oil will be shipped by tanker truck to the mine area tank farm.

8.4 Metallurgical Tests

With regard to the current metallurgical test results obtained, Met-Chem suggests initiating some metallurgical tests to look at others aspects such as ore liberation at different grain sizes, classification and a multi stage grinding possibility, to limit energy consumption by avoiding the grinding of all material at the same size, and to optimize process flow sheet, for iron recovery, and to increase the final grade of ore concentrate.

Also pelletizing tests are needed for the metallurgical process.




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9.0 MANPOWER

Based on similar type of operation it is expected that the various plants will have a total workforce of about 1,000 people, including supervision and management.

ImEr•cHEm December 2010

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10.0 COMPARABLE IRON ORE PROJECTS IN QUEBEC

As described earlier the Duncan Lake Iron Project analysis of capital and operating costs has been based on publications for other iron ore projects found on the SEDAR internet site and on other Met-Chem in-house documentation. The following similar iron ore projects were consulted for this report;

a) LabMag Iron Ore Project (2006);

b) Bloom Lake Project (2008)

c) Baffinland Iron Mines — Mary River Project (2008); and

d) KéMag Iron Ore Project (2009).

Short descriptions of the data of these projects are given below.

10.1 LabMag Iron Ore Project

The LabMag Iron Ore deposit contains the iron deposits in the Howells River area of Labrador. The property is located in Western Labrador, about 30 km northwest of the town of Schefferville, Quebec, situated 530 km north of Sept-Iles, Quebec. The deposit extends for about 30 km northwest-southeast parallel to the provincial boundary between Labrador and Quebec, is up to about 4 km wide and covers a total area of approximately 64 km2 (6,400 ha). The iron formation at Howells River consists mostly of recrystallized chert and jasper with bands (beds) and disseminations of magnetite, which is the target mineral for iron exploitation. The iron formation is exposed over the entire length of the property and is wedge shaped in cross section, perpendicular to the strike.

The by Watts, Griffis and McOuat (WGM) in 2006 estimated total mineral resource (at a cut-off grade of 18% DTWR) is shown in Table 10.1.

Table 10.1— Howells River Estimated Mineral Resources (2006)

Description Million Tonnes Fe % DTWR % % Fe

DTC SiO2

DTC Measured Resources 3,084.0 29.8 26.8 70.1 2.1

Indicated Resources 581.0 28.8 24.3 69.5 2.4

Inferred Resources 1,475.0 29.6 26.0 69.7 2.3

DTWR = Davis Tube Weight Recovery) (DTC = Davis Tube Concentrate)

The pilot plant testwork utilized a typical magnetic recovery flowsheet incorporating two stages of grinding, two stages of magnetic separation and screening. The concentrate slurry will be dewatered to 70% solids and pumped in a 230 km slurry pipe line to the pellet plant, located in Emeril, near Ross Bay Junction, on the QNS&L railway line, which is the northern limit of the IOCC portion of the railway line to Schefferville. The pellet plant will have a capacity of 15 Mtpy and includes two processing lines. The

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pellets will be shipped south by rail, to ship loading facilities at Pointe Noire, adjacent to Wabush Mine shipping facilities.

The mine will be a large scale open pit mine producing an average of 51.6 Mtpy of ore and 7.7 Mtpy of waste over a 25 year mine life. The estimated (2006) estimated capital cost of the project is shown in Table 10.2 and the (2006) estimated operating costs in Table 10.3.

Table 10.2 — Howells River Estimated Capital Cost (2006)

Description Million $

Mining Complex 1,031.5

Pipeline 307.6

Pellet Plant 702.6

Port Installations 264.4

Railroad , Helicopter Pad etc. 231.4

Total Direct Estimated Costs 2,537.5

Total Indirect Estimated Costs 701.0

Total Estimated Costs 3,238.5

Table 10.3 — Howells River Estimated Operating Costs (2006)

Description

Mining Ore 1.45 tonne Ore

Mining Waste 1.45 tonne Waste

Processing 9.12 tonne Concentrate

Pipeline Transport 0.86 tonne Concentrate

Pellet Plant 7.56 tonne Pellets

Rail Transport 9.40 tonne Pellets

Port Handling 0.78 tonne Pellets

Administration 0.83 tonne Pellets

Head Office 0.58 tonne Pellets

10.2 Bloom Lake Project

The Bloom Lake property forms part of the southwestern corner of the Labrador Trough iron range and is located in close proximity of a number of producing mines in the province of Quebec approximately 940 km northeast of Montreal, 8 km north of ArcelorMittal Mines of Canada's Mont-Wright iron mine and 13 km northwest of the town of Fermont.

The deposit is part of the highly folded and metamorphosed southwestern branch of the Labrador Trough or Labrador-Quebec fold belt. The Labrador Trough extends along the




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margins of the eastern boundary of the Superior-Ungava craton for more than 1,200 km and is up to 75 km wide.

Consolidated Thompson Iron Mines Ltd., the owner of the property, completed the drilling of the property in 2008 and had the mineral resources estimated for the Bloom Lake West Property. The report was published in March, 2009. An open pit optimization and pit design was prepared. The Mineral Reserves of the ultimate pit design and published in May 2006 and using a cut-off grade of 15% Fe, are shown in Table 10.4.

Table 10.4 — Mineral Reserves in Ultimate Pit Design (2006)

Description Million Tonnes

Fe /o WR % Mag% Fe CaO% MgO% Mag %

Proven 463.4 30.1 38.3 7.6 2.2 2.1 10.5

Probable 116.2 29.7 37.7 7.7 2.3 2.1 10.7

Total 579.6 30.0 38.2 7.6 2.3 2.1 10.5

Waste 563.8

(Waste includes the Inferred Resources within the pit limits. The estimated stripping ratio is 0.97 (tonnes/tonnes).

Based on the results of the metallurgical testing, the average calculated weight recovery is 38% in hematite and 41% with the magnetite recovery plant, equivalent to a concentration ratio of 2.63 tonnes of Run-of-mine (ROM) per tonne of concentrate in hematite and 2.43 with the magnetite recovery plant.

The Bloom Lake Property is mined using conventional open pit mining methods based on truck/shovel operation at a rate of 8 million tonnes of concentrate per year. The mining equipment is leased, operated and maintained by personnel of the Corporation. Ore is transported to a crusher, 1,000 m north of the open pit, stockpiled and fed to a single wet SAG mill, operating in closed circuit with vibrating screens. The liberated hematite and magnetite minerals will be recovered using three stages of spirals. Tailings of the spirals plant are passed over wet drum magnetic separators to recover unliberated magnetite particles. This material is reground in a ball mill followed by several more magnetic separator stages to produce a high-grade magnetite concentrate. Concentrate is filtered and loaded into railcars for transportation to the port.

The capital cost for the 8.0 Mtpy concentrate production project was estimated at $486.4 million (no further details have been shown in the report) which included the working capital. The estimated operating costs in US$ is shown in Table 10.5.

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Table 10.5 - Bloom Lake Estimated Operating Costs (in US$ - 2008)

Description Mining Ore US$ 1.76 tonne Ore

Mining Waste US$ 1.76 tonne Waste Processing US$ 4.18 tonne Concentrate

Rail Transport and Port Handling US$ 11.87 tonne Concentrate General and Administration US$ 1.54 tonne Concentrate

10.3 Baffinland Iron Mines - Mary River Direct Shipping Ore Project

The Mary River Properties, owned by the Baffinland Iron Mines Corporation, are located in the north central plateau of Baffin Island, approximately 160 km directly south of the settlement of Mittimatalik (Pond Inlet). The four deposits represent high-grade examples of the Algoma-type iron formation, characterized by zones of massive, layered or brecciated iron oxides, variable mixed with banded iron oxide units and iron silicate units. There are four deposits of potential economic interest.

Deposit No.1 is the largest and best known and it is expressed geographically as an arcuate ridge, some 3800 m in length, along which are exposed outcrops of magnetite and hematite. Deposit No.2 outcrops on a ridge 2.6 km east of Deposit No.1 and is revealed as a series of high-grade of specularite iron formation outcrops exposed over 500 m along strike. Iron grades are high, 66% - 68%, and sulphur and phosphorus levels are generally low. Deposit No.3 lies on a ridge 670 m south of Deposit No.2 and drilling and sampling to date indicates similar grades as Deposit No.2. Deposit No.4 outcrops on a low ridge 27 km to the northwest from the locality of the other three deposits. Surface sampling indicates again similar grades.

The planned mine output is 18 Mtpy and the Mineral resource estimate (for all three deposits) and proven and probable reserve estimates were made for the Deposit No.1. A summary of the Resource and Reserve estimates is shown in Table 10.6.

Table 10.6 - Deposit Nos.1, 2 and 3 Mineral Resources Exclusive of Reserves and Mineral Reserves (Deposit No.1)

Description Million Tonnes

Fe /o

P% S% SiO2')/0 A1203%

Mineral Reserves Proven 160 64.4 0.030 0.25 3.52 1.43

Probable 205 64.9 0.039 0.23 3.03 1.25

Total 365 64.7 0.035 0.24 3.24 1.32 Mineral Resources

Measured 0.4 65.4 0.010 0.60 1.86 1.09 Indicated 52 64.6 0.035 0.32 4.39 1.27 Inferred 448 65.5 0.044 0.18 2.56 1.21


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The process plant for the Mary River Direct Shipping Ore Project is designed with the maximizing of lump product as a primary objective. It is a two stage crushing and screening plant which produces two products , a +6.3 mm lump product and a -6.3 mm fine product, each of which is then conveyed to storage piles ahead of the rail loadout terminal.

The two products will be transported, separately, from the mine site to the Steensby Inlet port site via a 143 km long railway. A total of three train sets, each consisting of 2 locomotives and 110 railcars, will be used to transport the ore. At Steensby Inlet, railcars from the mine site are unloaded by a rotary car dumper into surge bins. Ore will be withdrawn from the surge bins and conveyed to the stockpile area. Fines generated in the transport from the mine site will be removed before shipment.

The port facility at Steensby Inlet will be designed to accommodate cape-size ore carriers for 12 months each year at one ore carrier berth. Two service berths are planned to be used by harbour tugs and for delivery of supplies. Other facilities include a diesel fuel tank farm, diesel power generation, camp, general warehousing and the railway maintenance yard and facilities.

The summary of the capital cost estimate is shown in Table 10.7 and the estimated operating costs in Table 10.8.

Table 10.7 — Mary River Project Estimated Capital Cost (2008)

Description Million $ Mining 23

Mary River Site 591

Railway 1,215

Steensby Site 706

Total Direct Estimated Costs 2,535 Support 988

Owner's Costs 86

Gravel 29

Contingency 437

Total Indirect Estimated Costs 1,540 Total Estimated Costs 4,075


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Table 10.8 — Mary River Project Estimated Operating Costs (2006)

Description $/tonne Mining 4.08

Crush/Screen/Load 4.89

Railway 1.45

Camp/Catering 0.59

Aircraft Services 0.60

Leased Equipment 0.40

G&A Onsite 2.25

G&A Offsite 0.36

Operating Cost Total 14.62

10.4 KéMag Iron Ore Project

New Millenium Capital Corporation (NML) is the owner of the KéMag magnetite rich taconite iron ore deposit at Lake Harris, Nunavik, approximately 50 km north of Schefferville, Quebec. The deposit is some 18 km north of the LabMag deposit at Howells River. The KéMag property covers a total area of approximately 81 km2. The deposit was explored by the Iron Ore Company of Canada from 1949 until 1971 and by NML from 2004 to 2008. A Pre-Feasibility Study was issued in 2009. The report covers the technical, commercial and financial aspects of the project to support the annual production of 21.2 million tonnes of concentrate, with downstream processing to result in 15 million tonnes of pellets and 7 million tonnes of concentrate for sale. The open pit mine at Lake Harris would have an average output of some 76 Mtpy of iron ore. The concentrate from the Lake Harris site would be transported through a 750 km long slurry pipeline to a pellet plant in Point-Noire. A jetty and ship loading facility would be designed to permit the loading of Laker-size ships and vessels.

The deposit is an iron formation of the Lake-Superior type which consists of banded sedimentary rocks composed principally of bands of iron oxides, magnetite and hematite within quartz (chert)-rich rock with variable amounts of silicate, carbonate and sulphide lithofacies. The portion of the property that has been explored by diamond drilling has a strike length of 9.5 km, oriented northwest-southeast, and iron formation is present along the entire length and continues beyond the property boundary.

The proposed KéMag iron ore mine will constitute a large-scale open pit operation using conventional drill, blast, load and haul process to mine an average of 76 Mtpy and 30 Mtpy. The open pit mine design, based on a cut-off grade calculation of operating cost and sales of the products, resulted in the total mine life proven and probable mine reserves shown in Table 10.9.

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Table 10.9 — Proven and Probable Reserves (Cut-off grade 18% DTWR)

Ore and Waste Concentrate

Description Million Tonnes DTWR % Crude Fe % Fe % SiO2 %

Mineral Reserves

Proven 1 347 26.9 31.2 69.1 2.8

Probable 794 27.0 31.4 69.1 2.5

Total 2 141 27.0 31.3 69.1 2.7

Inferred 73

Waste 1 015

Total 1 088

Waste : Ore Ratio 0.51

A summary of the capital cost estimate has been shown in Table 10.10 and that of the estimated operating costs in Table 10.11.

Table 10.10 — Summary Estimated Capital Cost of KéMag Project (2009)

Description Million $

Mine * 145.5

Crushers 335.8

Concentrator 690.9

Pipeline 1,023.1

Pellet Plant 879.5

Port Facilities 254.4

Power Supply 107.6

Railway 5.9

Total Direct Estimated Costs 3,442.7

EPCM 284.5

Owner's Costs 325.6

Contingencies 398.2

Total Indirect Costs 1,008.4

Total Estimated Costs 4,451.1

* Mining equipment is leased and not included in initial capital cost

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Table 10.11— KéMag Project Estimated Operating Costs (2009)

Description $/tonne concentrate

$/tonne BF pellets

$/tonne DR pellets

Mining 6.74 6.38 6.38

Mine & Conc. Services 0.41 0.38 0.38

Concentrator 10.15 9.61 9.61

Pipeline 1.09 1.03 1.03

Flotation -- 1.52

Pellet Plant - 8.97 8.97

Filtration Plant 0.82 - -

Port Site Services 0.34 0.34 0.34

Stocking/Loading 0.64 0.64 0.64

Administration 0.19 0.19 0.19

Operating Cost Total 20.38 27.54 29.06


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11.0 EXPECTED PROJECT SCHEDULE

The EPCM part of the project is likely to require about 36 months; however, nothing can start on the site until the Environmental Assessments Process is completed which will require 24 months or more. Since the project will have 2 or more processing lines, it is likely that the first pelletizing line could start a few months short of the 36 months construction period. The logistics of the project will have to be evaluated carefully. The highest priority should be on the delivery of construction equipment, materials and supplies to the concentrator and port site laydown areas, a road distance of about 1,300 km from Montreal.

Some of the major activities are shown on the bar-chart, below.


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Figure 11.1— Project Schedule

ID Task Name 2010 2011 2012 2013 2014 2015 2016 1121314 1121314

1 1121314 1121314 1121314 1121314 1121314

Exploration and Testing

1

I 1 I

2 Pre-Feasibility Study I I

• 3 Decision

4 Feaskiiity Study

5 Emnrmmental Assessment Process I

Concentrator Commission' issionin4

6 Pernik Appicatian for Construction I

7 Decision

8 EPCM

9 Power Transmission Line Canstuc6on I

10 Mine Development I I

11 Concentrator Construction I j

12 tlll■

13 PipelieConstuction I 1

14 Pipelne Carnmissianmrg

15 Pelet Rant Construction I I

16 Peiet Rant Commissioning

1T Stockyard and Shipkxading Construction I I

IM 18 Stockyard and Shiploading Commissioning

19 Production Startup I

• 20 Full Production

Task I

1 MMeatone •

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12.0 CAPITAL COST ESTIMATES

12.1 Scope

The capital cost estimates include all costs required for development, start-up and operation of a complete project starting with the mine and finishing with the shipping of the iron pellets produced. It includes the complete engineering, the labour and supplies required to construct a mine, crusher, concentrator, slurry pipeline, filtering and balling, induration, reception of bulk materials and supply and all support facilities at the mine/concentrator site as well as at the port/pellet plant site. In order to determine the best potential economic scenario various production rates were analyzed namely 3, 6, 9, 12, 15, 18 and 21 Mtpy operations. An earlier report (April 2010), showed the first four scenarios while this updated report includes the last three scenarios as well. The capital cost estimates for these last three scenarios have been based on the 12 Mtpy estimate increased by the "Six Tenth Factor" (or Proration Factor).

12.2 Cost Estimate

The Direct Cost estimate is based on earlier estimated costs of other recent large iron ore projects as described in Section 10.0. The equipment and labour costs were updated to 2010 dollars and as much as possible rationalized for the scope of the Duncan Lake Iron Ore Project.

The Indirect costs were expressed as a percentage of the equipment and transport cost to the site of the equipment.

Table 12.1 below shows the estimated cost per facility.

For additional clarity the estimates have been summarized for the four production areas namely:

• Mining Complex

• Mine, Crushers and Access Roads;

• Concentrator;

• Tailings;

• Water Systems;

• Electrical Supply;

• Heating Plant (half of the total estimated cost);

• Camp (half of the total estimated cost).

• Pipeline

• Pellet Plant Complex

• Pellet Plant;

• Water Systems;

• Electrical Supply.


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• Port Complex

• Port Installations;

• Heating Plant (half of the total estimated cost);

• Camp (half of the total estimated cost).

The summarized estimated capital costs for the 7 scenarios are shown in Table 12.1 and the more detailed capital cost estimates are shown in Tables 12.2 and 11.3.

Table 12.1— Capital Cost Estimates in M$ CAN

Description 3 Mtpy 6 Mtpy 9 Mtpy 12 Mtpy 15 Mtpy 18 Mtpy 21 Mtpy

Mining Complex 475 $ 685 $ 874 $ 1 020 $ 1 141 $ 1 301 $ 1 427 $

Pipeline 273 $ 291 $ 328 $ 357 $ 364 $ 455 $ 499 $

Pellet Plant 271 $ 400 $ 505 $ 597 $ 680 $ 761 $ 835 $

Port 218 $ 232 $ 242 $ 249 $ 250 $ 318 $ 348 $

Total Direct Estimated Costs

1 237 $ 1 608 $ 1 959 $ 2 223 $ 2 435 $ 2 835 $ 3 110 $

Total Indirect Estimated Costs

298 $ 388 $ 469 $ 536 $ 586$ 683 $ 749 $


1 535 $ 1 996 $ 2 418 $ 2 759 $ 3 019 $ 3 518 $ 3 859 $

12.3 Exclusions

The following items are not included in the cost estimate:

Cost increases due to the indexing of wage rates, the price of materials or the cost of equipment:

• Cost increases resulting from fluctuations in currency exchange rates;

• Training before and after hiring;

• Working capital;

• Project financing and interest during construction;

• Taxes and duties.


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Table 12.2 — Capital Cost Estimates in M$ CAN

Previous Studies Adjusted For Augyva Scope (Preliminary)

12 Mtpy 9 Mtpy 6 Mtpy 3 Mtpy

Item Description Cost Cost Cost Cost

Direct Cost

Mine, Crusher And Access Roads 242 $ 212 $ 152 $ 100 $

Concentrator 602 $ 506 $ 397 $ 262 $

Pipeline 357 $ 328 $ 291 $ 273 $

Pellet Plant 568 $ 478 $ 374 $ 247 $

Port Installation 219 $ 215 $ 208 $ 197 $

Tailings Disposal 52 $ 44 $ 34 $ 23 $

Heating Plant 8 $ 7 $ 5 $ 3 $

Camp, Administration 51 $ 46 $ 41 $ 38 $

Water, Concentrator Site 12 $ 11 $ 9 $ 8 $

Water Systems - Pellet Plant Site 6 $ 5 $ 5 $ 4 $

Electrical Power Supply - Concentrator Site 82 $ 74 $ 69 $ 61 $

Electrical Power Supply - Pellet Plant Site 23 $ 22 $ 21 $ 20 $

Sub-Total Direct Costs 2 222 $ 1 947 $ 1 608 $ 1 237 $

Indirect Costs

EPCM Sub-Total (Direct Cost %) 167 $ 146 $ 121 $ 93 $

Owner's Costs Sub-Total (Direct Cost %) 124 $ 108 $ 90 $ 69 $

Contingencies Sub-Total (Total Cost %) 245 $ 215 $ 177 $ 136 $

Total Indirect Costs (Total Cost %) 535 $ 469 $ 387 $ 298 $

Total Cost 2 757 $ 2 416 $ 1 995 $ 1 535 $

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Table 12.3 — Capital Cost Estimates in M $ CAN

Previous Studies Adjusted For Augyva Scope (Preliminary) 12 Mtpy 15 Mtpy 18 Mtpy 21 Mtpy

Item Description Cost Cost Cost Cost

Direct Cost

Mine, Crusher And Access Roads 242 $ 263 $ 309 $ 339 $

Concentrator 602 $ 688 $ 768 $ 842 $

Pipeline 357 $ 364 $ 455 $ 499 $

Pellet Plant 568 $ 649 $ 724 $ 795 $

Port Installation 219 $ 219 $ 279 $ 306 $

Tailings Disposal 52 $ 60 $ 66 $ 73 $

Heating Plant 8 $ 9 $ 10 $ 12 $

Camp, Administration 51 $ 51 $ 66 $ 72 $

Water, Concentrator Site 12 $ 12 $ 15 $ 17 $

Water Systems - Pellet Plant Site 6 $ 6 $ 8 $ 8 $

Electrical Power Supply - Concentrator Site 82 $ 87 $ 105 $ 115 $

Electrical Power Supply - Pellet Plant Site 23 $ 25 $ 29 $ 32 $

Sub-Total Direct Costs 2 222 $ 2 433 $ 2 834 $ 3 110 $

Indirect Costs

EPCM Sub-Total (Direct Cost %) 167 $ 183 $ 213 $ 233 $

Owner's Costs Sub-Total (Direct Cost %) 124 $ 135 $ 158 $ 173 $

Contingencies Sub-Total (Total Cost %) 245 $ 268 $ 312 $ 343 $

Total Indirect Costs (Total Cost %) 535 $ 586 $ 683 $ 749 $

Total Cost 2 757 $ 3 019 $ 3 517 $ 3.859 $

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13.0 OPERATING COST ESTIMATES

The Operating Costs were estimated based on recent published studies (see Section 10.0) and adjusted where necessary to reflect changes in condition such as stripping ratio, fuel cost, labour cost, etc. They include also all the maintenance and repair cost including grinding media and parts etc. The costs are based on the assumptions that all facilities with the exception of the Administration, work 24 hours, 7 days each week and manpower crews are build up for this. The operating philosophy will be developed at the Pre-Feasibility or the Feasibility Study level. Mining cost is estimated at 2.15$/t of ore and 1.60$/t of waste mainly related to more travelling distance from mines to milling facilities located at a specific area closer to Deposit 3, and not so far from Deposits 4 and 6. Concentrator and milling cost is estimated to 10.56$/t of concentrate; however, a figure of 11.00$ was used in the LG module for mining reserve simulation.

13.1 Summary of Estimated Operating Costs

Operating cost per tonne of concentrate is shown in Table 13.1. The ore mining operating cost is estimated at $2.15. Using a weight recovery of 29.03%, the operating cost of ore is estimated at CAN $7.42. The waste mining cost per tonne of waste is estimated at $1.60 and $11.54 per tonne of concentrate due to the high stripping ratio of 2.09. Cost of pellet plan is estimated at $8.97. Total cost per tonne of concentrate is CAD $32.60 and CAD $40.75 per tonne of pellets.

Table 13.1— Summary of Estimated Operating Costs

Description $/tonne $/tonne concentrate

$/tonne pellets

Mining of ore 2.15 7.42 7.42

Mining of waste 1.60 11.54 11.54

Concentrator 10.56 10.56

Pipeline 1.09 1.09

Pellet Plant - 8.97

Stocking/Loading 0.98 0.98

Administration 0.19 0.19

Filtration Plant 0.82 -

Operating Cost Total 32.60 40.75 Note that the Operating Costs are equal except for the Pellet Plant and Filtration. This is because 1 t of concentrate = 1 t of pellets. This will be have to be adjusted after Pot Grate test work

The operating cost is comparatively high due to the relatively low grade of the iron ore, and the high stripping ratio. The mining project would be more attractive at a lower stripping ratio of 1 to 1; however, at this lower stripping ratio, the ore mining reserve decreases substantially. More successful drilling and additional tonnage is required to optimize the mining project development.


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14.0 PRELIMINARY FINANCIAL ANALYSIS

It should be realized that the described preliminary financial analysis is based on data that is not yet compliant with NI 43-101 parameters, and the resources that are used in the analysis are not compliant with NI 43-101 requirements. The results of the analysis are therefore for guidance of the next steps of exploration only.

14.1 Constant Parameters

As shown in the section of capital costs four scenarios have been analyzed to determine the potential of the project. The constant parameters that were established from the preliminary test data and resource estimates are shown in Table 14.1.

Table 14.1— Preliminary Data Used in the Analysis

Description

Total Resources 720 300 000 tonnes

ROM Grade 24.65 % Fe

WR 28.88 %

ROM Grade 3.98 % SiO2

Total Waste 1 627 600 000 tonnes

Stripping Ratio 2.3 : 1

The other assumed constant parameters for the analysis are shown in Table 14.2.

Table 14.2 — Other Constant Parameters Used in the Analysis

Description

Grade of Pellets 67.5 % Fe

5.5 % Si02

Price per unit Fe 1.20 US $

Exchange Rate (US$) 0.98 CAN $

As shown in Section 12.0 the Capital Cost varies for the four production scenarios, but it has been assumed that the unit Operating Cost varies very little. Table 14.3 shows the unit operating costs that are used in the analysis.


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Table 14.3 - Estimated Unit Operating Costs used in the Analysis

Description

Mining Ore 2.15 tonne Ore

Mining Waste 1.60 tonne Waste

Processing 10.56 tonne Concentrate

Pipeline Transport 1.09 tonne Concentrate

Pellet Plant 8.97 tonne Pellets

Port Handling 0.98 tonne Pellets

Administration 0.19 tonne Pellets

It is assumed that the project's construction period will take 36 months and the capital costs are therefore expended over the first three years, one third of the total estimated capital costs each construction year.

A variable amount of sustaining capital is expected to be spent in Year 10 of the production period of the project.

14.2 Variable Parameters

The variable parameters for the analysis are the Capital Cost as well as the assumed Sustaining Capital for equipment replacements. The annual ore production, annual waste excavation and annual Fe pellet production vary for the four production scenarios. It has been assumed that the amount of annual concentrate production is the same as the annual amount of pellets produced. The variable parameters are shown in Table 14.4.

Table 14.4 - Variable Parameters used in the Analysis

Description 12 Mtpy 15 Mtpy 18 Mtpy 21 Mtpy

Ore 42.0 Mtpy 52.0 Mtpy 62.0 Mtpy 73.0 Mtpy

Waste 95.0 Mtpy 117.0 Mtpy 140.0 Mtpy 165.0 Mtpy

Concentrate Production 12.0 Mtpy 15.0 Mtpy 18.0 Mtpy 21.0 Mtpy

Pellets Production 12.0 Mtpy 15.0 Mtpy 18.0 Mtpy 21.0 Mtpy

Capital Cost 2 759.0 $ 3 021.0 $ 3 518.0 $ 3 858.0 $

Sustaining Capital 24.0 $ 29.0 $ 34.0 $ 39.0 $

14.3 Other Assumptions

The preliminary analysis has been prepared before Federal and Provincial Income taxes or any other tax implications.

No depreciation to the capital equipment has been applied.

No royalty payments have been assumed to be due for the project.


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The life of the project has been assumed to be until the total resource base of 720.3 million tonnes has been exhausted, while a required volume of resources has been calculated, for each of the four scenarios, to supply ore over a 25 year project life.

The estimate of the Net Present Value (NPV) has been prepared for an interest rate of 10%.

14.4 Results of the Analysis

The summary of the results of the preliminary analysis for the four scenarios is shown in Table 14.5, while the details have been attached as Appendix B.

Table 14.5 — Summary of the Results of the Preliminary Analysis

Description 12 Mtpy 15 Mtpy 18 Mtpy 21 Mtpy

NPV (Rate of 10%) $686.1 M $866.9 M $764.6 M $732.3 M

IRR 13.5% 15.0% 14.3% 14.1%

Payback Year 6 Year 5 Year 5 Year 5

Life of Project* 20.3 years 16.9 years 14.6 years 12.9 years

Ore Volume Required (25 years) 1,039 Mt 1,298 Mt 1,558 Mt 1,818 Mt * Includes 3 year construction period

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15.0 WORK PROGRAM

Met-Chem has prepared a preliminary work program which describes the exploration phases and the work to be performed on the property in order to bring the project to a Feasibility Study level with the objective to achieve a fast track mining development project.

Met-Chem also prepared a proposal for the drilling of Deposits 3, 4 and 6 in order to increase the mineral resources of the Inferred category for a Scoping Study, and of the Indicated and Measured category for the Final Feasibility Study.

Met-Chem recommends starting on the eastern part of the property, including Deposits 3, 4 and 6. The work program is divided into two (2) phases, passing directly from the scoping level to the complete feasibility study level to minimize delays, generally present between phases for decision making.

A summary of the proposed drilling program is shown in Table 15.1 for the Scoping Study (Inferred resources), and in Table 15.2 for the Feasibility Study (Measured and Indicated resources). A cost estimate for the drilling programs and the studies can be found in Table 15.3 and Table 15.4 for the Scoping Study and the Feasibility Study, respectively.

15.1 Drilling Proposal of Deposits 3, 4, and 6

The drilling proposal has been prepared using the geological model, section interpretation and geophysical survey data recently taken. For Deposit 4, the previous drilling and interpretation results provides relatively precise drilling targets; however, for Deposits 3 and 6, with very limited drilling information, the drilling program was prepared using the geophysical magnetic survey data that was verified by Geophysician, Mr. Joel Simard, consultant for Augyva.

15.1.1 Phase One Drilling Program

As described before the results of this drilling program would outline the Inferred Mineral Resources because of the proposed wide drill spacing. A total of eighty three (83) drill holes have been planned for an estimated total of 26,800 meters. A summary of the holes to be drilled on each of the three deposits is shown in Table 15.1, and details, as well as the map with the planned drill locations, are attached as Appendix C.

Table 15.1— Boreholes Proposal to determine Inferred Resources

Description Number of holes Total meters Deposit 4 27 7,700

Deposit 3 36 12,400

Deposit 6 20 6,700

Total 83 26,800

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15.1.2 Phase two Drilling Program

The results of this program would convert the earlier outlined Inferred Mineral Resources into the Indicated and Measured Mineral resources category which can be used for the Feasibility Study. Because of the lack of accurate information at this time, the number of holes as well as the estimated length of the holes are an estimate only and are used for the cost estimate of the program. Additional more precise details of this drilling phase will be determined during and after the first phase drilling program.

A total of eighty nine (89) drill holes are planned for a total of 28,700 meters. A summary of the holes to be drilled on each of the three deposits is shown in Table 15.2 and details of the planned drill locations are attached as Appendix C.

Table 15.2 — Boreholes Proposal to determine Measured and Indicated Resources

Description Number of holes Total meters Deposit 4 32 9,900

Deposit 3 36 12,400

Deposit 6 21 6,400

Total 89 28,700

15.2 Scoping Study

The major objective of the drilling is to be able to convert the mineral resources into the Inferred category for the Scoping Study. Additional drilling will be required to convert the Inferred Mineral Resources into the Indicated and Measured category required for the feasibility study.

A budget of approximately CAD $5.1 M is needed to develop the Scoping Study. The drilling program represents the major part of the estimated cost for this phase of the exploration.

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Table 15.3 — Work Program Cost Estimate for Scoping Study Level DUNCAN LAKE WORK PROGRAM

Phase 1 (Scoping Level)

Items Year Cost

Preliminary Environment Baseline, Mine site and Port areas 2010 and 2011 $150,000

Mine site and Port Areas

Stakeholders and Native issues 2010 and 2011 $50,000

Review of stakeholders concerns, native issues, meeting with First Nations

Geological mapping and structural details 2010 and 2011 $205,000

2 geologists on 6 months, room and transportation (1 pickup)

Technical Personnel for drilling 2010 and 2011 $265,000

1 geologist, 1 tech, 1 sampler, for 7 months, including room and transportation (2 pickups)

Drilling Direct Cost to reach Inferred category 2010 and 2011 $3,350,000

26800m,150m/day, 2 drill rigs, 6 months, at 125$/m, including mom and transportation

Chemical Assaying 2010 and 2011 $210,000

3000 samples (5m) at 70$

Core Storage and Logging facilities 2010 and 2011 $30,000

Radisson accommodation and storage facilities

Access roads and trails

2010 and 2011

$100,000

Metallurgical Testing 2010 and 2011 $375,000

2500 samples (6m), concentration and pelletizing test work, at 150$ each

Mineral Resources Re-Updated 2011 $80,000

External 43-101 issue, for scoping study

Scoping Study 2011 $250,000

Consultant assisted by external participation

SUB-TOTAL Scoping Study

$5,065,000)

It is recommended to start the environmental baseline study as soon as possible and to define the key points, to ensure success of the fast track development project. A budget of $200,000 is allocated for the preliminary environmental baseline study for the mine site area as well for the potential port area. Since extensive Environment Assessments have been conducted in the disturbed area for the Bay James and Radisson Hydro-Québec projects, it is strongly recommended to conduct a literature review of existing reports on fauna, flora and ecosystems. Preliminary investigation on fish habitat and species should be conducted as well as for water quality including sediment sampling. It should be noted that additional baseline studies requirement might be needed in 2011 following the submission of the Project Notice which is expected to be prepared shortly in 2010.

It is also important to discuss native issues with the local communities at the beginning of the project to ensure the long term objective. Government, municipalities, environment associations, health and social services should be included in the list of project stakeholders.

Geological mapping of Deposits 3, 4, and 6 needs to be conducted to better define the general geometry and the ore shape mineralisation and surface continuities. Previous

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geological mapping carried out in the 1960's indicates outcrops occurrences at many locations. More precise geological mapping will guide the general interpretation and the structural features may help to define megascopic features such as synform, antiform and plunging. Deposits 3, 4, and 6 are interpreted as major and favourable synform structures accompanied by isoclinals and squeezed antiform structures. Depth of the ore deposits is not yet defined, but expected to be great and at least more than 200 meters. Geological mapping and outcrops locations can be located directly using GPS positioning system.

The drilling program represents the main part of the budget cost. The provisional estimate is based on a 26,800 meters drilling program in a total of 83 boreholes for the Scoping Study level. The drilling objective is to substantially increase the tonnage of the mineral resources to sustain a long term mine development project, and to transfer all mineral resources to a minimum of the Inferred category which is considered acceptable for the scoping study. The total drilling program is a continuous process and no delay is expected between the scoping and the feasibility phase in order to minimize the total development period to less than 5 years, including the field development and the construction period. Direct drilling cost includes drill rigs, personnel, materials, consumables, fuel, drill pad preparation, set-up, and room and transportation for drill crews and foreman.

The allowance for the chemical assaying is for the determination of the head grade of the core samples, estimated to be taken at 5m length intervals. The assaying program needs to be reviewed and adjusted with the metallurgical testing results. At the moment, it is expected to carry out XR analyses (collection of elements, Satmagan and sulphur) on half split drill core.

Thereafter, composite samples will be prepared for metallurgical testing at a specified grain size liberation process and Davis Tube testing. More testing will be carried out to determine the minimum grain size liberation and the multi grinding stages process. The cost for the storage and logging facilities is included.

Met-Chem recommends restoring the trails and access roads on all deposits soon, for easy access by pick-up truck. Access to the deposits is considered easily to realize and not expensive taken into consideration the advanced exploration stage of the exploration project. This is considered the best choice for the environmental study, mapping as well as drilling. The access trails will limit delays and wasted travelling time. Many eskers are located close to the deposits and may be used as an easy access road. Only narrow swampy areas with little creeks are expected where crossovers using tubes and culverts would be needed.

Metallurgical testing has to be performed on composite core samples. More metallurgical testing needs to be conducted, however, to better define the final beneficiation process and to characterize the ore parameters used in the geological model. Metallurgical tests are needed to prepare a process flowsheet of the beneficiation plant in time for the Scoping Study, using limited number of representative samples. A


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preliminary review indicates that the required laboratory tests should include the following steps:

• Mineralogy and head assays using latest technology available for mineral studies and grain size liberation;

• High Pressure Grinding Roll (HPGR) testwork;

• Determination of appropriate size for primary (SAG) mill grind;

• Determination of optimum grind of the final magnetic concentrate;

• Grindability tests;

• Flotation tests;

• Setting and sedimentation tests, and

• Filtration Tests.

The pelletizing testwork has been planned for the Feasibility Study.

A new mineral resources model must be made and a resource estimate has to be prepared with all data of the boreholes and the detailed geological mapping from all deposits. Met-Chem recommends carrying out these studies applying the NI 43-101 standards.

The Scoping Study must be prepared compliant with NI 43-101. Met-Chem recommends not stopping or delaying the drilling program and other studies during the preparation of the Scoping report which might take several months to complete.

15.3 Feasibility Study

The major objective of the Feasibility Study is to increase the mineral resources classification to the Measured and Indicated categories, and to produce all information and data for the decision of management regarding the development of the mining project. Table 15.4 presents list of items to be carried out for the Feasibility Study, the time period and the cost estimate.


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Table 15.4 — Work Program Cost Estimate for Feasibility Study Level Phase 2 (Feasibility Level)

Items Year Cost

Base topographic map 2011 $100,000

mining and port areas, pipeline corridor, new digital map definition

Technical Personnel for drilling 2011 and 2012 $265,000

1 geologist, 1tech, 1sampler, for7 months, including room and transportation

Additional Drilling Program to reach Indicated and Measured categories 2011 and 2012 $3,587,500

28700m, 150m/day, 2 drill rigs 6 months, at 125$/m, including room and transportation

Chemical Assaying 2011 and 2012 $210,000

3000 samples (5m) at 70$

Core Storage and Logging facilities 2012 $15,000

Radisson accommodation and storage facilites

Additional Metallurgical Tests 2012 $500,000

beneficiation, pilot test, grinding, energy consumption and all

Environmental Impact Assessment 2013 $500,000

technical report, consultation, permits

Study of Mining Complex 2013 $2,000,000 'Feasibility

mine and benefi ciation, NI 43-101 report,

Port and Shipping Studies 2013 $1,000,000

bathymetry study, infrastructures

Pellet Plant Study 2013 $500,000

NI 43-101 report

ISU B-TOTAL Feasibility Study

$8,677,500 I

TOTAL WORK PROGRAM

$13,742,500)

The production of the topographic base map will involve an aerial survey of the deposits. At this time, it should be advantageous to survey all deposits located on the western and eastern part of the property. An accurate survey and control points in the field are necessary as well as an aerial photo survey. The satellite photo methodology should be investigated.

A drilling program is required to reclassify the Inferred mineral resources into Measured and Indicated categories, compliant with NI 43-101 to support the Feasibility Study.

Additional metallurgical tests are most likely required to support the Feasibility Study, and may include pilot testing on representative ore bulk samples, taken from each deposit as well as a composite sample. Generally samples of some 100 tonnes are used for pilot plant testing.

Environmental impact and others specific studies will be required to support decision making and permits and certification.

Laboratory testwork for the pellet plant will be performed to evaluate the quality and the productivity rate of the selected type furnace: Straight grate process and/or Grate-Kiln


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process. Preliminary review indicates that required tests would include the following items:

• Characterization of iron ore concentrate (size, chemical analysis and characteristics);

• Laboratory balling tests to determine guide line for pilot scale testwork;

• Pelletizing testwork (balling and firing tests) and

• Characterization of pellet quality.

The cost for the Feasibility Study has been estimated including the mining and beneficiation process. Specific studies are needed for the pipeline and concentrate transportation to the pellet plant to be located close to the port area and for the port installation and infrastructures. Finally a study will be required for transportation and sea shipment, period of the year and type of vessel. A provisional cost is estimated for these studies are included but due to the present level of understanding of the mine development project, they might not be very accurate.

The cost estimate of the Feasibility Study is approximately CAD $8.7 M. The total estimated cost for the Scoping and Feasibility Studies combined is approximately CAD $13.7 M. This cost should not be considered too high when compared to similar exploration projects in northern Québec. The exploration property is in fact very well located for exploration, close to a very good road, a town and existing infrastructures. Also, the exploration can be carried out all year, including winter, with no major logistic problems. Only the run-off season from April up to mid May might create difficulties due to muddy roads and difficulties to access specific areas.


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16.0 CONCLUSION AND RECOMMENDATIONS

The project is a challenge, but it is feasible. Major elements and key points are: to get an increase of mineral resources and costs of pellet transportation to customers.

Considering the critical timing of this large project it is urgent to:

• Initiate a drilling campaign to prove considerably more reserves, preferably in Deposits 3, 4 and 6;

• Initiate the Environmental Assessment studies. This item is critical and ultimately controls when and if work can start at the site;

• Select representative core samples for laboratory test work and a representative bulk sample for pilot plant test work to confirm the flow sheet from laboratory test work and allow more precise equipment sizing;

• Initiate bathymetry studies of the proposed port area in James Bay; and

• Initiate a Pre-Feasibility followed by a Feasibility Study to further define the main parameters of the project as well as establish a flow sheet, mine study, mechanical/electrical equipment list, preliminary layout for the mine, concentrator and pellet plant and port sites, develop arrangement drawings for each plant and establish all quantities for civil and steel of all types, develop CAPEX and OPEX and prepare a financial analysis to allow the Owners to make an informed decision on the project.

The Feasibility Study should include budget prices for all the major equipment and also quotations and prices for:

• Slurry pipeline,

• Pellet Plant; and

• Port and ship loading facilities.


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17.0 REFERENCES

The following sources of information were used in the preparation of the Opportunity Study for the Duncan Lake Iron Ore Project:

Met-Chem Canada Inc — Technical Report NI 43-101 on Mineral Resources of the Duncan Lake Iron Ore Project, James Bay Area, Quebec, Dated February 3, 2010;

BBA Inc. — NI 43-101 Technical Report on the Pre-Feasibility Study of the Kemag Iron Ore Project, Quebec, for New Millennium Capital Corp. March 2nd 2009;

Aker Kvaerner — Technical Report of the Definitive Feasibility Study, Mary River Iron Ore Project, Northern Baffin Island, Nunavut. Prepared for Baffinland Iron Mines Corporation dated February 2008;

Augyva Mining Resources Inc., Report on the Duncan Lake Property 2008-2009 Diamond Drilling, made by team's work, June 30, 2009, Internal Report.

Knight Piesold Consulting — Baffinland Iron Mines Corporation, Mary River Project, Development Proposal for the Mary River Project dated March 2008;

Watts, Griffins and McOuat Limited — A Technical Review of the Pre-Feasibility Study of the Labmag Iron Ore Project, Labrador, for New Millennium Capital Corp. Dated August 18, 2006.


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QPF-009-12/B



Appendix B Details of the Preliminary Financial Analysis


Duncan Lake Iron Ore Project

Preliminary Economic Evaluation

General Information

Company Name Project Name Scope Valuation Date

Augyva Mining Resources Inc. Duncan Lake Iron Ore Project Preliminary Economic Evaluation

19-May-10

Description Units Option 1 Option 2 Option 3 Option 4 Sensitivity

Factor Comments Mining

Total Resources Ore Tonnage tonnes 720,300,000 720,300,000 720,300,000 720,300,000 Ore Grade

Fe % 24.65% 24.65% 24.65% 24.65% Si02 % 3.98% 3.98% 3.98% 3.98%

Waste Tonnage tonnes 1,627,600,000 1,627,600,000 1,627,600,000 1,627,600,000 Average Stripping Ratio 2.3 2.3 23 2.3

Pre-production Period Years 3.0 3.0 3.0 3.0 Initial Production Year 1 1 1 1

Process Weight Recovery % 28.88% 28.88% 28.88% 28.88% 1.00 Grade in Concentrate

Fe % 67.5% 67.5% 67.5% 67.5% 1.00 Si02 % 5.5% 5.5% 5.5% 5.5%

Production Target Pellet Mtpy 12.0 15.0 18.0 21.0

Economics Capital Costs

Mining Complex CD$ 1,020,000,000 1,141,000,000 1,301,000,000 1,427,000,000 1.00 Equally divided in pre-production years Pipeline CD$ 357,000,000 364,000,000 455,000,000 499,000,000 1.00 Equally divided in pre-production years Pellet Plant CD$ 597,000,000 680,000,000 761,000,000 835,000,000 1.00 Equally divided in pre-production years Port CD$ 249,000,000 250,000,000 318,000,000 348,000,000 1.00 Equally divided in pre-production years Indirect Costs CDS 536,000,000 586,000,000 683,000,000 749,000,000 1.00 Equally divided in pre-production years

Sustaining Capex Mining Complex CD$ 20,000,000 25,000,000 30,000,000 35,000,000 1.00 10 years Pipeline CD$ 1,000,000 1,000,000 1,000,000 1,000,000 1.0010 years Pellet Plant CD$ 1,000,000 1,000,000 1,000,000 1,000,000 1.00 10 years Port CDS 2,000,000 2,000,000 2,000,000 2,000,000 1.0010 years

Operating Costs Mining

Ore CDS/ t 2.15 2.15 2.15 2.15 1.00 Waste CD$/t 1.60 1.60 1.60 1.60 1.00

Processing & Transportation Concentration CD$/t conc. 10.56 10.56 10.56 10.56 1.00 Pipeline CD$/t cone. 1.09 1.09 1.09 1.09 1.00 Pellet Plant CD$/t pellet 8.97 8.97 8.97 8.97 1.00 Port Handling CD$/t pellet 0.98 0.98 0.98 0.98 1.00 Administration CDS/t pellet 0.19 0.19 0.19 0.19 1.00

Revenue Unit Selling Price US$/unit Fe 1.20 1.20 1.20 1.20 1.00 Exchange Rate (US S) CDS/1 USS 0.98 0.98 0.98 0.98 1.00 Selling Price/Unit of Production

US Currency US$/t pellet 81.00 81.00 81.00 81.00 Canadian Currency CD$/t pellet 82.65 82.65 82.65 82.65

Discount Rate Base % 10% 10% 10% 10% 1.00

ET•CHEnI



Duncan Lake Iron Ore Project 900

850

800

750

700

650

600

550

500

12 15 18 21

Production rate (Mtpy)




Preliminary Economic Evaluation Summary of Results

Description Units Option 1 Option 2 Option 3 Option 4

Production rate Mtpy 12.0 15.0 18.0 21.0 Discount Rate % 10% 10% 10% 10% NPV MCD$ 686.1 866.9 764.6 732.3 IRR % 13.5% 15.0% 14.3% 14.1% Pay Back Years 6 5 5 5 Life of Mine Years 17.3 13.9 11.6 9.9

Resources for 25 years Mt 1,039 1,298 1,558 1,818


QPF-009-12/B


NP

V@

10%

(M

CD

$)





Option 1 - 12 Mtpy

repro ne ton rodorw,~~ Description Factor Coils Total -3 -2 -1 9 10 11 14 15

Mining Depletion

Ore Tonnage Fe% Si02%

Waste Tonnage

Production Ore

Tonnage Fe% Si02%

Waste Tonnage

Stripping Ratio

Mt % %

Mt

Mt % %

Mt

7203 24.65% 3.98%

1,627.6

720.3 24.65% 3.98%

1,627.6

23

720.3 720.3 720.3

24.65% 24.65% 24.65%

3.98% 3.98% 3.98%

1,627.6 1,627.6 1,627.6

0.0 0.0 0.0

0.0 0.0 0.0

0.0 0.0 0.0

0.0 0.0 0.0

720.3 678.7 637.2 595.6 554.1 512.5 471.0 429.4 387.9 346.3 304.8 263.2 221.7 180.1 138.6 97.0 55.5 13.9

24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65%

3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98%

1,627.6 1,533.7 1,439.8 1,345.9 1,252.0 1,1582 1,064.3 970.4 876.5 782.6 688.7 594.8 500.9 407.0 313.1 219.3 125.4 31.5

41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 41.6 13.9

24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65%

3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98%

93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 93.9 31.5

2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3

Process Weight Recovery

Concentrate Production Tonnage Fe Si02

Pellets Tonnage

1.0

1.0

%

Mt % %

Mt

28.88%

208.0 67.50% 530%

208.0

000% 0.00% 0.00%

0.0 0.0 0.0

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.0 0.0 0.0

28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88%

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 4.0

67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50%

5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50%

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 4.0


Mining Complex Pipeline Pellet Plant Port

1.0 MCDS MCDS MCDS MCDS

1,040 358 598 251

340 340 340 119 119 119 199 199 199 83 83 83

20 1 1 2

Direct Costs Indirect Costs

MCDS MCDS

2,247 536

741 741 741 179 179 179

0 0 0 0 0 0 0 0 0 24 0 0 0 0 0 0 0 0


Operating Costs Mining Ore Mining Waste Processing Pipeline Pellet Plant Port Handling Administration

1.0

MCDS

MCDS MCDS MCDS MCDS MCDS MCDS MCDS

2,783

1,549 2,604 2,197

227 1,866

204 40

920 920 920

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0 0 0 0 0 0 0 24 0 0 0 0 0 0 0 0

89 89 89 89 89 89 89 89 89 89 89 89 89 89 89 89 89 30

150 150 150 150 150 150 150 150 150 150 150 150 150 150 150 150 150 50

127 127 127 127 127 127 127 127 127 127 127 127 127 127 127 127 127 42

13 13 13 13 13 13 13 13 13 13 13 13 13 13 13 13 13 4

108 108 108 108 108 108 108 108 108 108 108 108 108 108 108 108 108 36

12 12 12 12 12 12 12 12 12 12 12 12 12 12 12 12 12 4

2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 I Total Operating Costs

Revenue

Net Revenue

Cumulative Revenue

Discount Rate 1.0

MCDS

MCDS

MCDS

%

8,686

17,194

5,725

-

10%

0 0 0

0 0 0

-920 -920 -920

-920 -1,839 -2,759

501 501 501 501 501 501 501 501 501 501 501 501 501 501 501 501 501 168

992 992 992 992 992 992 992 992 992 992 992 992 992 992 992 992 992 332

491 491 491 491 491 491 491 491 491 467 491 491 491 491 491 491 491 165

-2,268 -1,777 -1,287 -796 -305 186 677 1,167 1,658 2,125 2,616 3,107 3,597 4,088 4,579 5,070 5,561 5,725

NPV Cumulative NPV IRR Pay Back

MCDS MCDS

% Years

686 -

14% 6

-836 -1,596 -2,287 -1,952 -1,647 -1,370 -1,118 -889 -681 -492 -320 -163 -28 101 218 325 422 511 591 664 686

ET•CHER1 December 2010

QPF-009-12/B






Option 2 - 15 Mtpy

Pre-production Production Description Factor Units Total -3 -2 -1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 20

Mining Depletion


Waste Tonnage

Production Ore

Tonnage Fe% Si02%

Waste Tonnage

Stripping Ratio

MI % %

Mt

Mt % %

MI

720.3 24.65% 3.98%

1,627.6

720.3 24.65% 3.98%

1,627.6

2.3

720.3 720.3 720.3

24.65% 24.65% 24.65%

3.98% 3.98% 3.98%

1,627.6 1,627.6 1,627.6

0.0 0.0 0.0

0.0 0.0 0.0

0.0 0.0 0.0

0.0 0.0 0.0

720.3 668.4 616.4 564.5 512.5 460.6 408.7 356.7 304.8 252.8 200.9 149.0 97.0 45.1 0.0 0.0 0.0 0.0

24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65%

3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98%

1,627.6 1,510.2 1,392.9 1,275.5 1,158.2 1,040.8 923.4 806.1 688.7 571.3 454.0 336.6 219.3 101.9 0.0 0.0 0.0 0.0

51.9 51.9 51.9 51.9 51.9 51.9 51.9 51.9 51.9 51.9 51.9 51.9 51.9 45.1 0.0 0.0 0.0 0.0

24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65%

3.98% 3.98% 3.90% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98%

117.4 117.4 117.4 117.4 117.4 117.4 117.4 117.4 117.4 117.4 117.4 117.4 117.4 101.9 0.0 0.0 0.0 0.0

2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3



Pellets Tonnage

1.0

1.0

%

MI % %

MI

28.88%

208.0 67.50% 5.50%

208.0

0.00% 0.00% 0.00%

0.0 0.0 0.0

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.0 0.0 0.0

28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 0.00% 0.00% 0.00% 0.00%

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 13.0 0.0 0.0 0.0 0.0

67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 0.00% 0.00% 0.00% 0.00%

5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 0.00% 0.00% 0.00% 0.00%

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 13.0 11.0 0.0 0.0 0.0


Mining Complex Pipeline Pellet Plaid Port

1.0 MCD$ MCD$ MCD$ MCD$

1,166 365 681 252

380 380 380 121 121 121 227 227 227 83 83 83

25 1 1 2


MCD$ MCD$

2,464 586

812 812 812 195 195 195

0 0 0 0 0 0 0 0 0 29 0 0 0 0 0 0 0 0



1.0

MCD$

MCD$ MCDS MCD$ MCD$ MCD$ MCD$ MCD$

3,050

1,549 2,604 2,197

227 1,866

204 40

1,007 1,007 1,007

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0 0 0 0 0 0 0 29 0 0 0 0 0 0 0 0

112 112 112 112 112 112 112 112 112 112 112 112 112 97 0 0 0 0

188 188 188 188 188 188 188 188 188 188 188 188 188 163 0 0 0 0

158 158 158 158 158 158 158 158 158 158 158 158 158 138 0 0 0 0

16 16 16 16 16 16 16 16 16 16 16 16 16 14 0 0 0 0

135 135 135 135 135 135 135 135 135 135 135 135 135 117 0 0 0 0

15 15 15 15 15 15 15 15 15 15 15 15 15 13 0 0 0 0

3 3 3 3 3 3 3 3 3 3 3 3 3 2 0 0 0 0 Total Operating Costs

Revenue

Net Revenue

Cumulative Revenue

Discount Rate 1.0

MCD$

MCD$

MCDS

%

8,686

17,194

5,458

10%

0 0 0

0 0 0

-1,007 -1,007 -1,007

-1,007 -2,014 -3,021

626 626 626 626 626 626 626 626 626 626 626 626 626 544 0 0 0 0

1,240 1,240 1,240 1,240 1,240 1,240 1,240 1,240 1,240 1,240 1,240 1,240 1,240 1,076 0 0 0 0

613 613 613 613 613 613 613 613 613 584 613 613 613 533 0 0 0 0

-2,408 -1,794 -1,181 -567 46 660 1,273 1,887 2,500 3,085 3,698 4,312 4,915 5,458 5,458 5,458 5,458 5,458


MCD$ MCDS

% Tears

867 -

15% 5

-915 -1,748 -2,504 -2,085 -1,704 -1,358 -1,043 -757 -497 -260 -45 150 320 481 628 761 867 867 867 867 867

ET•CHEn1 P:\28078\Texte\Rapports\Opportunity Study Report\Final Report\28078 Opportunity Study Final Report-Rev.4.doc






Option 3 - 18 Mtpy

Pre-production Produchon Description Factor Units Total -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16

Mining Depletion


Waste Tonnage

Production Ore

Tonnage Fe% 0102%

Waste Tonnage

Stripping Ratio

Mt % %

Mt

Mt % %

Mt

720.3 24.65% 3.98%

1,627.6

720.3 24.65% 3.98%

1,627.6

2.3

720.3 720.3 720.3

24.65% 24.65% 24.65%

3.98% 3.98% 3.98%

1,627.6 1,627.6 1,627.6

0.0 0.0 0.0

0.0 0.0 0.0

0.0 0.0 0.0

0.0 0.0 0.0

720.3 658.0 595.6 533.3 471.0 408.7 346.3 284.0 221.7 159.4 97.0 34.7 0.0 0.0 0.0 0.0 0.0 0.0

24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65%

3.98% 3.98% 3.9t% 3.9t% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98%

1,627.6 1,486.8 1,345.9 1,205.1 1,064.3 923.4 782.6 641.8 500.9 360.1 219.3 78.4 0.0 0.0 0.0 0.0 0.0 0.0

623 62.3 62.3 62.3 62.3 62.3 62.3 62.3 62.3 62.3 62.3 34.7 0.0 0.0 0.0 0.0 0.0 0.0

24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65%

3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98%

140.8 140.8 140.8 140.8 140.8 140.8 140.8 140.8 140.8 140.8 140.8 78.4 0.0 0.0 0.0 0.0 0.0 0.0

2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3



Pellets Tonnage

1.0

1.0

%

Mt % %

Mt

28.88%

208.0 67.50%

5.50%

208.0

0.00% 0.00% 0.00%

0.0 0.0 0.0

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.0 0.0 0.0

28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00%

18.0 18.0 18.0 18.0 18.0 18.0 18.0 18.0 18.0 18.0 18.0 10.0 0.0 0.0 0.0 0.0 0.0 0.0

67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00%

5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00%

18.0 18.0 18.0 18.0 18.0 18.0 18.0 18.0 18.0 18.0 18.0 10.0 0.0 0.0 0.0 0.0 0.0 0.0


Mining Complex Pipeline Pellet Plant Port


1,331 456 762 320

434 434 434 152 152 152 254 254 254 106 106 106

30 1 1 2


MCD$ MCD$

2,869 683

945 945 945 228 228 228

0 0 11 0 0 0 0 0 0 34 0 0 0 0 0 0 0 0



1.0

MCD$

MCD$ MCD$ MCD$ MCD$ MCD$ MCD$ MCD$

3,552

1,549 2,604 2,197

227 1,866

204 40

1,173 1,173 1,173

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0 0 0 0 0 0 0 34 0 0 0 0 0 0 0 0

134 134 134 134 134 134 134 134 134 134 134 75 0 0 0 0 0 0

225 225 225 225 225 225 225 225 225 225 225 125 0 0 0 0 0 0

190 190 190 190 190 190 190 190 190 190 190 106 0 0 0 0 0 0

20 20 20 20 20 20 20 20 20 20 20 11 0 0 0 0 0 0

161 161 161 161 161 161 161 161 161 161 161 90 0 0 0 0 0 0

18 18 18 18 18 18 18 18 18 18 18 10 0 0 0 0 0 0


Revenue

Net Revenue

Cumulative Revenue

Discount Rate 1.0

MCD$

MCD$

MCD$

%

8,686

17,194

4,956

10%

0 0 0

0 0 0

1,173 -1,173 -1,173

-1,173 -2,345 -3,518

752 752 752 752 752 752 752 752 752 752 752 418 0 0 0 0 0 0

1,488 1,488 1,488 1,488 1,488 1,488 1,488 1,488 1,488 1,488 1,488 828 0 0 0 0 0 0

736 736 736 736 736 736 736 736 736 702 736 410 0 0 0 0 0 0

-2,782 -2,046 -1,309 -573 163 899 1,635 2,372 3,108 3,810 4,546 4,956 4,956 4,956 4,956 4,956 4,956 4,956


MCD$ MCD$

% Years

765 -

14% 5

-1,066 -2,035 -2,916 -2,413 -1,956 -1,541 -1,163 -820 -507 -223 35 269 473 666 765 765 765 765 765 765 765

ET•CHEn1 December 2010

QPF-009-12/B






Option 4 - 21 Mtpy

ro-prouenou ucnos

Descri tion Factor Units Total -2 -1 6 7 8 9

ro

10 ll 12 13 14 15 16 Mining Depiction


Waste Tonnage

Production Ore

Tonnage Fe% Si02%

Waste Tonnage

Stripping Ratio

Mt % %

Mt

Mt % %

Mt

7203 24.65% 3.98%

1,627.6

7203 24.65% 3.98%

1,627.6

23

720.3 720.3 720.3

24.65% 24.65% 24.65%

3.98% 3.98% 3.98%

1,627.6 1,627.6 1,627.6

0.0 0.0 0.0

0.0 0.0 0.0

0.0 0.0 0.0

0.0 0.0 0.0

720.3 647.6 574.9 502.2 429.4 356.7 284.0 211.3 138.6 65.9 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65%

3.98% 3.90% 3.98% 3.98% 3.98% 3.98% 3.90% 3.98% 3.98% 3.98%

1,627.6 1,463.3 1,299.0 1,134.7 970.4 806.1 641.8 477.5 313.1 148.8 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

72.7 72.7 72.7 72.7 72.7 72.7 72.7 72.7 72.7 65.9 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65% 24.65%

3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98% 3.98%

164.3 164.3 1643 164.3 164.3 164.3 164.3 164.3 164.3 148.8 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3 2.3



Pellets Tonnage

1.0

1.0

%

Mt % %

Mt

28.88%

208.0 67.50%

5.50%

208.0

0.00% 0.00% 0.00%

0.0 0.0 0.0

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.0 0.0 0.0

28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 28.88% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00%

21.0 21.0 21.0 21.0 21.0 21.0 21.0 21.0 21.0 19.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 67.50% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00%

5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 5.50% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00% 0.00%

21.0 21.0 21.0 21.0 21.0 21.0 21.0 21.0 21.0 19.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Economics Capital Costa

Mining Complex Pipeline Pellet Plant Pon


1,462 500 836 350

476 476 476 166 166 166 278 278 278 116 116 116

35 1 1 2


MCD$ MCD$

3,148 749

1,036 1,036 1,036 250 250 250

0 0 0 0 0 0 0 0 0 39 0 0 0 0 0 0 0 0


Operating Costs Mining Ore Mining Waste Processing Pipeline Pellet Plant Pon Handling Administration

1.0

MCDS

MCD$ MCD$ MCD$ MCD$ MCD$ MCD$ MCD$

3,897

1,549 2,604 2,197

227 1,866

204 40

1,286 1,286 1,286

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0

0 0 0 0 0 0 0 0 0 39 0 0 0 0 0 0 0 0

156 156 156 156 156 156 156 156 156 142 0 0 0 0 0 0 0 0

263 263 263 263 263 263 263 263 263 238 0 0 0 0 0 0 0 0

222 222 222 222 222 222 222 222 222 201 0 0 0 0 0 0 0 0

23 23 23 23 23 23 23 23 23 21 0 0 0 0 0 0 0 0

188 188 188 188 188 188 188 188 188 171 0 0 0 0 0 0 0 0

21 21 21 21 21 21 21 21 21 19 0 0 0 0 0 0 0 0


Revenue

Net Revenue

Cumulative Revenue

Discount Rate 1.0

MCD$

MCDS

MCD$

%

8,686

17,194

4,611

10%

0 0 0

0 0 0

-1,286 -1,286 -1,286

-1,286 -2,572 -3,858

877 877 877 877 877 877 877 877 877 794 0 0 0 0 0 0 0 0

1,736 1,736 1,736 1,736 1,736 1,736 1,736 1,736 1,736 1,572 0 0 0 0 0 0 0 0

859 859 859 859 859 859 859 859 859 739 0 0 0 0 0 0 0 0

-2,999 -2,140 -1,281 -422 436 1,295 2,154 3,013 3,872 4,611 4,611 4,611 4,611 4,611 4,611 4,611 4,611 4,611

NPV Cumulative NPV IRE Pay Back

MCD$ MCDS

% Years

732 -

14% 5

-1,169 -2,232 -3,198 -2,611 -2,078 -1,593 -1,153 -752 -388 -56 245 518 732 732 732 732 732 732 732 732 732

ET•CHEn1 December 2010

QPF-009-12/B



Appendix C — Planned Drill Locations

Drilling Proposal for Inferred Category

Deposit 4

Section

X

Coordinate

Y

Azimut Dip Lenght

(m)

DDHs 44

16000 328,350 5,940,840 320 -60 200 DUN10_ inf-01

16200 328,450 5,941,040 320 -60 200 DUN10_ inf-02

16400 328,541 5,941,250 320 -60 300 DUN10_ inf-03

16400 328,495 5,941,300 320 -60 150 DUN10_ inf-04

16600 328,625 5,941,455 320 -50 200 DUN10_ inf-05

16800 328,835 5,941,310 320 -50 250 DUN10_ inf-06

16900 329,010 5,941,468 320 -60 500 DUN10_ inf-07

16900 328,928 5,941,574 320 -60 300 DUN10_ inf-08

17100 329,210 5,941,541 320 -60 400 DUN10_ inf-09

17100 329,010 5,941,780 320 -60 350 DUN10_ inf-10

17300 329,180 5,941,887 320 -60 250 DUN10_ inf-11

17300 329,389 5,941,639 320 -60 350 DUN10_ inf-12

17300 329,438 5,941,581 320 -50 200 DUN10_ inf-13

17500 329,533 5,941,778 320 -60 350 DUN10_ inf-14

17500 329,360 5,941,985 320 -60 200 DUN10_ inf-15

17700 329,525 5,942,090 320 -60 200 DUN10_ inf-16

17700 329,682 5,941,912 320 -60 500 DUN10_ inf-17

17900 329,752 5,942,140 320 -60 200 DUN10_ inf-18

17900 329,833 5,942,043 320 -60 500 DUN10_ inf-19

18100 329,980 5,942,179 320 -60 300 DUN10_ inf-20

18100 330,064 5,942,079 320 -60 400 DUN10_ inf-21

18300 330,269 5,942,146 320 -60 350 DUN10_ inf-22

18500 330,378 5,942,327 320 -60 200 DUN10_ inf-23

18700 330,531 5,942,455 320 -60 200 DUN10_ inf-24

18900 330,614 5,942,668 320 -50 200 DUN10_ inf-25

17100 328,895 5,941,910 320 -60 250 DUN10_ inf-26

16900 328,845 5,941,670 320 -60 200 DUN10_ inf-27

Total

7700

27


QPF-009-12/B




Deposit 3

Section

X

Coordinate

Y

Azimut Dip Lenght

(m)

DDHs #

9200 323,955 5,935,588 315 -50 250 Dun10 Inf-28

9400 323,962 5,935,864 315 -50 200 Dun10_Inf-29

9400 324,085 5,935,742 315 -50 300 Dun10_Inf-30

9400 324,213 5,935,614 315 -50 400 Dun10_Inf-31

9600 324,135 5,935,974 315 -50 300 Dun10_Inf-32

9600 324,272 5,935,837 315 -50 400 Dun10_Inf-33

9800 324,015 5,936,377 315 -50 200 Dun10_Inf-34

9800 324,252 5,936,140 315 -50 400 Dun10 Inf-35

9800 324,438 5,935,956 315 -50 350 Dun10_Inf-36

10000 324,230 5,936,445 315 -50 450 Dun10 Inf-37

10000 324,520 5,936,155 315 -50 200 Dun10_Inf-38

10200 324,174 5,936,784 315 -50 300 Dun10_Inf-39

10200 324,472 5,936,485 315 -50 450 Dun10_Inf-40

10400 324,477 5,936,763 315 -50 450 Dun10_Inf-41

10400 324,946 5,936,294 315 -50 350 Dun10_Inf-42

10600 324,473 5,937,050 315 -50 250 Dun10_Inf-43

10600 324,679 5,936,844 315 -50 450 Dun10 Inf-44

10600 324,941 5,936,582 315 -50 250 Dun10_Inf-45

10600 325,147 5,936,376 315 -50 400 Dun10 Inf-46

10800 324,769 5,937,038 315 -50 450 Dun10_Inf-47

10800 325,007 5,936,800 315 -50 300 Dun10 Inf-48

10800 325,270 5,936,536 315 -50 450 Dun10_Inf-49

11000 324,709 5,937,380 315 -50 200 Dun10_Inf-50

11000 324,859 5,937,230 315 -50 450 Dun10_Inf-51

11000 325,419 5,936,670 315 -50 200 Dun10_Inf-52

11000 325,557 5,936,532 315 -50 450 Dun10_Inf-53

11200 324,878 5,937,494 315 -50 350 Dun10_Inf-54

11200 325,566 5,936,806 315 -50 250 Dun10 Inf-55

11200 325,773 5,936,599 315 -50 300 Dun10_Inf-56

11200 325,935 5,936,437 315 -50 450 Dun10 Inf-57

11400 324,902 5,937,753 315 -50 250 Dun10_Inf-58

11400 325,069 5,937,586 315 -50 450 Dun10_Inf-59

11400 325,548 5,937,107 315 -50 250 Dun10_Inf-60

11400 325,712 5,936,943 315 -50 450 Dun10_Inf-61

11500 324,988 5,937,808 315 -50 350 Dun10_Inf-62

11600 325,836 5,937,102 315 -50 450 Dun10_Inf-63

Total

12400

36

: mEr•cHEm December 2010

QPF-009-12/B




Deposit 6

Section

X

Coordinate

Y

Azimut Dip Lenght

(m)

DDHs #

100 326,523 5,935,215 340 -50 250 Dun10_Inf-64

300 326,350 5,936,274 340 -50 250 Dun10_Inf-65

300 326,452 5,935,995 340 -50 400 Dun10_Inf-66

300 326,708 5,935,290 340 -50 250 Dun10_Inf-67

500 326,521 5,936,389 340 -50 250 Dun10_Inf-68

500 326,627 5,936,097 340 -50 300 Dun10_Inf-69

500 326,695 5,935,911 340 -50 450 Dun10_Inf-70

500 326,777 5,935,686 340 -50 400 Dun10_Inf-71

700 326,758 5,936,323 340 -50 400 Dun10_Inf-72

700 326,815 5,936,165 340 -50 500 Dun10_Inf-73

900 326,942 5,936,400 340 -50 350 Dun10_Inf-74

900 327,012 5,936,209 340 -50 350 Dun10_Inf-75

900 327,086 5,936,007 340 -50 350 Dun10_Inf-76

1100 327,169 5,936,365 340 -50 400 Dun10_Inf-77

1100 327,239 5,936,171 340 -50 350 Dun10_Inf-78

1200 327,211 5,936,539 340 -50 250 Dun10_Inf-79

1300 327,363 5,936,415 340 -50 400 Dun10_Inf-80

1500 327,523 5,936,559 340 -50 350 Dun10_Inf-81

1700 327,685 5,936,700 340 -50 250 Dun10_Inf-82

1900 327,855 5,936,817 340 -50 200 Dun10_Inf-83

Total

6700

20


QPF-009-12/B



Drilling Proposal for Measured Category

Deposit 4

Section

X

Coordinate

Y

Azimut Dip Lenght

(m)

DDHs #

15900 328,227 5,940,846 320 -60 100 DUN10_ Mea-01

15900 328,275 5,940,789 320 -60 200 DUN10_ Mea-02

16000 328,300 5,940,914 320 -60 150 DUN10_ Mea-03

16000 328,412 5,940,780 320 -60 350 DUN10_ Mea-04

16200 328,388 5,941,121 320 -60 150 DUN10_ Mea-05

16200 328,519 5,940,965 320 -60 350 DUN10_ Mea-06

16400 328,606 5,941,172 320 -60 450 DUN10_ Mea-07

16600 328,738 5,941,326 320 -60 400 DUN10_ Mea-08

16800 328,887 5,941,459 320 -60 300 DUN10 Mea-09

16800 328,995 5,941,331 320 -60 250 DUN10_ Mea-10

16900 329,074 5,941,391 320 -50 450 DUN10 Mea-11

17100 328,920 5,941,887 320 -50 150 DUN10_ Mea-12

17100 329,106 5,941,675 320 -60 400 DUN10 Mea-13

17100 329,284 5,941,453 320 -60 400 DUN10_ Mea-14

17300 329,342 5,941,695 320 -60 300 DUN10 Mea-15

17300 329,144 5,941,930 320 -50 150 DUN10_ Mea-16

17300 329,252 5,941,802 320 -60 400 DUN10_ Mea-17

17300 329,438 5,941,581 320 -60 400 DUN10_ Mea-18

17500 329,324 5,942,028 320 -50 200 DUN10_ Mea-19

17500 329,401 5,941,936 320 -60 250 DUN10_ Mea-20

17500 329,482 5,941,839 320 -60 450 DUN10_ Mea-21

17500 329,582 5,941,720 320 -60 450 DUN10_ Mea-22

17700 329,581 5,942,032 320 -60 350 DUN10_ Mea-23

17700 329,637 5,941,965 320 -60 400 DUN10_ Mea-24

17700 329,723 5,941,863 320 -60 450 DUN10_ Mea-25

17900 329,790 5,942,094 320 -60 300 DUN10_ Mea-26

17900 329,872 5,941,997 320 -60 450 DUN10_ Mea-27

18300 330,203 5,942,225 320 -50 150 DUN10 Mea-28

18500 330,413 5,942,286 320 -60 250 DUN10_ Mea-29

18500 330,339 5,942,373 320 -60 150 DUN10 Mea-30

18700 330,581 5,942,396 320 -60 350 DUN10_ Mea-31

18900 330,677 5,942,593 320 -50 350 DUN10 Mea-32

Total

9900

32


QPF-009-12/B




Deposit 3

Section

X

Coordinate

Y

Azimut Dip Lenght

(m)

DDHs #

9200 323,905 5,935,638 315 -50 100 DUN10_ Mea-33

9200 324,002 5,935,542 315 -60 400 DUN10_ Mea-34

9400 324,027 5,935,799 315 -50 200 DUN10_ Mea-35

9400 324,140 5,935,870 315 -50 400 DUN10_ Mea-36

9600 324,071 5,936,038 315 -50 200 DUN10_ Mea-37

9600 324,195 5,935,915 315 -50 400 DUN10_ Mea-38

9600 324,352 5,935,758 315 -50 400 DUN10_ Mea-39

9800 324,127 5,936,265 315 -50 450 DUN10_ Mea-40

9800 324,517 5,935,875 315 -50 450 DUN10_ Mea-41

10000 324,112 5,936,563 315 -50 250 DUN10_ Mea-42

10000 324,358 5,936,317 315 -50 450 DUN10_ Mea-43

10000 324,702 5,936,177 315 -50 450 DUN10_ Mea-44

10200 324,332 5,936,626 315 -50 450 DUN10_ Mea-45

10200 324,693 5,936,355 315 -50 200 DUN10_ Mea-46

10200 324,844 5,936,317 315 -50 450 DUN10_ Mea-47

10400 324,376 5,936,864 315 -50 300 DUN10_ Mea-48

10400 324,565 5,936,675 315 -50 450 DUN10_ Mea-49

10400 324,858 5,936,382 315 -50 200 DUN10_ Mea-50

10400 325,021 5,936,220 315 -50 400 DUN10_ Mea-51

10600 324,569 5,936,954 315 -50 400 DUN10_ Mea-52

10600 325,055 5,936,468 315 -50 400 DUN10_ Mea-53

10800 324,618 5,937,189 315 -50 250 DUN10_ Mea-54

10800 325,179 5,936,627 315 -50 250 DUN10_ Mea-55

10800 325,356 5,936,450 315 -50 450 DUN10_ Mea-56

11000 324,774 5,937,315 315 -50 300 DUN10_ Mea-57

11000 325,484 5,936,605 315 -50 300 DUN10_ Mea-58

11200 324,806 5,937,566 315 -50 150 DUN10_ Mea-59

11200 324,957 5,937,414 315 -50 450 DUN10_ Mea-60

11200 325,680 5,936,692 315 -50 250 DUN10_ Mea-61

11200 325,857 5,936,515 315 -50 450 DUN10_ Mea-62

11200 326,015 5,936,357 315 -50 450 DUN10_ Mea-63

11400 324,981 5,937,674 315 -50 450 DUN10_ Mea-64

11400 325,478 5,937,177 315 -50 200 DUN10_ Mea-65

11400 325,622 5,937,033 315 -50 450 DUN10_ Mea-66

11500 324,909 5,937,887 315 -50 250 DUN10_ Mea-67

11600 325,755 5,937,183 315 -50 350 DUN10_ Mea-68

Total

12400

36


QPF-009-12/B




Deposit 6

Section

X

Coordinate

Y

Azimut Dip Lenght

(m)

DDHs #

300 326,307 5,936,393 340 -50 150 DUN10_ Mea-69

300 326,395 5,936,150 340 -50 300 DUN10_ Mea-70

300 326,677 5,935,377 340 -50 150 DUN10_ Mea-71

500 326,492 5,936,469 340 -50 150 DUN10_ Mea-72

500 326,550 5,936,309 340 -50 400 DUN10_ Mea-73

500 326,593 5,936,191 340 -50 150 DUN10_ Mea-74

500 326,657 5,936,015 340 -50 450 DUN10_ Mea-75

700 326,866 5,936,026 340 -50 350 DUN10_ Mea-76

700 326,898 5,935,938 340 -50 450 DUN10_ Mea-77

900 326,906 5,936,500 340 -50 200 DUN10_ Mea-78

900 326,978 5,936,302 340 -50 350 DUN10_ Mea-79

900 327,046 5,936,115 340 -50 350 DUN10_ Mea-80

1100 327,131 5,936,466 340 -50 250 DUN10_ Mea-81

1100 327,203 5,936,269 340 -50 350 DUN10_ Mea-82

1200 327,263 5,936,397 340 -50 350 DUN10_ Mea-83

1300 327,317 5,936,541 340 -50 250 DUN10_ Mea-84

1300 327,398 593,318 340 -50 450 DUN10_ Mea-85

1500 327,494 5,936,640 340 -50 250 DUN10_ Mea-86

1500 327,557 5,936,465 340 -50 350 DUN10_ Mea-87

1700 327,715 5,936,616 340 -50 350 DUN10_ Mea-88

1900 327,883 5,936,738 340 -50 350 DUN10_ Mea-89

Total

6400

21


QPF-009-12/B


-77°36' 328000

-77°35' -77°34' 330000 331000

-77°33' 329000

73265 9

Total Magnetic Field (nT)

66426.9 69113.6

S Ô S

ô~ S

S o

S o

~

AUGYVA MINING RESOURCES DUNCAN LAKE PROJECT, BLOCK 4

LOCATION OF PROPOSED INFERRED DDH

Joël Simard, P. Geo./Geoph.

~ ~

~

S S

79908 3 117592 8 93583.6 57812 5 643195 615937 56465.5

• BLOCK 4

°

329000 -77°35'

330000 -77°34'

331000 -77°33'

(meters) . NAD83 / UTM zone 18N 1-

328000 -77°36'

250 0 • 250 500 750

•

(7 80

< 8 78000 0

4, (7 ~00 4,

<76 87

4,I

( • (7>0

000 00 4,

<7,,<V004,

7 <7~os001 <7„,,,,,00,. 00 4,4,

( 10 7 <7) 0

7> 00 4, 6•0 ,1,

< • 7 (7)7 ~004, 4,

~04,

(776y ,000~(.04,

<7 s~ Op ti

00N1045

C76 < ~00 ~00 4, 4, <70

~00 <776`90

0,1, <7)7 4, ,

<. e oy,0s4Oti

4, 70 s~0ô 4, . < °o,~ . .

4, DUNK,

(<76, 64,4,

(76 7620 0

e <7 O00 4' `5. 4,

4,

(7 (lj °OO 400

(76 700 041,

uuNla.ixs

<76 • 9 OUNIUJ3°

0.10423


ET•CHEnI December 2010

QPF-009-12/B


324000 325000 326000 -77°40' -77°39' -77°38'

-77°40' -77°39' -77°38' -77°37' -77°36' 324000 325r000 326000 327,000 328000

I . I . ~ •

'v \ ~

1 i~

~ N~

j BLOCK 3 ;~4,~ ;

/ ~,° 1,\

~ s° " \ O ~ ,° °,~, � / ° ~~'~w

+ ! 1,

~ Co / tl, ~,$~

q1~ ~ / °~ 9 tq l ~Li°°W 1- 2

/ lo .w.

:BLOCK 6

.s ô a

•

�

ro l, °42 / lo°8i ti.

<'04,~~L °

ti- ••••

ti

ôs g -- m

{L225 S 300 5

L375 S • L4508

L600 S ° 1675 . L7505

1825 S L901 S

0975 S 11050 S

— 11125 L1200 S L1275 S 11350 S L1425 S L1300 S

01 CO

I 600 • • • • o- • • • 600 • • • 1080 • •â

+ °G — .~ ôc0, metersp ~

I ° NA08~ / UTM zone 18N â

3270▪ 00 I 328000

-77°37' -77°36' -77'35'

\ - ONER6li

\\\\IRPROl*166EC

l °$° L555 s • ~ 4. 1900 S

. . .~1 .°O° .•04, . L875$

J ti L1050 (0,8, °F • L1 S

(~ ~, • ,-~ü20~S ~ L1275 S F L1350 S

L1425 S

• \ l0

~ ~ ti

L1500 S

i~

~

8 0

~ N Ô


AUGYVA MINING RESOURCES DUNCAN LAKE PROJECT, BLOCKS 3 & 6

LOCATION OF PROPOSED INFERRED DDH

Total Magnetic Field ÇnT)

55815.0 57140.0 57895.0 58425.0 58895.0 59390.0 60015.0 60895.0 62225.0 64385.0 67775.0 73560.0 84710.0

Joël Simard, P. Geo./Gfi oph.


mET•cHEm P:\28078\Texte\Rapports\Opportunity Study Report\Final Report\28078 Opportunity Study Final Report-Rev.4.doc

opportunity study for the duncan lake iron ore project

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